DEVELOPMENT DOCUMENT FOR
EFFLUENT LIMITATIONS GUIDELINES
AND STANDARDS OF PERFORMANCE
FOR THE
ORE MINING AND DRESSING INDUSTRY
POINT SOURCE CATEGORY
£ £* \
| ^9&7 3
V
UNITED STATES ENVIRONMENTAL PROTECTION AGENCY
APRH 1975
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DRAFT
NOTICE
The attached document is a DRAFT CONTRACTOR'S REPORT. It includes tech-
nical information and recommendations submitted by the Contractor to the
United States Environmental Protection Agency ("EPA") regarding the sub-
ject industry. It is being distributed for review and comment only. The
report is not an official EPA publication, and it has not been reviewed
by the Agency.
The report, including the recommendations, will be undergoing extensive
review by EPA, Federal and State agencies, public interest organizations,
and other interested groups and persons during the coming weeks. The
report—and, in particular, the contractor's recommended effluent limita-
tion guidelines and standards of performance—is subject to change in any
and all respects.
The regulations to be published by EPA under Section 304 (b) and 306 of
the Federal Water Pollution Control Act, as amended, will be based to a
large extent on the report and the comments received on it. However,
pursuant to Sections 304 (b) and 306 of the Act, EPA will also consider
additional pertinent technical and economic information which is developed
in the course of review of this report by the public and within EPA. EPA
is currently performing an economic impact analysis regarding the subject
industry, which will be taken into account as part of the review of the
report. Upon completion of the review process, and prior to final pro-
mulgation of regulations, an EPA report will be issued setting forth EPA's
conclusions concerning the subject industry, effluent limitation guide-
lines, and standards of performance applicable to such industry. Judg-
ments necessary to promulgation of regulations under Sections 304 (b) and
306 of the Act, of course, remain the responsibility of EPA. Subject to
these limitations, EPA is making this draft contractor's report available
in order to encourage the widest possible participation of interested
persons in the decision making process at the earliest possible time.
The report shall have standing in any EPA proceeding or court proceeding
only to the extent that it represents the views of the Contractor who
studied the subject industry and prepared the information and recommenda-
tions. It cannot be cited, referenced, or represented in any respect in
any such proceedings as a statement of EPA's views regarding the subject
industry.
U.S. Environmental Protection Agency
Office of Water and Hazardous Materials
Effluent Guidelines Division
Washington, D. C. 20460
DRAFT
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ERRATA
Draft Development Document for Effluent Limitations Guidelines ana
Standards of Performance for the Ore Mining and Dressing Industry
Point Source Category dated April 1975.
Pg IX-12 Line 18, 1st word: change evaporation to precipitation.
Pg IX-13 Change 30-day average Hg waste load per unit ore milled
from 0.00035 kg/1000 metric tons to 0.0035 kg/1000
metric tons and from 0.0007 lb/1000 short tons to 0.007 lb/1000
short tons.
Pg IX-13 Change 24-hour maximum Ilg waste load per unit ore milled
from 0.0007 kg/1000 metric tons to 0.007 kg/1000 kg/1000 metric
tons and from 0.0015 lb/1000 short tons to 0.015 lb/1000
short tons.
Pg IX-21 Change 30-day average Zn concentration from 0.15 to 0.1.
Pg IX-21 Change 24-hour maximum Zn concentration from 0.25 to 0.2.
Pg IX-23 Change 30-day average Zn concentration from 0.15 to 0.1.
Pg IX-23 Change 24-hour maximum Zn concentration from 0.25 to 0.2.
Pg IX-23 Change 30-day average Zn waste load per unit ore milled
from 0.87 kg/1000 metric tons to 0.58 kg/1000 metric tons
and from 1.45 lb/1000 short tons to 1.16 lb/1000 short tons.
Pg IX-23 Change 24-hour maximum Zn waste load per unit ore milled
from 1.75 kg/1000 metric tons to 1.16 kg/1000 metric
tons and from 2.90 lb/1000 short tons to 2.32 lb/1000
short tons.
Pg IX-25 Change 30-day average Zn concentration from 0.15 to 0.1.
Pg IX-25 Change 24-hour maximum Zn concentration from 0.25 to 0.2.
Pg IX-25 Change 30-day average Zn waste load per unit ore milled
from 0.045 kg/1000 metric tons to 0.03 kg/1000 metric tons
and from 0.09 lb/1000 short tons to 0.06 lb/1000 short tons.
Pg IX-25 Change 24-hour maximum Zn waste load per unit ore milled
from 0.074 kg/1000 metric tons to 0.06 kg/1000 metric tons
and from 0.15 lb/1000 short tons to 0,12 lb/1000 short tons.
Pg IX-29 Change 30-day average Zn concentration from 0.15 to 0.1.
-------
Pg IX-29 Change 24-hour maximum Zn concentration from 0.25 to 0.2.
Pg IX-32 Change 30-day average Zn concentration from 0.15 to 0.1.
Pg IX-32 Change 24-hour maximum Zn concentration from 0.25 to 0.2.
Pg IX-32 Change 30-day average Zn waste load per unit ore milled
from 0.87 kg/1000 metric tons to 0.58 kg/1000 metric tons
and from 1.45 lb/1000 short tons to 1.16 lb/1000 short tons.
Pg IX-32 Change 24-hour maximum Zn waste load per unit ore milled
from 1.45 kg/1000 metric tons to 1.16 kg/1000 metric tons
and from 2.90 lb/1000 short tons to 2.32 lb/1000 short tons.
Pg I X-38 Change 24-hour maximum As concentration from 0.7 to 0.8.
Pg IX-38 Change 24-hour maximum Cd concentration from 0.07 to 0.1.
Pg IX-38 Change 30-day average Zn concentration from 0.15 to 0.1.
Pg IX-47 Change 30-day average As concentration from 0.5 to 0.4.
Pg IX-47 Change 24-hour maximum As concentration from 0.7 to 0.8.
Pg IX-47 Change 30-day average As waste load per unit ore milled
from 2.3 kg/1000 metric tons to 1.8 kg/1000 metric tons
and from 4.6 lb/1000 short tons to 3.7 lb/1000 short tons.
Pg IX-47 Change 24-hour maximum As waste load per unit ore milled
from 3.22 kg/1000 metric tons to 3.7 kg/1000 metric tons
and from 6.44 lb/1000 short tons to 7.4 lb/1000 short tons.
Pg X-13 Identification of BATEA for Gold Mines (alone) should
read: The best available technology economically
achievable is chemical (lime or sulfide) precipitation
with settling ponds. (Same as BPCTCA.)
Pg X-27 Change 30-day average As concentration from 0.5 to 0.4.
Pg X-27 Change 24-hour maximum As concentration from 0.7 to 0.8.
Pg X-27 Change 30-day average As waste load per unit ore milled
from 2.3 kg/1000 metric tons to 1.8 kg/1000 metric tons
and from 4.5 lb/1000 short tons to 3.7 lb/1000 short tons.
Pg X-27 Change 24-hour maximum As waste load per unit ore milled
from 3.2 kg/1000 metric tons to 3.7 kg/1000 metric tons
and from 6.4 lb/1000 short tons to 7.4 lb/1000 short tons.
-------
Pg X-32 Change 30-day average TSS concentration from 25 to 20.
Pg X-32 Change 30-day average TSS waste load per unit ore milled
from 0.36 kg/1000 metric tons to 0.3 kg/1000 metric tons
and from 0.72 lb/1000 short tons to 0.6 lb/1000 short tons.
Pg XI-5 Change 30-day average Cd concentration from 0.5 to 0.05.
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DRA^T
DEVELOPMENT DOCUMENT
for
EFFLUENT LIMITATIONS GUIDELINES
and
STANDARDS OF PERFORMANCE
ORE MINING AND DRESSING INDUSTRY
Prepared By:
Calspan Corporation
P. 0. Box 235
Buffalo, New York 14221
Contract No. 68-01-2682
April
DRAFT
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DRAFT
ABSTRACT
This document presents the findings of an extensive study
of the ore mining and dressing industry, for the purpose of
developing effluent limitations guidelines for existing point
sources and standards of performance and pretreatment standards
for new sources, to implement Sections 304, 306 and 307 of the
Federal Water Pollution Control Act, as amended (33 U.S.C. 1551,
1314, and 1316, 86 Stat. 816 et. seq.) (the "Act).
Effluent limitations guidelines contained herein set forth
the degree of effluent reduction attainable through the appJl-
cation of the best practicable control technology currently
available (BPCTCA) and the degree of effluent reduction attain-
able through the application of the best available technology
economically achievable (BATEA) which must be achieved by
existing point sources by July 1, 1977, and July 1, 1983,
respectively. The standards of performance and pretreatment
standards for new sources contained herein set forth the
degree of effluent reduction which is achievable through the
application of the best available demonstrated control tech-
nology, processes, operating methods, or other alternatives.
Based upon the application of the best practicable control
technology currently available, 16 of the 43 subcategories for
which separate limitations are proposed can be operated with no
discharge of process wastewater. With the best available
technology economically achievable, 22 of the 43 subcategories
for which separate limitations are proposed can be operated
with no discharge of process wastewater to navigable waters.
No discharge of process wastewater pollutants is also achiev-
able as a new source performance standard for 22 of the 43
subcategories.
Supporting data and rationale for development oC the proposed
effluent limitation guidelines and standards of performance
are contained in this report.
I I I
DRAFT
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DRAFT
CONTENTS
Section Page
I CONCLUSIONS L-l
II RECOMMENDATIONS I I-I
III INTRODUCTION Ill-l
PURPOSE AND AUTHORITY JIJ-I
SUMMARY OF METHODS USED FOR DEVELOPMENT OF
EFFLUENT LIMITATION GUIDELINES AND STANDARDS
OF TECHNOLOGY III-3
SUMMARY OF ORE-BENEFICIATION PROCESSES III-7
GENERAL DESCRIPTION OF INDUSTRY BY ORE CATEGORY II1-20
Iron Ore 111-20
^ Copper Ore 111-31
Lead and Zinc Ores 111-38
^-Gold Ore 111-50
^Silver Ores JII-54
Bauxite 111-60
'Ferroalloy Ores II[-61
Mercury Ores 111-85
Uranium, Radium, and Vanadium Ores II1-91
v. Metal Ores, Not Elsewhere Classified 1EI-1L8
IV INDUSTRY CATEGORIZATION 1V-J
INTRODUCTION IV-1
MINE IV-2
MILL IV-2
FACTORS INFLUENCING SELECTION OF SUBCATEGORIES IN
ALL METAL ORK CATEGORIES IV-)
DISCUSSION OK 1'KIMAKY FACTORS INFLUENCING
SUBCATEGORI/.ATLON BY ORK CATEGORY IV-9
Iron Ore IV-10
Copper On- IV-1A
Lead and X.inc Ori?s IV-13
Gold Ores IV-11)
Silver Ores JV-21
Bauxite Ores 1V-2J
DRAFT
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CONTENTS (cont.)
Section Page
IV (cont.) Ferroalloy Ores IV-23
Mercury Ores IV-26
Uranium, Radium, and Vanadium Ores IV-27
Metal Ores, Not Elsewhere Classified IV-32
SUMMARY OF RECOMMENDED SUBCATEGORIZAT10N IV-35
V WASTE CHARACTERIZATION V-l
INTRODUCTION V-l
SPECIFIC WATER USES IN ALL CATEGORIES V-3
Noncontact Cooling Water V-3
Wash Water V-4
Transport Water V-4
Scrubber Water V-4
Process and Product Consumed Water V-4
Miscellaneous Water V-4
PROCESS WASTE CHARACTERISTICS BY ORE CATEGORY V-5
Iron Ore V-5
Copper Ore V-21
Lead and Zinc. Ores V-69
Gold Ores V-78
Silver Ores V-87
Bauxite Ores V-102
Ferroalloy Ores V-lll
Mercury Ores V-137
Uranium, Radium, and Vanadium Ores V-149
Metal Ores - Not Elsewhere Classified V-173
Antimony V-l83
Beryllium V-185
Rare Earths V-188
Plat fnum-Groitp MeLals V-196
Tin V-J98
Titanium V-200
Zirconium V-207
VI SELECTION OK POLLUTANT I'AKAMKTKRS VI-1
INTRODUCTION VI-1
DRAFT
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DRAFT
vil
DRAFT
CONTENTS (cont.)
Section
VI GUIDELINE PARAMETER-SELECTION CRITERIA
(cont.)
SIGNIFICANCE AND RATIONALE FOR SELECTION OF
POLLUTION PARAMETERS VI-2
SIGNIFICANCE AND RATIONALE FOR REJECTJON OF
POLLUTION PARAMETERS VI-2'J
SUMMARY OF POLLUTION PARAMETERS SELECTED BY
CATEGORY VI-JI
VII CONTROL AND TREATMENT TECHNOLOGY Vll-l
INTRODUCTION VI1-2
CONTROL PRACTICES AND TECHNOLOGY VII-2
Mining Techniques VII-2
Surface-Water Control VII-6
Segregation or Combination of Mine and Mill
Wastewaters VII-7
Regrading VII-8
Erosion Control VII-10
Revegetatlon VII-12
Exploration, Development, and Pilot-Scale
Operations VI1-15
Mine and Mill Closure VI1-17
TREATMENT TECHNOLOGY VI L -1l)
Impoundment Systems VI1-20
Clarifiers and Thickeners V1I-23
Flocculation VII-25
Centrifugation VII-26
Hydrocyclones VII-26
Filtration V1I-27
Neutralization VI]-28
Chemical Free I pi tjt Ion Processes VT1-.10
ReducLlon VII-4J
Oxidation, Aeration, and Air Stripping Vll-AJ
Adsorption VIT-44
Ion Exchange V1I-40
UlLral I I tratlon and Reverse Osmosis VTl-S/i
High-Dens Ity-SLudge Acid Neutral Iz.iL Ion Vll-SS
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DRAFT
Section
VII (cont.)
VIII
CONTENTS (cont.)
Solvent Extraction VII-57
Evaporation and Distillation VII-57
Techniques for Reduction of Wastewater Volume VII-58
Electrodialysis VII-6]
Freezing VII-62
Biological Treatment VII-62
EXEMPLARY TREATMENT OPERATIONS BY ORE CATEGORY VII-64
Iron Ore VII-64
Copper Ores VII-69
Lead and Zinc Ores VII-84
Gold Ores VII-98
Silver Ores VII-109
Bauxite Ore VII-117
Ferroalloy Ores VII-121
Mercury Ores VII-148
Uranium, Radium, and Vanadium Ores VII-150
Metal Ores, Not Elsewhere Classified VII-159
Antimony Ores VII-159
Beryllium Ores VII-161
Platinum-Group Metals VII-161
Rare-Earth Ores VII-162
Tin Ores VII-164
Titanium Ores VII-164
Zirconium Ores VII-171
NONWATER-QUALITY ENVIRONMENTAL ASPECTS VII-171
COST, ENERGY, AND NONWATER-QUALITY ASPECTS VIII-1
INTRODUCTION VIII-1
SUMMARY OF METHODS USED VIII-2
Capital Costs VITI-2
Annual Costs VI LI-b
WASTEWATER-TREATMENT COSTS FOR IKON-OKK CATKCORY V1II-8
Iron-Ore Mines Vlll-S
Iron-Ore Mills Employing Chemlc.il/IMiys U:«l
Separation VIII-12
viJI
DRAFT
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DRAFT
CONTENTS (cont.)
Section
VIII WASTEWATER TREATMENT COSTS FOR COPPER-ORE CATEGORY VIII-17
(cont.) Copper Mines VIII-17
Copper Mills Using Froth Flocculation
(Precipitation Minus Evaporation = -76 cm
to Positive) (-30 in. to Positive) VlII-20
WASTEWATER-TREATMENT COSTS FOR LEAD- AND ZINC-ORE
CATEGORY VII1-25
Lead/Zinc Mines with No Solubility Potential VITI-25
Lead/Zinc Mines with Solubility Potential VIII-28
Lead/Zinc Mills VIIT-32
WASTEWATER-TREATMENT COSTS FOR GOLD-ORE CATEGORY VIII-38
Gold Mines (Alone) VIII-38
Gold Mills or Mine/Mills (Cyanidation Process) VIII-43
Gold Mills (Amalgamation Process) VIII-47
Gold Mills (Flotation) VIII-52
Gold Mine/Mills Employing Gravity Separation VIII-57
WASTEWATER-TREATMENT COSTS FOR SILVER-ORE CATEGORY VIII-62
Silver-Ore Mines VIII-62
Silver Mills Employing Cyanidation, Amalgama-
tion, Gravity Separation, and Byproduct
Recovery VIII-67
Silver Mines Employing Flotation Process VIII-68
WASTEWATER-TREATMENT COSTS FOR BAUXITE CATEGORY VII1-7L
Bauxite Mines VIII-7]
WASTEWATER-TREATMENT COSTS FOR FERROALLOY-ORE
CATEGORY
Ferroalloy-Ore Mines
Ferroalloy Mine/Mills Annually Processing
Less Than 5,000 Metric Tons (5,500
Short Tons) Ore by Methods Other Than
Ore Leaching
Ferroalloy Mills Annually Processing More
Than 5,000 Metric Tons (5,500 Short
Terns) Ore by 1'hyslcal Methods
Ferroalloy Mills Annually Processing More
Than 5.000 Metric Tons (5.500 Short
Tons) Ore by Flotation
Ferroalloy Mills Practicing Ore Leaching
VIII-7-S
vr.r I-TJ
VII1-7S
VLIT-83
VII1-87
VI1I-95
ix
DRAFT
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DRAFT
Section
VIII
(cont.)
IX
CONTENTS (cont.)
WASTEWATER TREATMENT COSTS FOR MERCURY-ORE CATEGORY
Mercury-Ore Mines
Mercury Mills Employing Flotation Process
Mercury Mills Employing Gravity Separation
WASTEWATER TREATMENT COSTS FOR URANIUM-ORE CATEGORY
Uranium Mines
Uranium Mills Using Acid or Combined Add/
Alkaline Leaching
Uranium Mills Using Alkaline Leaching
WASTEWATER TREATMENT COSTS FOR METAL ORES, NOT
ELSErfWERE CLASSIFIED
Antimony Mines
Titanium Mines
Titanium Mills Employing Electrostatic
and/or Magnetic Separation with Gravity
and/or Flotation Process
Platinum Mine/Mills Employing Dredging
BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY
AVAILABLE, GUIDELINES AND LIMITATIONS
INTRODUCTION
GENERAL WATER GUIDELINES
Process Water
Cooling Water
Storm-Water Runoff
BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY
AVAILABLE BY ORE CATEGORY AND SUBCATEGORY
Iron Ores
Copper Ores
Lead and ZJnc Ores
Gold Orc.s
Silver Orus
Bauxite Ores
Ferroalloy Ores
Mercury Ores
Uranium, Kudlum, ami Vaiim! lum Ore.::
VIII-103
VIII-103
VIII-107
V1II-112
V1II-115
VTII-115
VI11-125
VI IL-128
VIII-132
VIII-132
VIII-135
VIII-138
VIII-1A1
IX-1
IX-1
IX-2
IX-2
IX-2
IX-3
IX-4
IX-4
1X-8
IX-14
IX-18
IX-26
TX-J5
IX-J5
IX-4 6
1X-51
DRAFT
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DRAFT
CONTENTS (cont.)
Section Page
IX (cont.) Metal Ores, Not Elsewhere Classified IX-55
Antimony Ores IX-55
Beryllium Ores IX-58
Platinum Ores IX-59
Rare-Earth Ores JX-59
Tin Ores IX-62
Titanium Ores !X-d2
Zirconium Ores 1X-66
X BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE,
GUIDELINES AND LIMITATIONS X-]
INTRODUCTION X-J
GENERAL WATER GUIDELINES X-2
Process Water X-2
Cooling Water X-3
Storm-Water Runoff X-3
BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE,
BY ORE CATEGORY AND SUBCATEGORY X-5
Iron Ores X-5
Copper Ores x~7
Lead and Zinc Ores X-10
Gold Ores X-13
Silver Ores X-15
Bauxite Ores X-J 7
Ferroalloy Ores X-19
Mercury Ores X-26
Uranium, Radium, and Vanadium Ores X-28
Metal Ores, Not Elsewhere Classified X-29
Antimony Oros X-29
Beryllium Ores X-.ll
PLitinum Ores *--*'
Raro-Kartli Oros X-J3
Tin Orus X-JJ
Titanium Ores X-34
Zirconium Ores X-J5
xl
DRAFT
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DRAFT
Section
XI
XII
XIII
XIV
CONTENTS (cont.)
NEW SOURCE PERFORMANCE STANDARDS AND PRETREATMENT
STANDARDS
INTRODUCTION
GENERAL WATER GUIDELINES
NEW SOURCE STANDARDS BY ORE CATEGORY
Ferroalloy Ores
Uranium Ores
PRETREATMENT STANDARDS
ACKNOWLEDGMENTS
REFERENCES
GLOSSARY
CHEMICAL ELEMENTS
CONVERSION TABLE
Page
XI-1
XL-1
XI-2
X[-2
XI-4
XI-6
XI-9
XII-1
XIII-1
XIV-1
XIV-26
XIV-27
xlJ
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FIGURES
No. Title Page
III-l Beneficiation of Iron Ores 111-25
III-2 Iron-Ore Flotation-Circuit Flowsheet TTI-26
III-3 Magnetic Taconite Beneficiation Flowsheet IJ1-29
II1-4 Agglomeration Flowsheet FI1-30
III-5 Major Copper Mining and Milling Zones of the U.S. 111-32
III-6 General Outline of Methods for Typical Recovery
of Copper from Ore 111-37
III-7 Major Copper Areas Employing Acid Leaching in
Heaps, in Dumps, or In Situ 111-40
IH-fl Lead/Zinc-Ore Mining and Processing Operations 111-47
III-9 Cyanldation of Gold Ore: Vat Leaching of Sands
and "Carbon-in-Pulp1 Processing of Slimes 111-52
111-10 Cyanidation of Gold Ore; Agitation/Leach
Process 111-53
III-ll Flotation of Gold-Containing Minerals with
Recovery of Residual Gold Values by
Cyanidation I11-55
111-12 Recovery of Silver Sulfide Ore by Froth
Flotation 111-59
111-13 Gravity-Plant Flowsheet for Nigerian Columbite 111-72
111-14 Euxenite/Columbite Beneflciation-Plant Flowsheet 111-73
111-15 Representative Flow Sheet for Simple Gravity
Mill 111-74
111-16 Simplified Molybdenum Mill Flowsheet 111-77
111-17 Simplified Molybdenum Mill Flow Diagram 111-79
111-18 Simplified Flow Diagram for Small Tungsten
Concentrator 111-82
111-19 Mill Flowsheet for a Canadian Columbium
Operation 111-83
111-20 Flowsheet of Tristage Crystallization Process
for Recovery of Vanadium, Phosphorus, and
Chromium from Western Ferrophosphorus 111-86
HI-21 Arkansas Vanadium Process Flowsheet 111-87
111-22 Flowsheet of Dean-Leute Ammonium CarbamaLe
Process 111-88
IIL-23 Pachuca Tank for Alkaline Leaching 111-99
If1-24 Concentration Processes and Terminology 111-104
111-25 Simplified Schematic Diagram of Sulfurlc Acid
Digestion of Monazlte Sand for Recovery
of Thorium, Uranium, and Rare Earths III'-109
xlll
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DRAFT
FIGURES (cont.)
No. Title Page
111-26 Simplified Schematic Diagram of Caustic Soda
Digestion of Monazite Sand for Recovery
of Thorium, Uranium, and Rare Earths I1I-1LO
111-27 Effect of Acidity on Precipitation of Thorium,
Rare Earths and Uranium from a Monazite/
Sulfuric Acid Solution of Idaho and
Indian Monazite Sands Ill-Ill
111-28 Generalized Flow Diagram for Production of
Uranium, Vanadium, and Radium ILI-JIb
111-29 Beneficiation of Antimony Sulfide Ore by
Flotation JII-120
111-30 Gravity Concentration of Platinum-Group Metals III-124
HI-31 Beneficiation of Heavy-Mineral Beach Sands III-130
111-32 Beneficiation of Ilmenite Mined from a Rock Deposit III-131
V-l Flow Scheme for Treatment of Mine Water V-12
V-2 Water Flow Scheme in a Typical Milling Operation V-12
V-3 Water Balance for Mine/Mill 1105 (September 1974) V-14
V-4 Concentrator Flowsheet for Mill 1105 v-16
V-5 Flowsheet for Mill 1104 (Heavy-Media Plant) V-20
V-6 Simplified Concentration Flowsheet for Mine/Mill 1108 V-23
V-7 Wastewater Flowsheet for Plant 2120-B Pit V-28
V-8 Flowsheet of Hydrometallurgical Process Used in
Acid Leaching at Mine 2122 V-37
V-9 Reactions by Which Copper Minerals Are Dissolved In
Dump, Heap, or In-Situ Leaching v-38
V-10 Typical Design of Gravity Launder/Precipitation
Plant v-40
V-ll Cutaway Diagram of Cone Precipitator v-41
V-12 Diagram of Solvent Extraction Process for Recovery
of Copper by Leaching of Ore and Waste V-43
V-13 Vat Leach Flow Diagram (Mill 2124) V-4 7
V-14 Flow Diagram for Flotation of Copper (Mill 2120) V-S3
V-15 Addition of Flotation Agents to Modify Mineral
Surface V-54
V-16 Flowsheet for Miscellaneous Handling of Flotation
Tails (Mill 2124) V-6S
V-17 Dual Processing of Ore (Mill 2124) V-h(>
V-18 Leach/Pr€-clpItation/KLoLiit Ion Process V-dH
V-19 Water FJow Diagram for Mine 3105 V-71
V-20 Water Klow D Lag rum for Mint- 3104 V-7(.
V-21 Flow Diagram for Mill 'IJ01 V-7'l
xlv
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FIGURES (cont.)
— Title
V-22 Water Flow In Four Selected Gold Mining and
Milling Operations V-82
V-23 Water Flow in Silver Mines and Mills V-92
V-24 Process and Wastewater Flow Diagram for Open-Pit
Bauxite Mine 5101 V-104
V-25 Mill 6601 Flowsheet V-121
V-26 Simplified Mill Flow Diagram for Mill 6102 V-124
V-27 Internal Water Flow For Mill 6104 Through
Molybdenum Separation V-130
V-28 Internal Water Flow for Mill 6104 Following
Molybdenum Separation V-131
V-29 Water Use and Waste Sources for Vanadium Mill 6107 V-138
V-30 Water Flow in Mercury Mills 9101 and 9102 V-143
V-31 Typical Water-Use Patterns V-150
V-32 Alkaline-Leach Water Flow V-157
V-33 Ammonium Carbonate Leaching Process V-J58
V-34 Water Flow in Mills 9401, 9402, 9403, and 9404 V-161
V-35 Flowchart of Mill 9401 V-163
V-36 Flow Chart for Mill 9402 V-164
V-37 Flow Chart of Mill 9403 V-165
V-38 Flow Chart of Mill 9404 V-166
V-39 Water Flows and Usage for Mine/Mills 9901 (Antimony)
and 9902 (Beryllium) V-179
V-40 Water Flows and Usage for Mine/Mills 9903
(Rare Earths) and 9904 (Platinum) V-180
V-41 Water Flows and Usage for Titanium Mine/Mills
9905 and 9906 V-181
V-42 Beneficiation of Bertrandite, Mined from a Lode
Deposit by Flotation (Mill 9903) V-191
V-43 Beneficiation of Rare-Earth Flotation Concentrate
by Solvent Extraction (Mill 9903) V-192
V-44 Beneficiation and Waste Water Flow of IJmenite
Mine/Mill 9905 (Rock Deposit) V-202
V-45 Beneficiation of Heavy-Mineral Bcat-li Sands (RuLilc,
1Imenite, Zircon, and Monazltc) aL Mill 9906 V-J06
VII-1 Lime Neutralization and Prcrlplt.it Ion I'TOCL-HH for
Treatment o[ Mine Wat«.-r I'rlur to Dlscluir^i- VI I-12
VII-2 Theoretical Solubilities of MoLuJ.Lons .is .1
Function of pll VI I - H
xv
DRAFT
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DRAFT
FIGURES (cont.)
No. Title Page
VII-3 Minimum pH Value for Complete Precipitation
of Metal Ions as Hydroxides VII-34
VII-4 Heavy-Metal Precipitation vs pH for TaJ]ing-Pond
Effluent pH Adjustments by Lime Addition VI1-16
VII-5 Diagram of Modified Desal Process VI1-51
VII-6 Mill 1105 Water-Use System (Zero Discharge) VTI-68
VII-7 Control of Effluent by Reuse of Mine Water in
Leaching (Mine 2122) V1I-70
VII-8 Control of Mine-Water Effluent by Reuse in the
Concentrator (Mine/Mill 2119) VII-71
VII-9 Control of Effluent Through Reuse of Mill Flotation-
Process Water in Other Facilities
(Mine/Mill 2124) VII-78
VII-10 Reduction in Waste Pollutant Load in Discharge
by Separation of Minewater From Tailing Pond
for Separate Treatment (Mill 2121) VII-80
VII-11 Schematic Diagram of Treatment Facilities at
Mine 3107 VII-89
VII-12 Schematic Diagram of Water Flows and Treatment
Facilities at Mill 3103 VII-92
VII-13 Schematic Diagram of Water Flow and Treatment
Facilities at Mill 3102 (Tailing Pond/Stilling
Pond/Biological Treatment/Polishing Pond) VII-95
VII-14 Schematic Diagram of Water Flow and Treatment
Facilities at Mill 3105 VII-97
VII-15 Schematic Diagram of Treatment Facilities at
Mill 3101 VII-99
VII-16 Lime-Neutralization Plant for Open-Pit Mine 5102 VII-119
VII-17 Water-Flow Schematic Diagram for Mill 6102 VII-133
VII-18 Ion Exchange for Mercury and Uranium at Low
Loadings and Concentrations VII-154
VII-19 Chemical Changes in a Sequence of Tailing
Impoundments at Mill 9402 VII-158
xv i
DRAFT
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DRAFT
No.
Page
II-l Summary of Recommended BPCTCA Effluent Limitations
By Category and Subcategory—Ores for Which
Separate Limitations Are Proposed J1-2
II-2 Summary of Recommended BATEA Effluent Limitations
By Category and Subcategory—Ores for Which
Separate Limitations Are Proposed 11-4
II-3 Summary of Recommended NSPS Effluent Limitations
By Category and Subcategory—Ores for Which
Separate Limitations Are Proposed [1-6
III-l Iron-Ore Shipments for United States 111-21
III-2 Crude Iron-Ore Production for U.S. 111-22
III-3 Reagents Used for Flotation of Iron Ores 111-27
III-4 Various Flotation Methods Available for Pro-
duction of High-Grade Iron-Ore Concentrates 111-28
III-5 Total Copper-Mine Production of Ore by Year 111-33
III-6 Copper-Ore Production from Mines by State (1972) 111-33
III-7 Average Copper Content of Domestic Ore 111-35
III-8 Average Concentration of Copper in Domestic Ores
by Process (1972) 111-35
III-9 Copper Ore Concentrated in the United States
by Froth Flotation, Including LPF Process
(1972) IH-36
111-10 Heap or Vat Ore Leached in the United States (1972) 111-39
III-ll Average Price Received from Copper in the
United States III-A1
111-12 Production of Copper from Domestic Ore by
Smelters 111-42
111-13 Mine Production of Recoverable Lead in the
United States 111-44
111-14 Mine Production of Recoverable Zinc in the
United States (Preliminary) 111-45
111-15 Domestic Silver Production from Different
Types of Ores 111-57
111-16 Silver Produced at Amalgamation and Cyanldation
Mills in the U.S. and Percentage of Silver
Recoverable from All Sources 111-58
111-17 Production of Bauxite in the United States IH-62
111-18 Production of Ferroalloys by U. S. Mining and
Milling Industry 111-64
111-19 Observed Usage of Some Flotation Reagents 111-76
111-20 Probable Reagents Used in Flotation of Nickel
and Cobalt Ores 111-80
xvii
DRAFT
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DRAFT
TABLES (cont.)
No. Title Page
111-21 Domestic Mercury Production Statistics 111-89
111-22 Isotopic Abundance of Uranium ITI-93
111-23 Uranium Milling Activity by State, 1972 ITI-97
111-24 Uranium Concentration in IX/SX Eluates ITI-103
111-25 Decay Series of Thorium and Uranium 111-113
111-26 Uranium Milling Processes III-114
111-27 Uranium Production III-117
111-28 Vanadium Production 111-117
111-29 Vanadium Use III-117
111-30 Production of Antimony from Domestic Sources III-119
111-31 Domestic Platinum-Group Mine Production aiul Value I FT-123
111-32 Production and Mine Shipments of Titanium
Concentrates from Domestic Ores in the U.S. 111-128
IV-1 Summary of Industry Subcategorlzatlon Recommended fV-36
V-l Historical Constituents of Iron-Mine Discharges V-9
V-2 Historical Constituents of Wastewater from Iron-
Ore Processing V-9
V-3 Chemical Compositions of Sampled Mine Waters V-10
V-4 Chemical Compositions of Sampled Mill Waters V-10
V-5 Chemical Analysis of Discharge 1 (Mine Water)
and Discharge 2 (Mine and Mill Water) at
Mine/Mill 1104, Including Waste Loading
for Discharge 2 V-18
V-6 Chemical Characteristics of Discharge Water
from Mine 1108 V-22
V-7 Characteristics of Mill 1108 Discharge Water V-24
V-8 Principal Copper Minerals Used in the United States V-26
V-9 Mine-Water Production from Selected Major Copper-
Producing Mines and Fate(s) of Effluent V-29
V-10 Summary of Solid Wastes Produced by Plants
Surveyed V-30
V-ll Raw Waste Load In Water Pumped from Selected
Copper Mines V-.12
V-12 1973 Water Usage Ln Dump, Heap, and Tn-SLtu
Leaching Operations V-4S
V-13 Chemical Character!stJos of Barren Heap, Dump, or
In-Situ Acid Leach Solutions (Recycled: No
Waste Load) V-4h
V-14 Water Usage In Vat Leaching Process as a Function
of Amount of Product (Precipitate or Cdtliode
Copper) Produced V-49
V-15 Chemical Character 1stIcs of Vat-Leach Barren
Acid Solution (Recycled: No Waste Loud) V-50
xvlll
DRAFT
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DRAFT
TABLES (cont.)
No. Title Page
V-16 Miscellaneous Wastes from Special Handling uf
Ore Wash Slimes in Mine 2124 (No Effluent) V-51
V-17 Examples of Chemical Agents Which May be Employed
In Copper Flotation V-55
V-18 Water Usage in Froth Flotation of Copper V-58
V-19 Raw Mill Waste Loads Prior to Settling in Tailing
Ponds V-59
V-20 Wastewater Constituents and Waste Loads Resulting
from Discharge of Mill Process Waters V-63
V-21 Range of Chemical Characteristics of Sampled Raw
Mine Water from Lead/Zinc Mines 3102, 3103,
and 3104 V-72
V-22 Range of Chemical Characteristics of Raw Mine
Waters from Four Operations in SolubiJiza-
tion-Potential Subcategory V-77
V-23 Ranges of Constituents of Wastewaters and Raw Waste
Loads for Mills 3102, 3103, 3104, 3105, and
3106 V-80
V-24 Chemical Composition of Raw Mine Water from Mines
4105 and 4102 V-84
V-25 Process Reagent Use at Various Mills Beneficiating
Gold Ore V-88
V-26 Minerals Commonly Associated with Gold Ore V-88
V-27 Waste Characteristics and Raw Waste Loads at Four
Gold Milling Operations V-89
V-28 Raw Waste Characteristics of Silver Mining
Operations V-95
V-29 Major Minerals Found Associated with Silver Ores V-97
V-30 Flotation Reagents Used by Three Mills to Bene-
ficiate Silver-Containing Mineral Tetrahedrite
(Mills 4401 and 4403) and Native Silver and
Argentite (Mill 4402) V-99
V-31 Waste Characteristics and Raw Waste Loads at Mills
4401, 4402, 4403, and 4105 V-IOO
V-32 Concentrations of Selected Constituents in Acid
Raw Mine Drainage from Open-Pit Mine 5101 V-J07
V-33 Concentrations of Selected Constituents in Acid
Raw Mine Drainage from Open-Pit Mine 5J02 V-J07
V-34 Concentrations of Selected Constituents In Alkaline
Raw Mine Drainage from Underground Mine 5101 V-108
V-35 Wastewater and Raw Waste Load for Open-Pit Mine 5101 V-110
V-36 Wastewater and Raw Waste l.ot-ul for Underground
Mine 5101 V-110
V-37 Types of Operations Visited and Anl ic Ipaf eel —
Ferroal Joy-Ore Mini UK "»(1 Dressing Industry V-112
xlx
DRAFT
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DRAFT
TABLES (cont.)
No. Title page
V-38 Chemical Characteristics of Raw Mine Water in
Ferroalloy Industry V-116
V-39 Reagent Use in Molybdenum Mill 6101 V-122
V-AO Raw Waste Characterization and Raw Waste Load
for Mill 6601 V-122
V-41 Reagent Use for Rougher and Scavenger Flotation
at Mill 6102 V-J25
V-42 Reagent Use for Cleaner Flotation at Mill 6102 V-125
V-43 Reagent Use at Byproduct Plant of Mill 6102 (Based
on Total Byproduct Plant Feed) V-127
V-44 Mill 6102 Effluent Chemical Characteristics (Com-
bined-Tailings Sample) V-127
V-45 Chemical Characteristics of Acid-Flotation Step V-128
V-46 Composite Waste Characteristicis for Beneficiation
at Mill 6104 (Samples 6, 8, 9, and 11) V-132
V-47 Waste Characteristics from Copper-Thickener Over-
flow for Mill 6104 (Sample 5) V-132
V-48 Scheelite-Flotation Tailing Waste Characteristics
and Loading for Mill 6104 (Sample 7) V-133
V-49 50-Foot-Thickener Overflow for Mill 6104 (Sample 10) V-133
V-50 Waste Characteristics of Combined-Tailing Discharge
for Mill 6104 (Samples 15, 16, and 17) V-134
V-51 Waste Characteristics and Raw Waste Load at Mill
6105 (Sample 19) V-136
V-52 Chemical Composition of Wastewater, Total Waste,
and Raw Waste Loading from Milling and Smelter
Effluent for Mill 6106 V-136
V-53 Waste Characterization and Raw Waste Load for
Mill 6107 Leach and Solvent-Extraction Effluent
(Sample 80) V-139
V-54 Waste Characteristics and Waste Load for Dryer
Scrubber Bleed at Mill 6107 (Sample 81) V-140
V-55 Waste Characteristics and Loading for Salt-Roast
Scrubber Bleed at Mil] 6107 (Sample 77) V-I4I
V-56 Expected Reagent Use .it Mercury-Ore Flotation
Mill 9202 V-147
V-57 Waste Characteristics and Knw Waste Loadings at
Mills 9201 and 9202 V-148
V-58 Waste Constituents Expected V-153
V-59 Chemical and PliysicaJ Waste Constituents Observed
in Representative Operations V-L54
DRAFT
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DRAFT
TABLES (cont.)
No. Title Page
V-60 Water Use and Flows at Mine/Mills 9401, 9402, 9403,
and 9404 V-162
V-61 Water Treatment Involved In U/Ra/V Operations V-162
V-62 Radionuclides in Raw Waste-waters from Uranium/
Radium/Vanadium Mines and Mills V-169
V-63 Organic Constituents in U/Ra/V Raw Wastewater V-169
V-64 Inorganic Anions in U/Ra/V Raw Wastewater V-171
V-65 Light-Metal Concentrations Observed in U/Ra/V
Raw Wastewater V-171
V-66 Concentrations of Heavy Metals Forming Anionic
Species in U/Ra/V Raw Wastewater V-171
V-67 Concentrations of Heavy Metals Forming Cationic
Species in U/Ra/V Raw Wastewater V-172
V-68 Other Constituents Present in Raw Wastewater in
U/Ra/V Mines and Mills V-172
V-69 Chemical Composition of Wastewater and Raw Waste
Load for Uranium Mines 9401 and 9402 V-174
V-70 Chemical Composition of Raw Wastewater and Raw
Waste Load for Mill 9401 (Alkaline-Mill
Subcategory) V-174
V-71 Chemical Composition of Wastewater and Raw Waste
Load for Mill 9402 (Acid- or Combined Acid/
Alkaline-Mill Subcategory) V-175
V-72 Chemical Composition of Wastewater and Raw Waste
Load for Mine 9403 (Alkaline-Mill Subcategory) V-176
V-73 Chemical Composition of Wastewater and Raw Waste
Load for Mill 9404 (Acid- or Combined Acid/
Alkaline-Mill Subcategory) V-177
V-74 Reagent Use at Antimony-Ore Flotation Mill 9901 V-184
V-75 Chemical Composition of Raw Wastewater Discharged
From Antimony Flotation Mill 9901 V-186
V-76 Major Waste Constituents and Raw Waste Load at
Antimony Mill 9901 V-187
V-77 Chemical Composition of Raw Wastewater from
Beryllium Mill 9902 (No Discharge from
Treatment) V-189
V-78 Chemical Composition of Raw Wastewater from
Rare-Earth Mill 9903 V-194
V-79 Results of Chemical Analysis for Rare-Earth
Metals (Mill 9903—No Discharge) V-195
xxl
DRAFT
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DRAFT
TABLES (cont.)
No. Title page
V-80 Chemical Composition and Raw Waste Load from
Rare-Earth Mill 9903 V-197
V-81 Chemical Composition and Loading for Principal
Waste Constituents Resulting from Platinum
Mine/Mill 9904 (Industry Data) V-199
V-82 Chemical Composition of Raw Wastewater from
Titanium Mine 9905 V-201
V-83 Chemical Composition of Raw Wastewater From
Titanium Mill 9905 V-204
V-84 Reagent Use in Flotation Circuit of Mill 9905 V-204
V-85 Principal Minerals Associated with Ore of Mine 9905 V-205
V-86 Major Waste Constituents and Raw Waste Load at
Mill 9905 V-205
V-87 Chemical Composition of Raw Wastewater at Mills
9906 and 9907 V-208
V-88 Raw Waste Loads for Principal Wastewater Consti-
tuents from Sand Placer Mills 9906 and 9907 V-209
VI-1 Known Toxicity of Some Common Flotation Reagents
Used in Ore Mining and Milling Industry VI-27
VI-2 Summary of Parameters Selected for Effluent Limi-
tation by Metal Category VI-32
VII-1 Results of Coprecipitation Removal of Radium
from Wastewater VII-40
VII-2 Properties of Ion Exchangers for Metallurgical
Applications V1I-48
VII-3 Analytical Data for Modified Desal Process VIT-53
VII-4 Rejection of Metal Salts by Reverse-Osmosis
Membranes VTJ-56
VII-5 Chemical Characteristics of Settling-Pond Dis-
charge at Mine 1105 VTI-65
VII-6 Chemical Compositions of Raw and Treated Waste-
loading at Mine/Mill 1109 VTI-67
VII-7 Concentration of Parameters I'rosi>nf Ln Raw
Wastewater iind Effluent Following Lime
Precipitation nL Mine 212011 VI1-72
VII-8 Concentration of I'araineters I'ro.soiit In Knw W.isLe-
water and Effluent Following Lime I'reciplta-
tion at Mine 2120C VI1-73
VII-9 Dump, Heap, and In-SJ tu l.cdrh-Solut Inn Control
and Treatment Pr«.ii_tii-i- (I97J) VI 1-75
xx I i
DRAFT
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DRAFT
TABLES (cont.)
No- Title Page
VII-10 Solution-Control Practice in Vat Leaching of
Copper Ore VII-77
VII-11 Reduction of Pollutants in Concentrator Tails
by Settling at Various pH Levels VII-82
VII-12 Efficiency of Coagulation Treatment to Reduce
Pollutant Loads in Combined Waste (Includ-
ing Mill Waste) Prior to Discharge (Pilot
Plant) VII-83
VII-13 Chemical Compositions of Raw and Treated Mine-
waters from Mine 3105 (Historical Data Pre-
sented for Comparison) VII-85
VII-14 Chemical Compositions of Raw and Treated Waste-
waters from Mine 3107 (Historical Data Pre-
sented for Comparison) VII-87
VII-15 Chemical Compositions of Raw and Treated Mine
Waters from Mine 3101 VII-90
VII-16 Chemical Compositions and Waste Loads for Raw and
Treated Mill Wastewaters at Mill 3103 VII-93
VII-17 Chemical Composition and Waste Loading for Raw
and Treated Mill Wastewater Mill 3102 VII-96
VII-18 Waste Compositions and Raw and Treated Waste Loads
Achieved at Mill 4102 by Tailing-Pond Treat-
ment VII-103
VII-19 Chemical Compositions of Mill Wastewater and
Tailing-Pond Decant Water at Mill 4101 (No
Resultant Discharge) VII-105
VII-20 Waste Compositions and Raw and Treated Waste
Loads at Mill 4401 (Using Tailing-Pond
Treatment and Partial Recycle) VII-113
VII-21 Chemical Compositions of Mill Raw Wastewater
and Tailing-Pond Decant Water at Mill 4402 V1I-116
VII-22 Chemical Compositions of Raw and Treated Mine
Waters at Mine 5101 VII-120
VII-23 Chemical Compositions of Raw and Treated Mine
Waters at Mine 5102 VII-122
VII-24 Chemical Compositions of Raw Mine Wastewater
and Treated Effluent at Mine 6103 VII-124
VII-25 Chemical Compositions of R.iw and Treated Mine
Waters at Mine 6104 (Cl.irl riorcuLatnr
Treatment) VII-L26
VII-26 Chemical Compositions of KMW and Treated Waste-
waters at Mine 6L07 VII-127
xxlil
DRAFT
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DRAFT
TABLES (cont.)
No. Title Page
VII-27 Analyses of Intake and Discharge Waters From Mill
6101 (Company Data) VIT-130
V1I-28 Chemical Composition of WasLt-wa tor ami W;iste
Loading for Kill 6101 Vll-lll
VII-29 Chemical Composition and Calculated Waste Load for
Mill 6102 Tailing-Pond Surface Water, with
Analytical Data for Mi11-Reservoir Water VII-135
VII-30 Chemical Composition and Waste Loading for Discharge
at Mill 6102 (Company Data) VII-135
VII-31 Chemical Composition and Treated Waste Loads for
Overflow from First Settling Pond at Mill 6106 VII-138
VII-32 Characteristics of Surface Water from Second Settling
Pond at Mill 6106 VII-138
VII-33 Chemical Composition and Treated Waste Loads from
Final Effluent for Mine/Mill 6106 During
Rainy Season (Company Data) VII-139
VII-34 Chemical Composition ami Waste Loading from Area
Runoff and Reclamation-Pond Seepage at Mill
6107 (Company Data) VII-139
VII-35 Chemical Composition and Waste Loading for Cooling
Water Effluent at Mill 6107 (Company Data) VIi-140
VII-36 Chemical Composition and Waste Loading for Process
Effluent After AmmonLa Treatment at Mill 6107 VII-142
VII-37 Chemical Composition and Waste Loading for Drier
Scrubber Bleed Water After Settling Treatment
at Mill 6107 VII-143
VII-38 Chemical Composition and Waste Loading for Holding-
Pond Effluent (Process Water and Drier Scrubber
Bleed) at Mill 6107 (Company Data) VII-144
VII-39 Chemical Composition and Waste Loading for Roaster
Scrubber Bleed Water After Settling at Mill
6107 VII-145
VII-40 Chemical Composition and Waste Loading for Roaster
Scrubber Bleed Water After Settling .it Mill
6107 (Company Data) VI1-146
V1I-41 Chemical Composition an
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DRAFT
TABLES (cont.)
No. Title Page
VII-43 Chemical Compositions of Raw and Treated
Wastewaters at Mine 9402 (001) VII-154
VII-44 Chemical Compositions of Raw and Treated
Wastewaters at Mine 9402 (002) VII-155
VII-45 Chemical Compositions of Raw and Treated Waste-
waters and Effluent Waste Loading at Mill 9403
(Settling and BaC12^ Coprecipitation) Vfl-160
VII-46 Chemical Composition of Treated Effluent and
Waste Load from Mine/Mill 9904 (PJatinum) VTL-163
VII-47 Chemical Compositions of Raw Wastewatcr and Treated
Recycle Water at Mill 9903 (No Discharge) VII-163
VII-48 Chemical Compositions of Raw Wastewater and
Treated Recycle Water at Mill 9905 VII-165
VII-49 Chemical Compositions of Raw and Treated
Wastewaters at Mill 9906 VII-167
VII-50 Chemical Compositions of Raw and Treated
Wastewaters at Mill 9907 VII-168
VII-51 Wastewater Composition and Treated Waste Load
With Acid Flocculation and Settling at
Mill 9906 VII-169
VI1-52 Wastewater Composition and Treated Waste Load
With Acid Flocculation and Settling at
Mill 9907 VII-170
VIII-1 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mine 1105 VIII-9
VIII-2 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 1107 VIII-13
VIII-3 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mine 2120 VII1-18
VIII-4 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 2121 VIII-21
VIII-5 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mine (Hypothet Jciil)—l-cad/Zinc, No Solubility V11I-26
VIII-6 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mine (Hypotheticdl)--Lead/Zlnc, Solubility VIII-29
VIII-7 Water Effluent Treatment Costs and Resulting
Waste-Loud Characteristics For Typical
Mill (Hypothetical)—J.c.id/Zinc VllI-3/i
VIII-8 Water Effluent Treatment COHLS and Result Ing
Waste-Load Character 1 sites fur Typical
Mine (Hypothetical )—flo Id V1II-J9
XXV
DRAFT
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DRAFT
TABLES (cont.)
No. Title Page
VIII-9 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mil] 4105 VI FT-44
VIII-10 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 4102 V11F-48
VIII-11 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristirs for Mill 4104 V1JI-5J
VIII-12 Water Effluent Treatment Costs and Kt-sultiug
Waste-Load Characteristics fur Typical
Mine/Mill (Hypothc-t Leal) — Gold/Gravity VI11-58
VIII-13 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mine (Hypothetical)—Silver VIII-63
VIII-1A Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 4401 VIII-69
VIII-15 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mine 5102 VIII-72
VIII-16 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mine (Hypothetical)—Ferroalloy VIII-76
VIII-17 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mine/Mill (Hypothetical)— Ferroalloy/Limited VUI-79
VIII-18 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mill (Hypothetical)—Ferroalloy/Physical VII1-84
VIII-19 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristies for Typical
Mill (Hypothetical)— Fcrroalloy/F] oLation VI1I-88
VIII-20 Water Effluent Treatment Costs and ResuJting
Waste-Load Characteristics for Typical
Mill (Hypothetical)--Ferroalloy/Laaching VIII-96
VIII-21 Water Effluent Treatment Costs and Resulting
Waste-Load CharacteristIvs for Typlval
Mine (Hypothetical ) — Mc.n-ury VL 11-104
VIIT-22 Water Ef fluent Trt-.H iiii-nt (Justs .ind Rot.ii I L Inn
W.isU-l.oml Ui.ira. u-rlsLl. s for Mill 4J02 VI1I-HW
Vm-23 Water Kl'l'luunl TriMlmeiH Cn--,!-. ,md Kvsullinn
W;i.sti—Lo.itl Cli.ir.un>-risL U's lurMIII ').U)I VT11-.N3
VI 11-24 Wntor lifllutMit Tivnt-iniMH d'KLs .nul RosnUlnj;
WasLc-Lo.id ClULMcLoi i ht fi:^ Cur Typical
Mine (llypotlu-Llc.il)—ULMII iinn VUI-LIG
xxv I
DRAFT
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TABLES (cont.)
No. Title Page
VIII-25 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 9405 VIII-126
VIII-26 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 9403 VI1I-130
VIII-27 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Typical
Mine (Hypothetical)—Antimony VIII-133
VIII-28 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mine 9905 VIII-136
VIII-29 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mill 9905 VIII-139
VIII-30 Water Effluent Treatment Costs and Resulting
Waste-Load Characteristics for Mine/Mill 9904 VIII-142
IX-1 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Iron-Ore Mines TX-5
LX-2 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Iron-Ore Mills
Employing Physical Methods and/or
Chemical Reagents IX-7
IX-3 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Copper Mines IX-10
IX-4 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Copper Mills Using
Froth Flotation (Net Evaporation Less Than
76.2 cm (30 in.) per year) IX-13
IX-5 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Lead and Zinc
Mines Having No Solubilization Potential IX-15
IX-6 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Lead and Zinc Mines
Having Solubilization Potential IX-17
IX-7 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Lead and/or Zinc Mills IX-19
IX-8 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Gold Mines IX-21
IX-9 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Gold Mines Using
Amalgamation Process IX-23
IX-10 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Gold Mills Using
Flotation Process IX-25
xxv I i
DRAFT
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DRAFT
TABLES (cont.)
No. Title page
IX-11 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Gold Mines or Mills
Using Gravity-Separation Methods 1X-27
IX-12 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Silver Mines (Alone) IX-29
IX-13 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Silver Mills Using
Amalgamation Process IX-32
IX-14 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Silver Mills Using
Gravity Separation IX-34
IX-15 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Bauxite Mines (Acid
or Alkaline Mine Drainage) IX-36
IX-16 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Ferroalloy-Ore Mines IX-38
IX-17 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Ferroalloy-Ore Mills
Treating Less Than 5,000 Metric Tons (5,512
Short Tons) Per Year IX-40
IX-18 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Ferroalloy-Ore Mills
Treating More Than 5,000 Metric Tons (5,512
Short Tons) Per Year by Physical Processing IX-43
IX-19 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Ferroalloy-Ore Mills
Using Flotation Process IX-45
IX-20 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Ferroalloy-Ore Mills
Using Leaching Process IX-47
IX-21 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Mercury Mines IX-49
IX-22 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Uranium Mines IX-53
IX-23 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Antimony Mines IX-57
IX-24 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Platinum Mills IX-60
IX-25 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Titanium Mines IX-63
xxviii
DRAFT
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DRAFT
TABLES (cont.)
No. Title Page
IX-26 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Titanium Mills IX-65
IX-27 Parameters Selected and Effluent Limitations
Recommended for BPCTCA—Titanium Dredge Mine
With Wet Separation Mill IX-67
X-l Parameters Selected and Effluent Limitations
Recommended for BATEA—Iron-Ore Mines X-6
X-2 Parameters Selected and Effluent Limitations
Recommended for BATEA—Iron-Ore Mills Employing
Physical Methods and/or Chemical Reagents X-8
X-3 Parameters Selected and Effluent Limitations
Recommended for BATEA—Lead and Zinc Mines
Having Solubilization Potential X-12
X-4 Parameters Selected and Effluent Limitations
Recommended for Alkaline Mine Drainage
BATEA—Bauxite Mines (Acid or Alkaline Mine
Drainage) X-18
X-5 Parameters Selected and Effluent Limitations
Recommended for BATEA—Ferroalloy-Ore Mines X-20
X-6 Parameters Selected and Effluent Limitations
Recommended for BATEA—Ferroalloy-Ore Mills
Treating Less Than 5,000 Metric Tons (5,512
Short Tons) Per Year X-22
X-7 Parameters Selected and Effluent Limitations
Recommended for BATEA—Ferroalloy-Ore Mills
Treating More Than 5,000 Metric Tons (5,512
Short Tons) Per Year by Physical Processing X-22
X-8 Parameters Selected and Effluent Limitations
Recommended for BATEA—Ferroalloy-Ore Mills
Using Flotation Process X-25
X-9 Parameters Selected and Effluent Limitations
Recommended for BATEA—Ferroalloy-Ore Mills
Using Leaching Process X-27
X-10 Parameters Selected and Kffluent Limitations
Recommended for BATKA—Ur.mium Mines X-JO
X-ll Parameters Selected and Kf fluent Li mi tat Urns
Recommended Cor HATKA—Platinum Mi Us X- 12
XI-1 Parameters Selected and lifflueut Limitations
Recommended for NSl'S — Ferroalloy-Ore Mines X I-5
XI-2 Parameters Selected and Effluent Llmlt.it ions
Recommended for NSPS—Ferroalloy-Ore Mills
Using Flotation Process XT-7
XI-3 Parameters Selected and Effluent Limitations
Recommended for NSPS—Uranium Mines XI-8
xxix
DRAFT
-------
DRAFT
SECTION 1
CONCLUSIONS
To establish effluent limitation guidelines and standards i>f
performance, the ore mining and dressing industry was divided
into 43 separate categories and subcategories for which separ-
ate limitations were recommended. This report deals with the
entire metal-ore mining and dressing industry and examines
the industry by ten major categories: iron ure; copper ore;
lead and zinc ores; gold ore; silver ore; bauxite ore; ferro-
alloy-metal ores; mercury ores; uranium, rudltun and vanadium
ores; and metal ores, not elsewhere classified (orus of anti-
mony, beryllium, platinum, rare earths, tin, tJtjn lum, and
zirconium). The subcatcgorlzation of the ore c.ilcgnrlc.s Is
based primarily upon ore mineralogy and processing or extrdc-
tion methods employed; however, other factors (such as size,
climate or location, and method of mining) are used in some
Instances.
Based upon the application of the best practicable control
technology currently available, mining or milling facilities
in the 16 of 43 subcategories for which separate limitations
are proposed can be operated with no discharge of process
wastewater. With the best available technology economically
achievable, facilities in 22 of the 43 subcategories can be
operated with no discharge of process wastewater to navigable
waters. No discharge of process wastewater is also achiev-
able as a new source performance standard for facilities in
22 of the 43 subcategories.
Examination of the wastewater treatment methods employed in
the ore mining and dressing industry indicates tli.it mil ing
ponds or other types of sedimentation impoundments an- the
most commonly used methods of suspended-sol id remov;i I, .and
that these Impoundments provide the additional benefit of
reduction of dissolved parameters as well. Tailing impound-
ments also servo to equalize- flow rates and concent r.it inns of
wastewater parameters.
It is concluded that, for areas of excess water l>.i 1 .nu-e, the
practices of runoff diversion, segregdtiou of wdsie *troams,
and reduction in the use of process water will assist in the
attainment of no discharge for the speciI led snlu .n cgoi Ics.
Effective chemlcdl -treatment, methods which will result in
significant Improvement in d ischaryo-wat IT <|n
-------
DRAFT
SECTION II
RECOMMENDATIONS
The recommended effluent limitation guidelines based on the
beat practicable control technology currently available (BPCTCA)
are summarized in Table II-l. Based on information contained
in Sections III through VIII, it is recommended that facilities
in 16 of the 43 subcategorles achieve no discharge of process
wastewater.
The recommended effluent limitation guidelines based upon the
best available technology economically achievable (BATEA) are
summarized in Table IJ-2. Of the 43 subcategorics listed for
which separate limitations are recommended, it Is recommended
that facilities in 22 subcategories achieve no discharge of
process wastewater by 1983.
The new source performance standards (NSPS) recommended for
operations begun after the promulgation of recommended guide-
lines for the ore mining and dressing industry are summarized
in Table II-3. With the exception of three subcategories, new
source performance standards are identical to BPCTCA and BATEA
recommended effluent limitations.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
11-1
DRAFT
-------
DRAFT
TABLE 11-1. SUMMARY OF RECOMMENDED BPCTCA EFFLUENT LIMITATIONS BY
CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
LIMITATIONS ARE PROPOSED (Sheet 1 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
IRON ORES
Mines
Mills
{Physical/Chemical Separation
Magnetic and Physical Separation (Mesabi Range)
X
IX-1
IX-2
COPPER ORES
Mines
Mills
( Open-Pit, Underground, Stripping
1 Hydrometallurgical (Leaching)
{Vat Leaching
Flotation (Net Evaporation >76.2 em'}
Flotation (Net Evaporation < 76.2 cm'}
X
X
X
IX-3
1X4
LEAD AND ZINC ORES
Mines
( No Solubilization Potential
I SolubHIzation Potential
Mills
IX-5
1X6
IX-7
GOLD ORES
Mines
Mills
! Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation
X
IX-8
IX-9
1X10
IX-11
SILVER ORES
Mines
Mills
{Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
X
X
IX-12
IX 13
IX 14
BAUXITE ORE
Mines ||
IX 15
•76.2 cm - 30 In.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
II-2
DRAFT
-------
DRAFT
TABLE 11-1. SUMMARY OF RECOMMENDED BPCTCA EFFLUENT LIMITATIONS BY
CATEGORY AND SUBCATEGORY -ORES FOR WHICH SEPARATE
LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills <
< 6.000 metric tonjt/year
> 6.000 metric tons* /year by Physical Processes
> 5.000 metric tons'/year by Flotation
Leaching
IX-16
IX 17
IX 18
IX 19
IX -20
MERCURY ORES
Mines
Mills <
Gravity Separation
Flotation Process
X
X
IX-21
URANIUM. RADIUM. VANADIUM ORES
Mines
Mills <
Acid or Acid/Alkaline Leeching
Alkaline Leaching
X
X
IX-22
ANTIMONY ORES
Mines
Mills -
Flotation Process
X
1X23
BERYLLIUM ORES
Mines
Mills
X
X
PLATINUM ORES
Mines or Mine/Mills
IX-24
RARE-EARTH ORES
Mines
Mills -
Flotation or Leaching
X
X
TITANIUM ORES
RfllfMI
Mills i
Electrostatic/Magnetic end Gravity/Flotation Processes
Physical Processes with Dredge Mining
1X25
IX-26
IX 27
6.000 metric tons • 6,612 short tons
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
II-3
DRAFT
-------
DRAFT
TABLE 11-2. SUMMARY OF RECOMMENDED BATEA EFFLUENT LIMITATIONS BY
CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
LIMITATIONS ARE PROPOSED (Sheet 1 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
IRON ORES
Mines
Mills
{Physical/Chemical Separation
Magnetic and Physic.il Separation (Mesabi Rdngo)
X
X-1
X-2
COPPER OHfcS
Mines
Mills
{Open-Pit, Underground. Stripping
Hydrometallurgical (Laachmgl
{Vat Leaching
Flotation (Net Evaporation > 76.2 cmVyear)
Flotation (Net Evaporation < 76.2 cmVyear)
X
X
X
X
(Same ai BPCTCA)
LEAD AND ZINC ORES
Mines
J No Solubilization Potential
| Solubilization Potential
Mills
X
(SameasBPCTCA)
X-3
GOLD ORES
Mines
Mills
ICyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation
X
X
X
(Same as BPCTCA)
(Same as BPCTCA)
SILVEH ours
Mines
Mill*
I Flnt.ilitiii Procuss
* Cydiiiclulion I'rniuss
\ Amalgamation PIOIIISS
' Gravity Sepdrnlitui
X
X
X
(SameasBPCTCA)
(Same as BPCTCA)
UAUXITE OHt
Mines
X-4
•76.2 cm • 30 in
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IT-4
DRAFT
-------
DRAFT
TABLE 11-2. SUMMARY OF RECOMMENDED BATEA EFFLUENT LIMITATIONS BY
CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
!< 5,000 metric tons'/year
> 5.000 metric tons'/year by Physical Processes
> 5.000 metric tons* /year by Flotation
Leaching
X-5
X-6
X-7
X-8
X-9
MERCURY ORES
Mines
Mills
{Gravity Separation
Flotation Process
X
X
(Same es BPCTCA)
URANIUM. RADIUM. VANADIUM ORES
Mines
Mills
1 Acid or Acid/Alkaline Leaching
| Alkaline Leaching
X
X
X-10
ANTIMONY ORES
Mines
Mills
— Flotation Process
X
(Same as BPCTCA)
BERYLLIUM ORES
Mines
Mills
X
X
PLATINUM ORI S
Mines or
Mum/Milk
1
XII
HAHt L AMI HOMES
Minns
Mills
1 loltifioit 01 LiMitimg
X
X
TITANIUM OKCS
Mines
Mills
j Eluclrusiiilic/M.iflnnli< ,IIH| Gr.ivilv/Floldliuii 1'iticusvs
| Physical I'liicunes willi (JitKlgu MIIIMII)
X
(Same as BPCTCAI
(Same as BPCTCA)
5.00O miitrir Ions b.SIV slum Inns
NOTICE THESE ARE TENTATIVE HECOMMtNDATIONS BASED UPON INFUKMATIOM IN THIS REPOR T AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVf.D AND FURTHER INTERNAL REVIEW HY EPA
I l-'i
DRAFT
-------
DRAFT
TABLE 11-3. SUMMARY OF RECOMMENDED NSPS EFFLUENT LIMITATIONS BY
CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
LIMITATIONS ARE PROPOSED (Sheet 1 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
IRON ORES
Mines
Mills
/ Physical/Chemical Separation
| Magnetic and Physical Separation (Mesabi Range)
X
(Same as BPCTCA)
(Same as BPCTCA)
COPPER ORES
Mines
Mills
J Open-Pit. Underground. Stripping
| Hydrometallurgical (Leaching)
/ Vat Leaching
< Flotation (Net Evaporation > 76.2 cmVyear)
' Flotation (Net Evaporation < 76.2 cm'/year)
X
X
X
X
(Same as BPCTCA)
LEAD AND ZINC ORES
Mines
{No Solubilization Potential
Solubilization Potunlial
Mills
X
(Same as BPCTCA)
(SameasBATEA)
GOLD ORES
Mines
Mills
iCyamdation Process
Amalgamation Process
Flotation Process
Gravity Separation
X
X
X
(Same as BPCTCA)
(Same as BPCTCA)
SILVER ORES
Mines
Mills
(Flotation Process
Cyanidalion Process
1 Amalgamation PIOCIISI
' Gravity Separation
X
X
X
(Same as BPCTCA)
(Same as BPCTCA)
BAUXITE ORE
Mines
I
(Same as BATEA)
•76 2 cm - 30 in.
NOTICE THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
ARE
TJ-ft
DRAFT
-------
DRAFT
TABLE 11-3. SUMMARY OF RECOMMENDED NSPS EFFLUENT LIMITATIONS BY
CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
!< 5.000 metric tons' /year
> 5.000 metric tons' /year by
> 5,000 metric tons'/year by
Leaching
Physical Processes
Flotation
Xl-l
(Same us BATEA)
(Same as BATEA)
XI-2
(Same as BATEA)
MERCURY ORES
Mines
Mills
{Gravity Separation
Flotation Process
X
X
(Same as BPCTCA)
URANIUM. RADIUM. VANADIUM ORES
Mines
Mills
1 Acid or Acid/Alkaline Leaching
{ Alkaline Leaching
X
X
XI-3
ANTIMONY ORES
Mines
Mills
— Flotation Process
X
(Same as BPCTCA)
BERYLLIUM ORES
Mines
Mills
X
X
PLATINUM ORES
Mines or Mine/Mills
(Same as BATEA)
RARE 1 AHTHORES
Mines
Mills
Flolnlion <» Limrhing
X
X
TITANIUM ORES
Mines
Mills
{Elactrostatic/Miignetir .nicl Clrnviiy/riolnlioii Pinrussiit
Physical Processes with Oimlgo Mining
X
(Some ns BPCTCA)
(Sumo us BPCTCAI
5.000 mulnc tons - b.512 ihon tout
NOTICE THESE ARE TENTATIVE RECOMMENDATIONS RASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
I 1-7
DRAFT
-------
DRAFT
SECTION III
INTRODUCTION
PURPOSE AND AUTHORITY
The United States Environmental Protection Agency (EPA) is
charged under the Federal Water Pollution Control Act Amend-
ments of 1972 with establishing effluent limitations which
must be achieved by point sources of discharge into the
navigable water of the United States.
Section 301(b) of the Act requires the achievement, by not
later than July 1, 1977, of effluent limitations for point
sources, other than publicly owned treatment works, which
are based on the application of the best practicable control
technology currently available as defined by the Adminis-
trator pursuant to Section 304(b) of the Act. Section 301(b)
also requires the achievement, by not later than July 1, 1983,
of effluent limitations for point sources, other than publicly
owned treatment works, which are based on the application of
the best available technology economically achievable which
will result in reasonable further progress toward the national
goal of eliminating the discharge of all pollutants, as deter-
mined in accordance with regulations issued by the Administrator
pursuant to Section 304(b) to the Act. Section 306 of the
Act requires the achievement by new sources of a Federal
standard of performance providing for the control of the
discharge of pollutants which reflects the greatest degree
of effluent reduction which the Administrator determines to
be achievable through the application of the best available
demonstrated control technology, processes, operating methods,
or other alternatives, including, where practicable, a stan-
dard permitting no discharge of pollutants. Section 304(b)
of the Act requires the Administrator to publish, within one
year of enactment of the Act, regulations providing guide-
lines for effluent limitations setting forth the degree of
effluent reduction attainable through the application of the
best practicable control technology currently available and
the degree of effluent reduction attainable through the
application of the best control measures and practices
achievable including treatment techniques, process and pro-
cedure innovations, operating methods and other alternatives.
The regulations proposed herein set forth effluent limitations
guidelines pursuant to Section 304(b) of the Act for the Ore
Mining and Dressing Industry point source category.
III-l
DRAFT
-------
DRAFT
Section 306 of the Act requires the Administrator, within one year
after a category of sources is included in a list published pur-
suant to Section 306(b) (1) (A) of the Act, to propose regulations
establishing Federal standards of performance for new sources
within such categories. Section 307 of the Act requires the Admin-
istrator to promulgate pretreatment standards for new sources
simultaneously with the promulgation of standards of performance
under Section 306. The Administrator published, in the Federal
Register of January 16, 1973 (38 F.R. 1624), a list of 24 source
categories. Publication of an amended list will constitute announce-
ment of the Administrator's intention of establishing, under Section
306, standards of performance applicable to new sources within the
ore mining and dressing Industry, and under Section 307, pretreatment
standards. The list will be amended when proposed regulations for
the Ore Mining and Dressing Industry are published in the Federal
Register.
The subgroups of the metal mining industries are identified as major
group 10 in the Standard Industrial Classification (SIC) Manual, 1972,
published by the Executive Office of the President (Office of Manage-
ment and Budget). This industry category includes establishments
engaged in mining ores for the production of metals, and includes all
ore dressing and beneficiating operations, whether performed at mills
operating in conjunction with the mines served or at mills operated
separately. These include mills which crush, grind, wash, dry, sinter,
or leach ore, or perform gravity separation or flotation operations.
The industry categories covered by this report include the following:
SIC 1011 - Iron Ores
SIC 1021 - Copper Ores
SIC 1031 - Lead and Zinc Ores
SIC 1041 - Gold Ores
SIC 1044 - Silver Ores
SIC 1051 - Bauxite Ores
SIC 1061 - Ferroalloy Ores
SIC 1092 - Mercury Ores
SIC 1094 - Uranium/Radium/Vanadium Ores
SIC 1099 - Metal Ores, Not Elsewhere Classified
The guidelines in this document identify, in terms of the chemical.
physical, and biological characteristics of pollutants, the lovol
of pollutant reduction attainable through application of tin? best
practicable control technology currently available, and best avail-
able technology economically achievable. Standards of perfornuinre
for new sources and pretreatment are also presented. The guidelines
also consider a number of other factors, such as the costs of
achieving the proposed effluent limitations and nonwater-quality
environmental impacts (Including energy requirements resulting from
application of such technologies).
III-2
DRAFT
-------
DRAFT
SUMMARY OF METHODS USED FOR DEVELOPMENT OF EFFLUENT
LIMITATION GUIDELINES AND STANDARDS OF TECHNOLOGY
The effluent limitations guidelines and standards of per-
formance proposed herein were developed in a series of
systematic tasks. The Ore Mining and Dressing Industry
was first studied to determine whether separate limitations
and standards would be appropriate for different SIC cate-
gories. Development of reasonable Industry categories nnd
subcategories and establishment of effluent guidelines and
treatment standards require a sound understand Ing and know-
ledge of the Ore Mining and Dressing Industry, i he mining
techniques and milling processes Involved, tin- mineralogy
of the ore deposits, water use, wastewater generation and
characteristics, and the capabilities of existing control
and treatment technologies.
Approach
This report describes the results obtained from application
of the above approach to the mining of metals and ore min-
erals for the ore mining and dressing industry. The survey
and extensive sampling and analysis covered a wide range of
processes, products, and types of wastes. In each SIC cate-
gory, slightly different evaluation criteria were applied
initially, depending upon the nature of the extraction pro-
cesses employed, locations where mining activities occur,
mineralogical differences, treatment and control technology
employed, and water usage in the industry category. The
following discussion Illustrates the manner In which the
effluent guidelines and standards of performance were
developed.
Data Base
Each SIC category was first examined to determine the range
of activities incorporated by the industry classification.
Information used as a data base for detailed examination
of each category was obtained from a wide variety of sources
including published data from journals and trade literature,
mining industry directories, general business publications,
texts on mining/milling technology, texts on industrial
wastewater control, summaries of production of the partIculur
metals of interest, U.S. Bureau of Mines annual snmniarU-s,
III-3
DRAFT
-------
DRAFT
U.S. Environmental Protection Agency publications, U.S.
Geological Survey publications, surveys performed by industry
trade associations, NPDES permits and permit applications,
and numerous personal contacts. Additional information was
supplied by surveys of research performed in the application
of mining, extractive processing, and effluent control tech-
nology. Various mining company personnel, independent
researchers, and state and federal environmental officials
also supplied requested information.
Categorization and Waste Load Characterization
After assembly of an extensive data base, each SIC code
group or subgroup was examined to determine whether differ-
ent limitations and standards would be appropriate. In
several categories, it was determined that further subdivi-
sion' was unnecessary. In addition, after further study
and site visits, subcategory designations were later reduced
within a category in some instances. Where appropriate,
subcategorization consideration was based upon whether the
facility was a mine or a concentrating facility (mill),
and further based upon differences such as raw material
extracted or used, milling or concentration process employed,
waste characteristics, treatability of wastes, reagents used
in the process, treatment technology employed, water use and
balance, end products or byproducts. Other factors considered
were the type of mine (surface or underground), geographic
location, size, age of the operation, and climate.
Determination of the wastewater usage and characteristics
for each subcategory as developed in Section IV and discussed
in Section V included: (1) the source and volume of water
used in the particular process employed and the source of
waste and wastewaters in the plant, and (2) the constituents
(including thermal) of all waste waters, including pollutants,
and other constituents which result in taste, odor, and color
in water or aquatic organisms. Those constituents discussed
in Section V and Section VI which are characteristic of t lie-
industry and present In measurable quantities won- se I cited
as pollutants subject to effluent limitation guide.! Ines .nul
standards.
Site Visits and Sampling Program
Based upon information gathered as part of the dssembly ol
a data base, examination of NPDES permits and permit appli-
cations, surveys by trade associations, and examination ol
texts, Journals, and the literature available on treatment
practices in the Industry, selection of mining and milling
III—'«
DRAFT
-------
DRAFT
operations which were thought to embody exemplary treatment
practice was made for the purpose of sampling and verifica-
tion, and to supplement compiled data. All factors poten-
tially influencing industry subcategorization were represented
by the sites chosen. Detailed information on production,
water use, wastewater control, and water treatment practices
was obtained. As a result of the visits, many subcategorles
which had been tentatively determined were found to be unnec-
essary. Flow diagrams were obtained indicating the course of
wastewater streams. Control and treatment plant design and
detailed cost data were compiled.
Sampling and analysis of raw and treated effluent streams,
process source water, and ititermediate process or treatment
steps were performed as part of the site visits. In-situ
analyses for selected parameters such as temper.iture, pll,
dissolved oxygen, jnd specific conductance wure performed
whenever possible. Historical data for the same waste sireams
was obtained when available.
Raw waste characteristics were then identified for each sub-
category. This included an analysis of all constituents of
wastewaters which might be expected in effluents from mining
and milling operations. In addition to examination of can-
didate control parameters, a reconnaissance investigation
of some 55 chemical parameters was performed upon raw and
treated effluent for each site visited. Additionally, limited
sampling of mine waters for several radiological parameters
was accomplished at selected sites. Raw and treated waste
characterization during this study was based upon a detailed
chemical analysis of the samples and historical effluent
water quality data supplied by the industry and Federal and
State regulatory agencies.
Cost Data liase
Cost information contained in this report was obtained
directly from industry during plant visits, from enj; I neei -
ing firms, equipment suppliers, .uul from Liu- 1 11 era t ure .
The lnforin.il ion obtained I ruin them- sourei-s h.is lu-en n^i-il
tci develop goner.11 capital, opera l lug, ami overall iosl^
for etich 11:0. it men L mid control method. Where d.ila was l.u-k-
ing, costs were developed p.irameir n-.i I I v I rom knowledge nl
equipment required, processes employed, constriu-iIon, and
maintenance requirements. This generalized cost ilain plus
the specific information obtained from plant visits was then
used for cost effectiveness estimates in Section Vlll .uul
wherever else costs nre mentioned iti this report.
I I L-"J
DRAFT
-------
DRAFT
Treatment and Control Technologies
The full range of control and treatment technologies exist-
ing within each subcategory was identified. This included
an identification of each control and treatment technology,
including both in-plant and end-of-process technologies,
which is existent or capable of being designed for each
subcategory. It also included an identification of the
amounts and the characteristics of pollutants resulting
from the application of each of the control and treatment
technologies. The problems, limitations, and reliability
of each control and treatment technology were also
Identified. In addition, the nonwater-quality environ-
mental impact—such as the effects of the application of
such technologies upon other pollution problems, including
air, solid waste, noise, and radiation—was also Identified.
The energy requirements of each of the control and treat-
ment technologies were identified, as well as the cost of
the application of such technologies.
Selection of_ BPCTCA, BATEA, and New Source Standards
All data obtained were evaluated to determine what levels of
treatment constituted "best practicable control technology
currently available" (BPCTCA), "best available technology
economically achievable" (BATEA), and "best demonstrated con-
trol technology, processes, operating methods, or other
alternatives." Several factors were considered in identi-
fying such technologies. These included the application of
costs of the various technologies in relation to the effluent
reduction benefits to be achieved through such application,
engineering aspects of the application of various types of
control techniques or process changes, and nonwater-quality
environmental Impact. Efforts were also made to determine
the feasibility of transfer of technology from subcategory
to subcategory, other categories, and other industries where
similar effluent problems might occur. Consideration of the
technologies was not limited to those presently employed In
the industry, but Included also those processes in pilot-
plant or laboratory-research stages.
LII-6
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SUMMARY OF ORE-BENEF1CIATION PROCESSES
General Discussion
As mined, most ores contain the valuable metals, whose recovery
is sought, disseminated in a matrix of less valuable rock, called
gangue. The purpose of ore beneficiation is the separation of
the metal-bearing minerals from the gangue to yield a more useful
product—one which is higher in metal content. To accomplish
this, the ore must generally be crushed and/or ground small
enough so that each particle contains mostly the mineral to be
recovered or mostly gangue. The separation of the particles
on the basis of some difference between the ore mineral and
the gangue can then yield a concentrate high in metal value, as
well as waste rock (tailings) containing very little metal.
The separation is never perfect, and the degree of success
which is attained is generally described by two numbers:
(1) percent recovery and (2) grade of the concentrate. Widely
varying results are obtained in beneficiating different ores;
recoveries may range from 60 percent or less to greater than
95 percent. Similarly, concentrates may contain less than
60 percent or more than 95 percent of the primary ore mineral.
In general, for a given ore and process, concentrate grade
and recovery are inversely related. (Higher recovery is
achieved only by including more gangue, yielding a lower-grade
concentrate.) The process must be optimized, trading off
recovery against the value (and marketability) of the concen-
trate produced. Frequently, depending on end use, a particular
grade of concentrate is desired, and specific gangue components
are limited as undesirable impurities.
Many properties are used as the basis for separating valuable
minerals from gangue, including: specific gravity, conductivity,
magnetic permeability, affinity for certain chemicals, solubility,
and the tendency to form chemical complexes. Processes for
effecting the separation may be generally considered as: gravity
concentration, magnetic separation, electrostatic separation,
flotation, and leaching. Amalgamation and cyanidation are
variants of leaching which bear special mention. Solvent
extraction and ion exchange are widely applied techniques for
concentrating metals from leaching solutions, and for separating
them from dissolved contaminants. All of these processes are
discussed in general terms—with examples—in the paragraphs
that follow. This discussion is not meant to be all-inclusive;
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rather, its purpose is to discuss Che primary processes in
current use in the ore mining and milling industry. Details
of processes used in typical mining and milling operations
are provided, together with process flowcharts, under
"General Description of Industry By Ore Category."
Gray it y-Conc en t ra t ion Processes
General. Gravity-concentration proe-esses exploit di f rereiuA-s
in density to separate valuable ore minerals from gangue.
Several techniques (jigging, tabling, spirals, sink/float
separation, etc.) are used to achieve the separation. Each
is effective over a somewhat limited range of particle sizes,
the upper bound of which is set by the size of the apparatus
and the need to transport ore within it, and the lower bound,
by the point at which viscosity forces predominate over
gravity and render the separation ineffective. Selection of
a particular gravity-based process for a given ore will be
strongly influenced by the size to which the ore must be crushed
or ground to separate values from gangue, as well as by the
density difference and other factors.
Most gravity techniques depend on viscosity forces to suspend
and transport gangue away from the (heavier) valuable mineral.
Since the drag forces on a particle depend on its area, and its
weight on its volume, particle size as well as density will
have a strong influence on the movement of a partii-le in .1
gravity separator. Smaller particles of ore mlncrdJ may he-
carried with the gangue, despite their higher density, or
larger particles of gangue may be included in the gravity
concentrate. Efficient separation thus depends on a feed to
the process which contains a small dispersion of particle
sizes. A variety of classifiers—spiral and rake classifiers,
screens, and cyclones—is used to assure a reasonably uniform
feed. At some mills, a number of sized fractions of ore are
processed in different gravity-separation units.
Viscosity forces on the particles set a lower limit, for
effective gravity separation by any technique*. lroi sufficiently
small partle-le.s, eve«n the smallest t urhu le-nce* suspends tin*
p.i rl It-It- lor long p<>r Joels of liim-, ren-inl le>ss ol lU-nsily.
Sin li slimes, ont-i- lormeil, i-.mnnt. lie- I n i>vi-i i-il l*v j.'.i.ivllv
Le-chn I c|ii«". .mil m.iy cause very low recuve i i e-. in f.i.-ivilv
professing ol IllK-'ly frl.iMe* ores, sin li .1:. -,i heel Me li.ili I inn
I unysl
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Jigs. Jigs of many different designs are used to achieve
gravity separation of relatively coarse ore (generally, a
secondary crusher product between 0.5 mm and 25 mm—up to
1 in.—in diameter). In general, ore Is fed as a thick slurry
to a chamber in which agitation is provided by a pulsating
plunger or other such mechanism. The feed separates into layers
by density within the Jig, the Jlghter gangue being drawn off
at the top with the water overflow, and the denser mineral,
at a screen on the bottom. Often, a bed of coarser ore or iron
shot is used to aid the separation; the dense ore mineral
migrates down through the bed under the influence of the
agitation within the jig. Several jigs are most often used,
in series, 'to achieve both acceptable recovery and high concen-
trate grade.
Tables. Shaking tables of a wide variety of designs have
found widespread use as an effective means of achieving gravity
separation of finer ore particles (0.08 to 2.5 mm—up to 0.1
in.—in diameter). Fundamentally, they are, as the name implies,
tables over which water carrying ore particles flows. A series
of ridges or riffles, approximately perpendicular to the water
flow, traps heavy particles, while lighter ones are suspended
by shaking the table and flow over the obstacles with the water
stream. The heavy particles move along the ridges to the edge
of the table and are collected as concentrate (heads), whiJe
the light material which follows the water flow is generally
a waste stream (tails) . Between these streams is generally some
material (termed "middlings") which has been diverted somewhat
by the riffles, although less than the heads. These are often
collected separately and returned to the table feed. Repro-
cessing of either heads or tails, or both, and multiple stages
of tabling are not uncommon. Tables may be used to separate
minerals differing relatively little in density, but uniformity
of feed becomes extremely important in such cases.
Spirals. Humphreys spiral separators, a relatively recent'
development, provide an efficient means of gravity separation
for large, volumes of material between 0.1 mm and 1 mm (up to
approximately O.OJ In.) In diameter und have been wldoJy
appiled—particularly, in the processing of heavy sands for
llmenlte (FeTlOJ) and mon.i7.lte. (a rare-earth phosphate) .
They consist of a helical conduit (usually, of five turns)
about a vertical axis. A slurry of ore is feil to the conduit
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at the top and flows down the spiral under gravity. The
heavy minerals concentrate along the inner edge of the spiral,
from which they may be withdrawn through a series of ports.
Wash water may also be added through ports along the inner
edge to improve the separation efficiency. A single spiral
may, typically, be used to process 0.5 to 2.4 metric tons
(0.55 to 2.64 short tons) of ore per hour; In large plants,
as many as several hundred spirals may be run in parallel.
Sink/Float Separation. Sink/float separators differ from
most gravity methods in that buoyancy forces are used to
separate the various minerals on the basis of density. The
separation is achieved by feeding the ore to a tank containing
a medium whose density is higher than that of the gangue and
less than that of the valuable ore minerals. As a result,
the gangue floats and overflows the separation chamber, and
the denser values sink and are drawn off at the bottom—often,
by means of a bucket elevator or similar contrivance. Because
the separation takes place in a relatively still basin and
turbulence is minimized, effective separation may be achieved
with a more heterogeneous feed than for most gravity-separation
techniques. Viscosity does, however, place a lower bound on
particle size for practicable separation, since small particles
settle very slowly, limiting the rate at which ore may be fed.
Further, very fine particles must be excluded, since they mix
with the separation medium, altering its density and viscosity.
Media commonly used for sink/float separation in the are
milling industry are suspensions of very fine ferrosilicon or
galena (PbS) particles. Ferrosilicon particles may be used to
achieve medium specific gravities as high as 3.5 and are used
in "Heavy-Medium Separation." Galena, used in the "Huntington-
Heberlein" process, allows the achievement of somewhat higher
densities. The particles are maintained in suspension by a
modest amount of agitation in the separator and are recovered
for reuse by washing both values and gangue after separation.
Magnetic Separation
Magnetic separation is widely applied In the ore milling Industry,
both for the extraction of values from orc> mid fur llu- separation nl"
different valuable minerals recovered from complex orr^. Kxl t;nsl ve
use of magnetic, separation Is inudc In I IK- processing ol ores <>l lion,
columbium and tantalum, and tungsten, in mime a lew. The Repartition Is
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based on differences in magnetic permeability (which, although
small, is measurable for almost all materials) and is effective
in handling materials not normally considered magnetic. The
basic process involves the transport of ore through a region of
high magnetic-field gradient. The most magnetically permeable
particles are attracted to a moving surface, behind which is
the pole of a large electromagnet, and are carried by it out of
the main stream of ore. As the surface leaves the high-field
region, the particles drop off—generally, into a hopper or onto
a conveyor leading to further processing.
For large-scale applications—particularly, in the iron-ore
industry—large, rotating drums surrounding the magnet are
used. Although dry separators are used for rough separations,
these drum separators are most often run wet on the slurry
produced in grinding mills. Where smaller amounts of material
are handled, wet and crossed-belt separators are frequently
employed.
Electrostatic Separation
Electrostatic separation is used to separate minerals on the
basis of their conductivity. It is an inherently dry process
using very high voltages (typically, 20,000 to 40,000 volts).
In a typical implementation, ore is charged to 20,000 to 40,000
volts, and the charged particles are dropped onto a conductive
rotating drum. The conductive particles discharge very rapidly
and are thrown off and collected, while the non-conductive
particles keep their charge and adhere by electrostatic attraction.
They may then be removed from the drum separately.
Flotation Processes
Basically, flotation is a process whereby particles of one
mineral or group of minerals are made, by addition of chemicals,
to adhere preferentially to air bubbles. When air is forced
through a slurry of mixed minerals, then, the rising bubbles
carry with them the particles of the mineral(s) to be separated
from the matrix. If a foaming agent Ls added which prevents
the bubbles from bursting when they reach the surface, a layer
of mineral-laden foam Is buJlt up at flie surface of the flotation
cell which may be removed to recover the mineral. Requirements
for the success of the operation are Lluit particle size be
small, that reagents compatible with the mineral to he recovered
he u«fd, and that water conditions in the cell noL Interfere
with attachment of reagents to mineral or Lo nlr bubbles.
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Flotation concentration has become a mainstay of the ore
milling industry. Because it is adaptable to very fine
particle sizes (less than 0.001 cm), it allows high rates of
recovery from slimes, which are inevitably generated in crushing '
and grinding and which are not generally amenable to physical
processing. As a physico-chemical surface phenomenon, it can
often be made highly specific, allowing production of high-grade
concentrates from very-low-grade ore (e.g., over 95-percent
MoS2. concentrate from 0.3-percent ore). Its specificity also
allows separation of different ore minerals (e.g., CuS, PbS,
and ZnS), where desired, and operation with minimum reagent
consumption, since reagent interaction is typically only with
the particular materials to be floated or depressed.
Details of the flotation process—exact suite and dosage of
reagents, fineness of grinds, number of regrinds, cleaner-
flotation steps, etc.—differ at each operation where it is
practiced and may often vary with time at a given mill.
A complex system of reagents is generally used, including
four basic types of compounds: collectors, frothers, activators,
and depressants. Collectors serve to attach ore particles to
air bubbles formed in the flotation cell. Frothers stabilize
the bubbles to create a foam which may be effectively recovered
from the water surface. Activators enhance the attachment of
specific kinds of particles to the air bubbles, and depressants
prevent it. Frequently, activators are used to allow flotation
of ore depressed at an earlier stage of the milling process.
In almost all cases, use of each reagent in the mill is low
(generally, less than 0.5 kg—approximately 1 Ib—per ton of
ore processed) , and the bulk of the reagent adheres to tailings
or concentrates.
Sulfide minerals are all readily recovered by flotation using
similar reagents in small doses, although reagent requirements
and ease of flotation do vary throughout the class. Sulfide
flotation is most often carried out at alkaline pH. Collectors
are most often alkaline xanthates having two to five carbon
atoms—for example, sodium ethyl xanthate (NaS2COC2H5).
Frothers are generally organics with a soluble hydroxyl group
and a "non-wettable" hydrocarbon. Pine oil (C6H120H), for
example, is widely used to allow separate recovery of metal
values from mixed-sulfide ores. Sodium cyanide is widely
used as a pyrite depressant. Activators useful in sulfide-ore
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flotation may include cuprous sulfide and sodium sulfide.
Sulfide minerals of copper, lead, zinc, molybdenum, silver,
nickel, and cobalt are commonly recovered by flotation.
Many minerals in addition to sulfides may be, and often are,
recovered by flotation. Oxidized ores of iron, copper,
manganese, the rare earths, tungsten, titanium, and columbium
and tantalum, for example, may be processed in this way.
Flotation of these ores Involves a very different suite of
reagents from sulfide flotation and has, in some cases, required
substantially larger dosages. Experience has shown these
flotation processes to be, in general, somewhat more sensitive
to feed-water conditions than sulfide floats; consequently,
they are less frequently run with recycled water. Reagents
used Include fatty acids (such as oleic acid or soap skimmings),
fuel oil, and various amines as collectors; and compounds such
as copper sulfate, acid-dichromate, and sulfur dioxide as
conditioners.
Leaching
General. Ores can be beneficlated by dissolving away either
gangue or values in aqueous acids or bases, liquid metals, or
other special solutions. The examples which follow illustrate
various possibilities.
(1) Water-soluble compounds of sodium, potassium, and
boron which are found in arid climates or under
impervious strata can be mined, concentrated,
and separated by leaching with water and recrystal-
lizing the resulting brines.
(2) Vanadium and some other metals form anlonic species
(e.g., vanadates) which occur as Insoluble ores.
Roasting of such insoluble ores with sodium compounds
converts the values to soluble sodium salts (e.g.,
sodium vanadate) . After cooling, the water-soluble
sodium salts are removed from the gangue by Jcaching
In water.
(3) Uranium ores are only mildly solnliio In water, hut
they dissolve quickly Jn ai-id or .ilk.-illne solutions.
Native gold which Is found Ln a Iliu-ly divided
Is soluble in mercury and can hi- t-xlrdited by
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amalgamation (I.e., leaching with a liquid metal).
One process of nickel concentration involves reduction
of the nickel by ferrosilicon at a high temperature
and extraction of the nickel metal into molten iron.
This process, called skip-ladling, Is related to
liquid-metal leaching.
(5) Certain solutions (e.g., potassium cyanide) dissolve
specific metals (e.g., gold) or their compounds, and
leaching with such solutions immediately concentrates
the values.
Leaching solutions can be categorized as strong, general
solvents (e.g., acids) and weaker, specific solvents (e.g., cyanide)
The acids dissolve any metals present, which often include
gangue constituents (e.g., calcium from limestone). They are
convenient to use, since the ore does not have to be ground very
fine, and separation of the tailings from the value-bearing
(pregnant) leach is then not difficult. In the case of sulfuric
acid, the leach is cheap, but energy is wasted in dissolving
unsought-for gangue constituents.
Specific solvents attack only one (or, at most, a few) ore
constituent(s) , including the one being sought. Ore must be
ground finer to expose the values. Heat, agitation, and pressure
are often used to speed the action of the leach, and considerable
effort goes into separation of solids—often, in the form of
slimes—from the pregnant leach.
Countercurrent leaching, preneutralization of lime in the
gangue, leaching in the grinding process, and other combinations
of processes are often seen in the industry. The values contained
in the pregnant leach solution are recovered by one of several
methods, including precipitation (e.g., of metal hydroxides from
acid leach by raising pH) , electrowinning (which is a form of
electroplating), and cementation. Ion exchange and solvent
extraction are often used to concentrate- values before recovery.
Ores can be exposed to Le.icli In n v.irieLy of w;iys. In v.iL
leaching, the process Is carried otiL In .1 eontalner (v.it),
often equipped with facilities for .1^1 Kit Ion, he;Uln>',, .if r;iL inn.
and presaurizat Ion (e.g., P.ieliuc .1 (.inks). .i"----jj_iti le.uhlnn
takes place in the ore body, with Uu- lejeli .ipplloil either by
plumbing or by percolation through overburden. The pregnant
leach solution 1s pumped to the recovery facility ;incl c.in oil en
be recycled. In-situ leaching Is most economic-iij when Lhe ore
body is surrounded by an impervious nuiLrlx. When water sufflees
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as a leach solution and Is plentiful, in-situ leaching is
economical, even in pervious strata. Ore or tailings stored on
the surface can be treated by heap or dump leaching. In this
process, the ore is placed on an impervious layer (plastic
sheeting or clay) that is furrowed to form drains and launders
(collecting troughs), and leach solution is sprinkled over the
resulting heap. The launder effluent is treated to recover
values. Gold, using cyanide leach, and uranium, with sulfuric
acid leach, are recovered in this fashion.
Amalgamation. Amalgamation is the process by which mercury
is alloyed with some other metal to produce an amalgam. This
process is applicable to free milling of precious-metal ores,
which are those in which the gold is free, relatively coarse,
and has clean surfaces. Lode or placer gold/silver that Is
partly or completely filmed with iron oxides, greases, tellurium,
or sulfide minerals cannot be effectively amalgamated. Hence,
prior to amalgamation, auriferous ore is typically washed and
ground to remove any films on the precious-metal particles.
Although the amalgamation process has, in the past, been used
extensively for the extraction of gold and silver from pulverized
ores, it has, due to environmental considerations, largely
been superseded, in recent years, by the cyanidation process.
The properties of mercury which make amalgamation such a
relatively simple and highly efficient process are: (a) its
high specific gravity (13.55 at 20 degrees Celsius, 68 degrees
Fahrenheit); (b) the fact that mercury is a liquid at room
temperature; and (c) the fact that it readily wets (alloys)
gold and silver in the presence of water.
In the past, amalgamation was frequently implemented in
specially designed boxes containing plates (e.g., sheets of
metal such as copper or Muntz metal (Cu/Zn alloy), etc.) with
an adherent film of mercury. These boxes, typically, were
located downstream of the grinding circuit, and the gold was
seized from the pulp as it flowed over the amalgam plates.
In the U.S., this process has been abandoned to prevent stream
pollution.
The current practice of amalgamation in the U.S. Js limited
to barrel amalgamation of a relatively small quantity of
high-grade, gravity-concentrated ore. This form of amalgamation
is the simplest method of treating an enriched gold- or silver-
bearing concentrate. The gravity concentrate Is ground for
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several hours in an amalgam barrel (e.g., a small cylinder
batching mill) with steel balls or rods before the mercury
is added. This mixture is then gently ground to bring the
mercury and gold into intimate contact. The resulting amalgam
is collected in a gravity trap.
Cyanidation. With occasional exceptions, lode gold and silver
ores now are processed by cyanidation. Cyanidation is a
process for the extraction of gold and/or silver from finely
crushed ores, concentrates, tailings, and low-grade mine-run
rock by means of potassium or sodium cyanide, used In dilute,
weakly alkaline solutions. The gold is dissolved hy the
solution according to the reaction:
4Au + SNaCN + 2H20 > 4NaAu(CN)2_ + 4NaOH
and subsequently sorbed onto activiated carbon ("Carbon-.ln-1'u I p"
process) or precipitated with metallic zinc according to the-
react ion:
NaAu(CN)^ + NaCN + Zn + 2H20 > NaZn(CN)^ + Au + ftf + NaOH
The gold particles are recovered by filtering, and the filtrate
is returned to the leaching operation.
A recently developed process to recover gold from cyanide solu-
tion is the Carbon-in-Pulp process. This process was developed
to provide economic recovery of gold from low-grade ores or
slimes. In this process, gold which has been solubilized with
cyanide is brought into contact with 6 x 16 mesh activated
coconut charcoal in a series of tanks. The pulp and enriched
carbon are air lifted and discharged on small vibrating screens
between tanks, where the carbon is separated and moved to the
next adsorption tank, counter-current to the pulp flow. Gold-
enriched carbon from the last adsorption tank Is leached wjth
hot caustic cyanide solution to desorb the gold. This hot,
high-grade solution containing the leached gold is then sent
to electrolytic cells, where the gold and silver nri- deposited
onto stainless sU:i-l wool cathodes.
l»rotreatment of ores containing only llnt-ly divided gold ,nul
sllvcir usually Includes multistage crushing, I Iiu- grind Lug, .uul
classification of the ore pulp into SHIR!
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is treated by agitation leaching in mechanically or air
agitated tanks, and the pregnant solution is separated from
the slime residue by thickening and/or filtration. Alternatively,
the entire finely ground ore pulp may be leached by counter-
current decantation processing. Gold or silver is then
recovered from the pregnant leach solutions by the methods
discussed above.
Different types of gold/silver ore require modification of the
basic flow scheme presented above. At one domestic operation,
the ore is carbonaceous and contains graphitic material, which
causes dissolved gold to adsorb onto the carbon, thus causing
premature precipitation. To make this ore amenable to cyanlda-
tion, the refractory material is oxidized by chlorine treatment
prior to the leaching step. Other schemes which have been
employed Include oxidation by roasting and blanking the carbon
with kerosene or fuel oil to inhibit adsorption of gold from
solution.
Other refractory ores are those which contain sulfides.
Roasting to liberate the sulfide-enclosed gold and precondi-
tioning by aeration with lime of ore containing pyrrhotite are
two processes which allow conventional cyanidation of these ores.
The cyanidation process is comparatively simple, and is applicable
to many types of gold/silver ore, but efficient low-cost
dissolution and recovery of the gold and silver are possible onJy
by careful process control of the unit operations involved.
Effective cyanidation depends on maintaining and achieving several
conditions:
(1) The gold and silver must be adequately liberated
from the encasing gangue minerals by grinding and,
if necessary, roasting or chlorine treatment.
(2) The concentration of "free" cyanide and dissolved
oxygen in the leaching solution must be kept at a
level that will enable reasonably fast dissolution
of the gold and silver.
(3) The "protective" alkalinity of the Leach solution
must be maintained aL a level that wiJl minimize
consumption of ryanJde by the dissolution of oi hor
meta] •+•". Ur Inn mlner.ils.
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(A) The leach residues must be thoroughly washed without
serious dilution to reduce losses of dissolved
values and cyanide to acceptable limits.
Ion Exchange and Solvent Extraction
These processes are used on pregnant leach solutions to
concentrate values and to separate them from impurities.
Ion exchange and solvent extraction are based on die same
principle: Polar organic molecules tend to exchange a
mobile ion in their structure — typically, C1-, NCtt-, HSCM-, or
C03. — (anions) or H+ or Na+ (cations) — for an ion with a
greater charge or a smaller ionic radius. For example, let
R be the remainder of the polar molecule (in the case, of a
solvent) or polymer (for a resin), and let X be the mobile
ion. Then, the exchange reaction for the example of the
uranyltrisulfate complex is:
4RX + (UO£(S04J3_) ---- ~> R4U02^S04)3_ + 4X-
Thls reaction proceeds from left to right in the loading
process. Typical resins adsorb about ten percent of their
mass In uranium and increase by about ten percent in density.
In a concentrated solution of 'the mobile ion (for example, in
N-hydrochloric acid), the reaction can be reversed, and the
uranium values are eluted (in this example, as hydrouranyl
trisulfuric acid). In general, the affinity of cation-exchange
resins for a metallic cation increases with increasing valence:
Cr-l-H- > Mg-H- > Na+
and, because of decreasing ionic radius, with atomic number:
92U > 42Mo > 23V
and the separation of hexavalent 92U cations by Lon exchange
or solvent extraction should prove to be easier than that of
any other naturally occurring element.
Uranium, vanadium, and molybdenum (the. latter being a common
ore constituent) a J most always appear In aqueous solutions us
oxidized Lons (uranyl, vanadyl , or molylulale radicals), wll.li
uranium and vanadium additionally cuinplcxc.il with anlonlc radicals
to form trisul fates or trlcarbonates In Lhc leach. The com-
plexes react an Ionic-ally, and the affinity n| exchange, resins
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and solvents is not simply related to fundamental properties
of the heavy metal (U, V, or Mo), as is the case in cationic-
exchange reactions. Secondary properties, including pH and
reduction/oxidation potential, of the pregnant solutions
influence the adsorption of heavy metals. For example, seven
times more vanadium than uranium was adsorbed on one resin
at pH 9; at pK 11, the ratio was reversed, with 33 times as
much uranium as vanadium being captured. These variations
in affinity, multiple columns, and control of leaching time
with respect to breakthrough (the time when the interface
between loaded and regenerated resin arrives at the end of
the column) are used to make an ion-exchange process specific
for the desired product.
In the case of solvent extraction, the type of polar solvent
and its concentration in a typically nonpolar diluent (e.g.,
kerosene) affect separation of the desired product.
The ease with which the solvent is handled permits the con-
struction of multistage, cocurrent and countercurrent, solvent-
extraction concentrators that are useful even when each stage
effects only partial separation of a value from an interferent.
Unfortunately, the solvents are easily polluted by slimes,
and complete liquid/solid separation is necessary. Ion-
exchange and solvent-extraction circuits can be combined to take
advantage of the slime resistance of resin-in-pulp ion exchange
and of the separatory efficiency of solvent extraction (Eluex
process).
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GENERAL DESCRIPTION OF INDUSTRY BY ORE CATEGORY
The ore groups categorized in SIC groups 1011, 1021, 1031,
1041, 1044, 1051, 1061, 1092, 1094, and 1099 vary considerably
in terms of their occurrence, mineralogy arid mineralogical
variations, extraction methods, and end-product uses. For
these reasons, these Industry areas generally are treated
separately except for groups SIC 1061, Ferroalloys (members
of which are differently occurring ore minerals but are
classed as one group), and SIC 1099, Metal Ores, Not Elsewhere
Classified (a grouping of ore minerals whose mining and pro-
cessing operations bear little resemblance to one another).
Iron Ore
American iron-ore shipments increased from 82,718,400 metric
tons (91,200,000 short tons) in 1968 to 92,278,180 metric tons
(101,740,000 short tons) in 1973, an increase of 11.56% (Refer-
ence 1). In this period, the shipments of agglomerates, most of
which were produced by processing low-grade iron formations,
increased by 19.1%. Total consumption of iron ore in the
United States in 1973 was 139,242,640 metric tons (153,520,000
short tons), with 76.5% produced domestically. Domestic
agglomerates accounted for 66,256,350 metric tons (73,050,000
short tons), or 47.6% of United States consumption. A summary
of U.S. iron-ore shipments is shown in Table III-l. A break-
down of crude iron-ore production in the U.S. is shown in
Table III-2. A breakdown of U.S. iron-ore shipments by pro-
ducing company is given in Supplement B to this document.
Except for a very small tonnage, iron ores are beneficiated
before shipping.
Beneficlation of iron ore includes such operations as crush-
ing, screening, blending, grinding, concentrating, classify-
ing, and agglomerating and is most often carried on at or
near the mine site. Methods selected are based on physical
and chemical properties of the crude ore. A noticeable trend
has been developing in furthering efforts to use lower-grade
ores and to recover more of the secondary minerals in an ore
deposit. As with many other natural resources, future avail-
ability will largely be a matter of cost rather than of abso-
lute depletion as these lower-grade ores are utilized. Bene-
fication methods have been developed to upgrade 20-30% iron
'taconite' ores into high-grade materials.
In most cases, open-pit mining is more economical than con-
ventional underground methods. It provides the lowest cost
operation and is employed whenever the ratio of overburden
(either consolidated or unconsolidated) to ore does not
111-20
DRAFT
-------
DRAFT
TABLE 111-1. IRON-ORE SHIPMENTS FOR UNITED STATES
a. QUANTITIES SHIPPED BY REGION
REGION
Great Lakes
NortfiBBStB rn
Southern
•Hf--»-— —
YYMiarn
TOTAL U.S.
REGION
Groat Lakes
Northeastern
Southern
UUAM*IUM
VW5I6fn
TOTAL U.S.
AMOUNT SHIPPED
1968
METRIC TONS
66.093,239
3.602.706
3.474.203
10.666.860
82.736.806
LONG TONS
64,065.185
3.645.805
3.419.333
10.399.872
81.430.195
1969
METRIC TONS
72.S34.630
3.453.486
4.733.087
10.464.364
91.176.567
LONG TONS
71,389.050
3.398.943
4.658.335
10.289.252
89.735.580
1970
METRIC TONS
70.180.666
3.043357
5.022.36B
10.644.782
88.791.674
LONG TONS
69,072.263
2,995,784
4.943.048
10.378.242
87.389.337
AMOUNT SHIPPED
1971
METRIC TONS
62.766.873
2.B59.973
4.240,720
8,253,243
78.120,810
LONG TONS
61.775.561
2.814304
4.173.744
8.122,895
76.887,004
1972
METRIC TONS
65.769.357
2.362.067
4.032,651
7,397,815
79.537,162
LONG TONS
64.720.783
2.324,762
3.966.961
7,266,471
78,280.977
1873
METRIC TONS
77.604365
2,405,456
3.923,518
8.462,679
92.296,418
LONG TONS
76.280,787
2.367,465
3,861 ,562
8,328,925
90,838.729
b. SHIPMENTS FROM GREAT LAKES REGION AS PERCENTAGES OF TOTAL US. SHIPMENTS
YEAH
1968
1869
1970
1971
1972
1973
GREAT LAKES SHIPMENTS
AS PERCENTAGE OF
TOTAL U.S. SHIPMENTS
78.7
79.6
79.0
80.4
82.7
84.0
AGGLOMERATES AS
PERCENTAGE OF
GREAT LAKES SHIPMENTS
61.9
63.6
66.2
70.1
74.8
73.6
GREAT LAKES AGGLOMERATES
AS PERCENTAGE OF TOTAL
U.S. SHIPMENTS
48.7
60.6
62.3
66.3
61.8
61.7
c. PERCENTAGES OF TOTAL U.S. SHIPMENTS
CATEGORY
Direct Shipping
CoaneOres
Fine Ores
Screened Ores
Concentrates
Agglomeratei
YEAR
1968
8.2
3.2
283
60.3
100.0
1968
7.0
3.1
27.6
62.4
100.0
1670
6.0
27
28.2
64.1
100.0
1971
4.3
3.1
23.7
689
100.0
1972
2.0
12.8
11.9
73.3
100.0
1973
2.4
129
129
71.8
100.0
SOURCE: Reference 1
111-21
DRAFT
-------
DRAFT
TABLE II1-2. CRUDE IRON-ORE PRODUCTION FOR U.S.
a. QUANTITIES PRODUCED
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION BY REGION
GREAT LAKES
METRIC TONS
169.349.027
169.328.625
172.799.698
161.947.509
1 68,183.907
186,627.840
LONG TONS
166.832.339
166.654.225
170.070.772
159.389.781
155.685.620
183.680.322
NORTHEASTERN
METRIC TONS
10.236.712
9.728.661
9.173.800
7.774.210
6.721.672
6.916.338
LONG TONS
10.075.038
9.575.01 1
9.028.913
7.651.428
6.615.513
6.806.120
SOUTHERN
METRIC TONS
7.743.542
9.135.961
10.542.987
9.414.016
9.333.043
8.629.278
LONG TONS
7.621.244
8.991.662
10.376.387
9.265.335
9.185.641
8.492,991
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION BY REGION
WESTERN
METRIC TONS
19,671.003
19.270.778
19.981.771
18.422,861
13.347.447
18.080.995
LONG TONS
19.360.328
18.966.424
19.666.188
18.131.898
13,136.643
17.795.432
TOTAL U.S. PRODUCTION
METRIC TONS
197,000.285
207463.916
212.498.366
197558,696
187.686.069
220,263,451
LONG TONS
193.888.949
204.187.322
209.142.260
194.438.442
184.623.417
216.774.865
b. PERCENTAGE OF U.S. CRUDE IRON-ORE PRODUCTION
REGION
Great Lakes
Northeastern
Southern
Western
YEAR
1968
80.9
5 1
4.0
10.0
100.0
1969
81.6
4.7
4.4
9.3
100.0
1970
81.3
4.3
6.0
9.4
1000
1971
82.0
3.9
4.8
93
100.0
1972
84.3
3.6
6.0
7.1
1000
1973
84.7
32
39
82
1000
SOURCE. Reference 1
111-22
DRAFT
-------
DRAFT
exceed an economical limit. The depth to which open pit
mining can be carried depends on the nature of the overburden
and the stripping ratio (volume of overburden/crude ore).
Economic stripping ratios vary widely from mine to mine and
from district to district, depending upon a number of factors.
In the case of direct shipping ores, it may be as high as 6
or 7 to 1; in the case of taconite, a stripping ratio of less
than 1/2 to 1 may become necessary. Stripping the overburden
necessitates continually cutting back the pit walls to permit
deepening of the mine to recover ore in the bottom. Power
shovels, draglines, power scrapers, hydraulicking, and hydraulic
dredging are used to recover ore deposits. Drilling and blast-
ing may be necessary to remove consolidated overburden and to
loosen ore banks directly ahead of power shovels. Iron ore
is loaded into buckets ranging in size from 0.75 to 7.5 cubic
meters (1 to 10 cubic yards). The ore is transported out of
the pit by railroad cars, trucks, truck trailers, belt con-
veyors, skip hoists, or a combination of these. It is then
transferred to a loading dock for hauling to a crushing plant
for size reduction, to a screening plant for sizing, or to a
concentrating plant for treatment by washing (wet size classi-
fication and tailings rejection) or by gravity separation.
Special problems are associated with the mining of taconite.
The extreme hardness of the ore necessitates additional drill-
ing/blasting operations and specialized, more rugged equipment.
The low iron content makes it necessary to handle two or four
times as much mined material to obtain a given quantity of
iron as compared to higher grade ore deposits.
Water can cause a variety of problems if allowed to collect
in mine workings. Therefore, means must be developed to
collect water and pump it out of the mine. This drainage
water is often used directly to make up for water losses in
concentration operations.
Underground methods are utilized only when stripping ratios
become too high for economical open pit mining. Mining tech-
niques consist of sinking vertical shafts adjacent to the
deposit but far enough away to avoid the effects of surface
subsidence resulting from mining operations. Construction
of shafts, tunnels, underground haulage and development work-
Ings, and elaborate pumping facilities usually requires expen-
sive capital Investments. Production in terms of iron ore/day
is much lower than in the case of open pit production, neces-
sitating the presence of very high grade ores for economic
recovery. The health, and safety of operators and the ecological
Impacts of underground mining operations are of increasing concern
to the public, while pressures from the public sector have recently
been forcing Important changes in mine operations and planning.
111-23
DRAFT
-------
DRAFT
General techniques utilized in the beneficiation of iron ore
are illustrated in Figure III-l. Processes enhance either
the chemical or physical characteristics of the crude ore
to make more desirable feed for the blast furnace.
Crushing and screening reduce the size of crude ore not
requiring processing in order to eliminate handling problems
and to increase heat transfer and, hence, rate of reduction
in the blast furnace. Blending produces a more uniform prod-
uct to comply with blast furnace requirements.
Physical concentrating processes such as washing remove un-
wanted sand, clay, or rock from crushed or screened ore.
For those ores not amenable to simple washing operations,
other physical methods such as Jigging, heavy-media separa-
tion, flotation, and magnetic separation are used. Jigging
Involves stratification of ore and gangue by pulsating water
currents. Heavy-media separation employs a water suspension
of ferrosilicon in which iron ore particles sink while the
majority of gangue (quartz, etc.) floats. Air bubbles attracted
to ores conditioned with flotation reagents separate out iron
ore in the froth during the flotation process, while magnetic
separation techniques are used where ores containing magnetite
are encountered.
At the present time, there are only three chemical process-
ing iron ore flotation plants in the United States. Figure
III-2 illustrates a typical flowsheet used in an iron ore
flotation circuit, while Table III-3 lists types and amounts
of flotation reagents used per ton of ore processed. Various
flotation methods which utilize these reagents are listed in
Table II1-4. The most commonly adopted flowsheet for the
beneficiation of low grade magnetic taconite ores is illus-
trated in Figure III-3. Low grade ores containing magnetite
are very susceptible to concentrating processes, yielding a
high quality blast furnace feed. Initially higher grade ores
containing hematite, on the other hand, while containing more
iron cannot be upgraded much above 55% iron.
Agglomerating processes follow concentration operations and
serve to Increase permeability of furnace feed and reduce
"fines" which normally would be lost in the flue gases.
Sintering, pelletizing, briquetting, and nodullzlng are all
possible operations involved in agglomeration. Sintering
involves the mixing of small portions of coke and limestone
with the iron ore, followed by combustion. A granular,
coarse, porous product is formed, with a reduction Ln impuri-
ties. Pelletizing involves the formation of pellets or ballh
of iron ore fines, followed by heating. (Figure IIT-4 illus-
trates a typical palletizing operation.) Nodules or lumps are
111-24
DRAFT
-------
DRAFT
Figure 111-1. BENEFICIATION OF IRON ORES
ORE
CRUSHING AND
SCREENING
*
BLENDING
t
CONCENTRATING PROCESSES:
PHYSICAL
* i
WASHING' JIGGIN
1
| SINTERING 1
* *
r MAGNETIC VJfpn
G SEPARATION SEpM*°
*
AGGLOMERATION
PROCESSES
PELLETIZING NODULIZING
CHEMICAL
*
'Y-
A FLOTATION
TION
*
1
BRIQUETTING
t
T
TO STOCK PILE AND/OR SHIPPING
111-25
DRAFT
-------
DRAFT
Figure 111-2. IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
DENSIFYING THICKENER
UNDERFLOW
CONDITIONERS
I
ROUGHER FLOTATION
ROUGHER
TAIL
TO
TAILING
BASIN
ROUGHER
CONCENTRATE
(10 CELLS)
1
FROTH OF
'FIRST 2 CELLS
I
CLEANER FLOTATION
CLEANER
TAIL
I
CLEANER
CONCENTRATE
(8 CELLS)
FROTH OF
FIRST 2 CELLS
1
RECLEANER FLOTATION
RECLEANER
TAIL
RECLEANER
CONCENTRATE
(7 CELLS)
I
TOTAL
FLOTATION
CONCENTRATE
J
TO AGGLOMERATION
(FIGURE 111-4)
111-26
DRAFT
-------
DRAFT
TABLE 111-3. REAGENTS USED FOR FLOTATION OF IRON ORES
(Reagent quantities represent approximate maximum usages. Exact chemical composition of reagent
may be unknown.)
1. Anionlc Flotation of Iron Oxides (from crude ore)
Petroleum sulfonate: 0.5 kg/metric ton (1 Ib/short ton)
Low-rosin, tall oil fatty acid: 0.25 kg/metric ton (0.5 Ib/short ton)
Sulfuric acid: 1.25 kg/metric ton (2.5 Ib/short ton) to pH3
No. 2 fuel oil: 0.15 kg/metric ton (0.3 Ib/short ton)
Sodium silicate: 0.5 kg/metric ton (1 Ib/short ton)
2. Anionlc Flotation of Iron Oxides (from crude ore)
Low-rosin tall oil fatty acid: 0.5 kg/metric ton (1 Ib/short ton)
3. Cetionlc Flotation of Hematite (from crude ore)
Rosin amine acetate: 0.2 kg/metric ton (0.4 Ib/short ton)
Sulfuric acid: 0.15 kg/metric ton (0.3 Ib/short ton)
Sodium fluoride: 0.15 kg/metric ton (0.3 Ib/short ton)
(Plant also includes phosphate flotation and pyrite flotation steps. Phosphate flotation employs
sodium hydroxide, tall oil fatty acid, fuel oil. and sodium silicate. Pyrita flotation employs
xanthate collector.)
4. Cationic Flotation of Silica (from crude ore)
Amine: 0.15 kg/metric ton (0.3 Ib/short ton)
Gum or starch (tapioca fluor): 0.5 kg/metric ton (1 Ib/short ton)
Methylisobutyl carbinol: as required
5. Cationic Flotation of Silica (from magnetite concentrate)
Amine: 5 g/metric ton (0.01 Ib/short ton)
Methylisobutyl carbinol: as required
111-27
DRAFT
-------
DRAFT
TABLE 111-4. VARIOUS FLOTATION METHODS AVAILABLE FOR PRODUCTION
OF HIGH-GRADE IRON-ORE CONCENTRATES
1. Anionic flotation of specular hematites
2. Upgrading of natural magnetite concentrated by cationic flotation
3. Upgrading of artificial magnetite concentrates by cationic flotation
4. Cationic flotation of crude magnetites
5. Anionic flotation of silica from natural hematites
6. Cationic flotation of silica from non-magnetic iron formations
III-28
DRAFT
-------
DRAFT
Figure 111-3. MAGNETIC TACONITE BENEFICIATION FLOWSHEET
CRUSHED CRUDE ORE
T _
ROD MILL |
COBBER MAGNETIC SEPARATION
CONCENTRATE
i
| BALL MILL |
V
CLEANER MAGNETIC SEPARATION
CONCENTRATE
I
HYDROCYCLONE
OVERSIZE UNDERSIZE
HYDROSEPARATOR
CONCENTRATE
FINISHER MAGNETIC SEPARATION
CONCENTRATE
THICKENING
I
LTI
T
TO PELLETIZING
(FIGURE MM)
TAILING
TAILING
TAILING
TAILING
TO TAILING BASIN
111-29
DRAFT
-------
DRAFT
Figure 111-4. AGGLOMERATION FLOWSHEET
CONCENTRATE FILTER CAKE
BENTONITE
BALLING DRUM
1
SCREEN
I
UNDERSIZE OVERSIZE
I
FUEL
I
AGGLOMERATION FURNACE
PELLETS EXHAUST GASES
t *
TO STOCK PILE TO ATMOSPHERE
AND/OR SHIPPING
III-JO
DRAFT
-------
DRAFT
formed when ores are charged into a rotary kiln and heated
to incipient fusion temperatures in the nodulizing process.
Hot ore briquetting requires no binder, is less sensitive
to changes in feed composition, requires little or no grind-
ing and requires less fuel than sintering. Small or large
lumps of regular shape are formed.
Copper Ore
The copper ore segment of the ore mining and dressing indus-
try includes facilities mining copper from open pit and
underground mines, and those beneficiating the ores and
wastes by hydrometallurgical and/or physical-chemical pro-
cesses. Other operations for processing concentrate and
cement copper, and for manufacturing copper products (such
as smelting, refining, rolling, and drawing) are classified
under other SIC codes and are covered under limitations and
guidelines for those industry classifications. However,
to present a comprehensive view of the history and statistics
of the copper production in the United States, statistics
pertaining to finished copper are included with those for
ore production and beneficiation.
Evidence of the first mining of copper in North America,
in the Upper Peninsula of Michigan, has been found by
archeologists. Copper was first produced in the colonies
at Simsbury, Connecticut, in 1709. In 1820, a copper ore
body was found in Orange County, Vermont. In the early
1840's, ore deposits located in Northern Michigan accounted
for extensive copper production in the United States. Other
discoveries followed in Montana (1860), Arizona (1880), and
Bingham Canyon, Utah (1906). Since 1883, the United States
has led copper production in the world. As indicated by
the tabulation which follows, seven states presently produce
essentially all of the copper mined in the U.S. (See also
Figure III-5.)
Arizona
Utah
New Mexico
Nevada
MLchlgan
Tennessee
A series of tables follow which give statistics for the U.S.
copper industry. Table III-5 lists total copper mine produc-
tion of ore by year, and Table III-6 gives copper ore produc-
tion by state for 1972. The average copper content of domestic
111-31
DRAFT
-------
Figure 111-5. MAJOR COPPER MINING AND MILLING ZONES OF THE U.S.
U!
to
MINING AND MILLING COPPER AS A PRIMARY METAL • '
|y:::j::| MINING AND MILLING COPPER AS A COPRODUCT
O
33
-------
DRAFT
TABLE 111-5. TOTAL COPPER-MINE PRODUCTION OF ORE BY YEAR
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION
1000 METRIC TONS
154,239
202.943
233,760
220.089
242,016
263.088
1000 SHORT TONS
170.054
223.752
257,729
242.656
266.831
290,000
SOURCE: REFERENCE 2
TABLE 111-6. COPPER-ORE PRODUCTION FROM MINES BY STATE [1972]
STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
TENNESSEE
ALL OTHER
TOTAL U.S.
PRODUCTION
1000 METRIC TONS
150,394
32,250
18,077+
15,531*
12.052*
7.483
1,598
< 4.631
242.016
1000 SHORT TONS
165,815
35,557
19,930+
17,126+
13.288+
8,250
1.762
< 5,106
266,831
SOURCE: REFERENCE 2
III-'JJ
DRAFT
-------
DRAFT
ores is given by Table III-7. The average concentration of
copper recovered from domestic ores, classified by extraction
process, is listed in Table III-8. Copper concentrate produc-
tion by froth flotation is given in Table III-9, while pro-
duction of copper concentrate by major producers in 1972 is
given as part of Supplement B.
Twenty-five mines account for 95% of the U.S. copper output,
with more than 50% of this output produced by three companies
at five mines. Approximately 90% of present reserves (77.5
million metric tons, 85.5 million short tons, of copper metal
as ore) average 0.86% copper and are contained in five states:
Arizona, Montana, Utah, New Mexico, and Michigan. Mining
produced 154 million metric tons (170 million short tons) of
copper ore and 444 million metric tons (490 million short tons)
of waste in 1968.
Open pit mines produce 83% of the total copper output with the remain-
der of U.S. production from underground operations. Ten percent of
mined material is treated by dump (heap) and in-situ leaching producing
229,471 metric tons (253,000 short tons) of copper. Recovery of copper
from leach solutions by iron precipitation accounted for 87.5% of the
leaching production. Electrowon copper amounted to 13.5%.
Approximately 98% of the copper ore was sent to concentrators
for beneficiation by froth flotation, a process at least 60
years old. Copper concentrate ranges from 11% to 38% copper
as a result of approximately 83% average recovery from ore.
Secondary or coproduction of other associated metals occurs
with copper mining and beneficiation. For instance, in 1971,
41% of U.S. gold production was as base-metal byproducts.
Fourteen copper plants in 1971 produced molybdenum as well.
From 63.5 million metric tons (70 million short tons) of moly-
bdenum byproduct ore, 18,824 metric tons (20,750 short tons)
of byproduct molybdenum were produced.
Processes Employed to Extract Copper from Ore. The mining
processes employed by the copper industry are open pit or
underground operations. Open pit mining produces step-like
benched tiers of mined areas. Underground mining practice
is usually by block-caving methods.
Beneficiation of copper ores may be hydrometallurgical or
physical-chemical separation from the gangue material. A
general scheme of methods employed for recovery of copper
from ores is given as Figure III-6. Hydromotallurgica.1 pro-
cesses currently employ sulfurlc acid (5-10%) or iron su]fate
to dissolve copper from the oxldlc or mixed oxldlc-sulfidic
IIL-34
DRAFT
-------
DRAFT
TABLE 111-7. AVERAGE COPPER CONTENT OF DOMESTIC ORE
YEAR
1968
1969
1970
1971
1972
1973
PERCENT COPPER
0.60
0.60
0.59
0.55
0.55
0.53
SOURCE: REFERENCE 2
TABLE 111-8. AVERAGE CONCENTRATION OF COPPER IN DOMESTIC
BY PROCESS (1972)
ORES
STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
IDAHO
TENNESSEE"
COLORADO
ALL OTHER
TOTAL U.S.
CONCENTRATION (%)
FLOTATION*
0.51
0.58
0.70
0.55
0.54
0.82
-
0.64
-
1.35
0.55
DUMP/HEAP
LEACH
0.47
1.10
-
-
0.38
N/A
-
N/A
-
-
0.47
DIRECT SMELTER
FEED
1.94
—
0.07f
4.06
0.68
-
2.65
-
10.24
2.30
1.68
• INCLUDES FROTH FLOTATION AND LEACH-REDUCTION/FLOTATION
•• FROM COPPER/ZINC ORE
t JUST AS A FLUXING MATERIAL
SOURCE: REFERENCE 2
111-35
DRAFT
-------
DRAFT
TABLE 111-9. COPPER ORE CONCENTRATED IN THE UNITED STATES BY
FROTH FLOTATION, INCLUDING LPF PROCESS (1972)
STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
TENNESSEE*
ALL OTHER
TOTAL U.S.
PRODUCTION
1000 METRIC TONS
138.998
31.702
18.019
15,508
12,003
7,483
1.598
228
225.537
1000 SHORT TONS
153.250
34,952
19.866
17,098
13,234
8,250
1,762
251
248.663
FROM COPPER/ZINC ORE
SOURCE: REFERENCE 2
TII - 10
DRAFT
-------
DRAFT
Figure 111-6. GENERAL OUTLINE OF METHODS FOR TYPICAL RECOVERY OF
COPPER FROM ORE
ORE KOIHCul
WASTE
DUMP "«
LEACH ««
IACIDI IAC
ACID ACID
SOLUTION SOL'N
ACID
RECVCLED
1 1
PRECIPI
PLA
ORE 10 144% Cu) ORE ^ OVERBURDEN AND
ORE -'»""« WASTE DUMPS
1 i
AP MMTJJ CRUSHER *j SCREENING
101 IACIDI MINED I
i ORE CRUSHER
ACID 1 1
Ml'M «
RECVCLED RECVCLED J SCREENING
I |
[y,""1 HEAP _^. TERTIARY
HT* [^ CRUSHER MIXED
(, ACID i &ULFID
ACID RECYCLED 'OLUTION [_j ^^
SPONGE IRON
CEMENT 1 "•"•".'"..""" COPPER AND
rnPVFR 1 KUAHI 100N
n
TOSM
1
REFIf
ELTER TAII.IMO 1 , TmrnriiirnT ^. FLOTATION ^ TA
POND P^ THICKENERS ^ CEtLS ••
* 1 t
,ERV RECYCLED WATER «IIHI*»IH»IB
THICKENERS 'j'^
REFINERY
TO SMtLTER
CEMENT
COPPER
t
PRECIPITATION
PLANT
OXIDE/
E ORE
WASH
WATER
VAT LEACH
IACIDI
LS '
BARREN
PF
SI
l
ELECTRO
WINNING
FACIIITV
ARD '"
OW
T SULFIDE
| FILTER | 4
1 TO DUMP
BYPRODUCT COPPEHISI
MOLYBDENUM CONCENTRATE
t TO SMELTER
TO SMELTER |
Y REFINERY
MARKET
t
CATHOOI
COPPER
EGNANT
3LUTION
1UMAKKLI
ion HI IINERVI
111-37
DRAFT
-------
DRAFT
ores in dumps, heaps, vats or in-situ (Table 111-10). Major
copper areas employing heap, dump, and in-situ leaching are
shown in Figure III-7. The copper is then recovered from
solution in a highly pure form by the iron precipitation,
electrolytic deposition (electrowinning), or solvent extrac-
tion-electrowinning process.
Ore may also be concentrated by froth flotation, a process
designed for extraction of copper from sulfide ores. Ore is
crushed and ground to a suitable mesh size and is sent through
flotation cells. Copper sulfide concentrate is lifted in
the froth from the crushed material and collected, thickened,
and filtered. The final concentrate, containing 15-30% copper,
is sent to the smelter for production of blister copper (98%
Cu) . The refinery produces pure copper (99.88-99.9% Cu) from
the blister copper, which retains impurities such as gold,
silver, antimony, lead, arsenic, molybdenum, selenium, tel-
lurium, and iron. These are removed in the refinery.
A combination of the hydrometallurgical and physical-chemical
processes, termed LPF (leach-precipitation-flotation) has
enabled the copper industry to process oxide and sulfide
minerals efficiently. Tailings from the vat leaching process,
if they contain significant sulfide copper, can be sent to
the flotation circuit to float copper sulfide, while the
vat leach solution undergoes iron precipitation or electro-
winning to recover copper dissolved from oxide ores by acid.
A major factor affecting domestic copper production is the
market price of the material. Historically, copper prices
have fluctuated but have generally Increased over the long
term (Table III-ll). Smelter production of copper from
domestic ores has continuously risen and has increased in
excess of a factor of three over the last 68 years (Table
111-12).
Lead and Zinc Ores
Lead and zinc mines and mills in the U.S. range in age from
over one hundred years to essentially new. The size of
these operations ranges from several hundred metric tons
of ore per day to complexes capable of moving about ten
thousand metric tons of ore per day. Lead and zinc ores
are produced almost exclusively from underground mines.
There are some deposits wlilcli are amenable to open pit oper-
ations; a number of minus during Llielr early opening stages
of operation are started «ns open-pit mines and then developed
into underground mines. At present., only one small open-pit
mine is in operation, nnd Its useful life Is estimated in
months. Therefore, for all practical purposes, jll miniiiR
II1-J8
DRAFT
-------
DRAFT
TABLE 111-10. HEAP OR VAT ORE LEACHED IN THE UNITED STATES (1972)
STATE
ARIZONA
UTAH
NEW MEXICO/NEVADA
MONTANA
TOTAL U.S.
PRODUCTION
1000 METRIC TONS
11.071
549
4.400
N/A
16.039
1000 SHORT TONS
12,228
605
4.851
N/A
17.684
SOURCE: REFERENCE 2
111-39
DRAFT
-------
0
3J
--
o
Figure 111-7. MAJOR COPPER AREAS EMPLOYING ACID LEACHING IN HEAPS.
IN DUMPS, OR IN SITU
O
3J
ma LEACHING ZONES
-------
DRAFT
TABLE 111-11. AVERAGE PRICE RECEIVED FROM COPPER
IN THE UNITED STATES
YEAR
PRICE IN CENTS PER KILOGRAM (CENTS PER POUND)
LAKE COPPER*
ELECTROLYTIC COPPERt
1865-1874
1907
1910
1916
1917
1920
1926
1930
1932
1936
1940
1946
1960
1966
1960
1966
1970
1972
1973
60.94 (27.70)
46.86 (21.30)
28.86 (13.12)
38.81 (17.64)
64.20 (29.18)
39.62 (18.01)
31.77 (1444)
29.48 (1340)
13.00 ( 6.91)
19.62 ( 8.92)
26.66(11.66)
26.40 (12.00)
40.96- 64.16(18.62-24.62)
66.00- 94.60(30.00-43.00)
66.00- 72.60(30.00-33.00)
74.80- 83.60(34.00-38.00)
116.6 -132.0(63.00-60.00)
109.7 -114.7(49.88-62.13)
110.3 -169.2(60.13-72.38)
42.90- 63.90(19.60-24.60)
69.30- 94.60(31.50-43.00)
66.00 • (30.00)
77.00- 81.40(35.00-37.00)
116.9 -132.3(63.12-60.12)
111.4 -116.8(60.63-62.63)
116.9 -151.1 (63.13-68.70)
• COPPER FROM NATIVE COPPER MINES OF LAKE SUPERIOR DISTRICT: MINIMUM 99.90%
PURITY, INCLUDING SILVER.
t ELECTROLYTIC COPPER RESULTS FROM ELECTROLYTIC REFINING PROCESSES:
MINIMUM 99.90% PURITY, SILVER COUNTED AS COPPER
SOURCE: REFERENCE 3
111-41
DRAFT
-------
DRAFT
TABLE 111-12. PRODUCTION OF COPPER FROM DOMESTIC ORE
BY SMELTERS
YEAR
1905
1910
1915
1916
1919
1921
1925
1929
1930
1932
1935
1937
1940
1943
1946
1950
1955
1960
1965
1970
1971
1972
1973
ANNUAL PRODUCTION
METRIC TONS
403.064
489.853
629.463
874.280
583,391
229,283
759,554
908.299
632,356
246,709
345,834
757,038
824.539
991.296
543.888
826.596
913,631
1.036.563
1.272.345
1.455.973
1.334,029
1.513,710
1.569.110*
SHORT TONS
444,392
540,080
694,005
963,925
643,210
252,793
837,435
1,001,432
697,195
272,005
381,294
834,661
909,084
1,092.939
599,656
911,352
1.007,311
1,142,848
1,402,806
1,605,262
1,470,815
1,668,920
1,730,000*
•PRELIMINARY BUREAU OF MINES DATA
SOURCE: REFERENCE 3
111-42
DRAFT
-------
DRAFT
can be considered to be underground.
In general, the ores are not rich enough in lead and zinc to
be smelted directly. Normally, the first step in the conver-
sion of ore into metal is the milling process. In some cases,
preliminary gravity separation is practiced prior to the
actual recovery of the minerals of value by froth flotation,
but, in most cases, only froth flotation is utilized. The
general procedure is to Initially crush the ore and then
grind it, in a closed circuit with classifying equipment,
to a size at which the ore minerals are freed from the gangue.
Chemical reagents are then added which, in the presence of
bubbled air, produce selective flotation and separation of
the desired minerals. The flotation milling process can be
rather complex depending upon the ore, its state of oxidation,
the mineral, parent rock, etc. The recovered minerals are
shipped in the form of concentrates for reduction to the
respective metals recovered.
The most common lead mineral mined in the U.S. Is galena (lead
sulflde). This mineral is often associated with zinc, silver,
gold, and iron minerals.
The principal zinc ore mineral is zinc sulfide (sphalerite).
There are, however, numerous other minerals which contain zinc.
The more common include zlncite (zinc oxide), zinc willemite
(silicate), and franklinite (an iron, zinc, manganese oxide
complex). Sphalerite is often found in association with sul-
fides of iron and lead. Other elements often found in associa-
tion with sphalerite Include copper, gold, silver, and cadmium.
Mine production of lead increased during 1973 and 1974, as
illustrated in Table 111-13, which has been modified from
the Mineral Industry Surveys, U.S. Department of the Interior,
Bureau of Mines, Mineral Supply Bulletin (Reference 4).
Missouri was the foremost state with 80.78% of the total
United States production, followed by Idaho with 10.24%,
Colorado with 4.66%, Utah with 2.28%, and other states with
the remaining 2.04%. This same trend continues with the pre-
liminary figures for 1974 for the period of January through
June. Based on this information and the estimated 60-year
life for the lead ores in the "Viburnum Trend" of the "New
Lead Belt" of southeast Missouri, it Is likely that this jreu
will be the predominant lead source for many years to come.
Mine production of zinc during 1973 and preliminary production
figures for December and January 1974 and January through May
1974 are presented in Table 111-14, which has been modified
from the Mineral Industry Surveys, U.S. Department of Interior,
111-43
DRAFT
-------
DRAFT
TABLE 111-13. MINE PRODUCTION OF RECOVERABLE LEAD
IN THE UNITED STATES
STATE
Alaska
Arizona
California
Colorado
Idaho
Illinois
Maine
Missouri
Montana
New Mexico
New York
Utah
Virginia
Washington
Wisconsin
Other States
1973
RANK
3
2
1
4
%
4.66
12.24
80.78
2.28
Total
Daily average*
1973
JAN.-OEC.
METRIC TONS
S
692
40
25.497
56.002
491
185
441.839
160
2.318
2,090
12,456
2.392
2.011
765
-
546.943
1.498
SHORT TONS
6
763
44
28.112
61,744
541
204
487,143
176
2.556
2.304
13,733
2.637
2.217
844
—
603.024
1,652
1974 (PRELIMINARY)
JAN JUNE
METRIC TONS
...
357
11
11.317
25.667
122
98
251371
51
1.078
1.331
5,674
1.359
443
596
486
300.163
1.658
SHORT IONS
394
12
12.478
28.299
135
108
277.366
56
1,189
1.467
6.256
1.499
489
657
536
330.941
1.828
"Based on number of days in month without adjustment for Sundays or holidays.
lit-44
DRAFT
-------
DRAFT
TABLE 111-14. MINE PRODUCTION OF RECOVERABLE ZINC
IN THE UNITED STATES (PRELIMINARY)
STATE
Arizona
California
Colorado
Idaho
Illinois
Kantueky
Main*
Missouri
Montana
New Jartay
New Mexico
New York
Pennsylvania
Tennessee
Utah
Virginia
Washington
Wisconsin
1973
RANK
4
5
7
1
B
2
3
9
8
%
11.94
9.65
4.13
17.27
6.94
17.4
13.32
3.48
331
Total
Daily average*
1973
JAN.-OEC. TOTALS
METRIC TONS
7.638
16
51.533
41.216
4323
245
17343
74376
379
29,955
11.147
73361
17.104
57.474
15323
15.131
5,768
7365
431399
1.183
SHORT TONS
8,421
18
66317
45,442
5.318
270
19.672
82,223
418
33327
12,290
81,435
18358
63,367
16364
16.682
6.359
8.672
475.853
1.304
1974
JAN. TOTALS
METRIC TONS
600
-
3.961
3,279
224
-
1,238
6389
82
2.361
863
6361
1,575
7,239
1,130
1.281
528
733
38.644
1,246
SHORT TONS
662
-
4,367
3,615
247
...
1.365
7.266
90
2.603
951
7,675
1,737
7381
1.246
1.412
582
808
42.606
1.374
"Based on number of days in month without ad|ustment for Sundays or holidays.
111-45
DRAFT
-------
DRAFT
Bureau of Mines, Mineral Supply Bulletins.
The mine production figures by state for zinc in 1973, how-
ever, are misleading, because Tennessee was ranked third
due to prolonged strikes, the replacement of some older
mine-mills, and the development and construction of new
production facilities. Therefore, note that Tennessee led
the nation in the production of zinc for 15 consecutive
years (until 1973) and should regain the number one ranking
back from Missouri (1973), based on the preliminary produc-
tion figures given for the first half of 1973.
Description of Lead/Zinc Mining and Milling Processes. The
recovery of useful lead/zinc minerals involves the removal
of ores containing these minerals from the earth (mining)
and the subsequent separation of the useful mineral from
the gangue material (concentration). A generalized flow
sheet for such a mine/mill operation is presented in Figure
III-8.
Mine Operations. The mining of lead- and zinc-bearing ores
is generally accomplished in underground mines. The mineral-
containing formation is usually fractured utilizing explosives
such as ammonium nitrate-fuel oil (AN-FO) or slurry gels,
placed In holes drilled in the formation. After blasting,
the rock fragments are transported to the mine shaft where
they are lifted up the shaft in hoppers. Primary
or rough crushing equipment is often operated underground.
The drilling and transportation equipment is, of course,
highly mechanized and employs diesel power.. At some locations,
the equipment is maintained in underground shops, constructed
in tnined-out areas of the workings.
Water enters a mine naturally when aquifers are intercepted;
in highly fractured and fissured formations, water from the
surface may seep into the mine. Minor amounts of water
are introduced from the surface by evaporation of cooling
water and through water expired by workers. At some loca-
tions, water enters with sand or tailings used in hydraulic
backfill operations.
The water is pumped from the mine at a rate necessary Lo
maintain operations in the mine. The amount of water puinpcJ
does not bear any necessary relationship to the output of
ore or mineral. The amount pumped may vary from thousands
of liters per day to 120 to 160 miJJIon liters (JO to 40
million gallons) per day. In many rases, llirre Is
-------
DRAFT
Figure 111-8. LEAD/ZINC-ORE MINING AND PROCESSING OPERATIONS
| ORE MINING | DRAINAGE
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS •
FUELS
LUBRICANTS
TO POND
> AND/OR
MILL
WATER FROM MINE.
RECYCLE OR OTHER
REAGENTS
CONCENTRATE
FINAL LEAD
CONCENTRATE
THICKENING
AND FILTRATION
USUALLY
RECYCLED
TO PROCESS
WATER SYSTEM
1
TAILING ZINC "OUGHER
Wl> CONCENTRATE
i *
TAILING
THICKENER
ZINC CLEANER
FLOTATION
1
TO TAILING FINAL ZINC
DAM CONCENTRATE
r
o ,
VCLE
t
~1
THICKENING
AND FILTRATION
' EFFLUENT 4
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS
EXCESS REAGENTS
I
TO SUBSURFACE
DRAINAGE
CONCENTRATE
TO ZINC
SMELTER
USUALLY RECYCLED
TO PROCESS
WATER SYSTEM
111-47
DRAFT
-------
DRAFT
The water pumped from a mine may contain fuel, oil,
and hydraulic fluid from spills and leaks, and, perhaps,
blasting agents and partially oxidized blasting agents.
The water, most certainly, will contain dissolved solids
and suspended solids generated by the mining operations.
The dissolved and suspended solids may consist of lead,
zinc, and associated minerals.
Milling Operations. The valuable lead/zinc minerals are
recovered from the ore brought from the mine by
froth flotation. In some cases, the ore is precon-
centrated using mechanical devices based on specific gravity
principles. The ore or preconcentrate is initially crushed
to a size suitable for introduction into fine grinding equip-
ment, such as rod mills and ball mills. These mills run wet
and are usually run in circuit with rake or cyclone classi-
fers to recycle to the mill material which is coarser than
the level required to liberate the mineral particles. The
fineness of grind is dependent on the degree of dissemination
of the mineral in the host rock. The ore is ground to a size
which provides an economic balance between the additional
metal values recovered versus the cost of grinding.
In some cases, the reagents used in the flotation process
are added in the mill; in other cases, the fine material
from the mill flows to a conditioner (mixing tank), where
the reagents are added. The particular reagents utilized
are a function of the mineral concentrates to be recovered.
The specific choice of reagents at a facility is usually
the result of determining empirically which reagents result
in an economic optimum of recovered mineral values which
reagents result in an economic optimum of recovered mineral
values versus reagent costs. In general, lead and zinc as
well as copper sulfide flotations are run at elevated pH
(8.5 to 11, generally) levels so that frequent pH adjustments
with hydrated lime (CaOH^) are common. Other reagents
commonly used and their purposes are:
Reagent Purpose
Methyl Isobutyl-carbinol Frother
Propylene Glycol Methyl Ether Frother
Long-Chain Aliphatic Alcohols Frother
Pine Oil Frother
Potassium Amyl Xanthate Collector
Sodium Isopropol Xanthate Collector
Sodium Ethyl Xantliate Collector
Oixanthogen Collector
TII-48
DRAFT
-------
DRAFT
Reagent Purpose
Isopropyl Ethyl Thionocarbonate Frother
Sodium Dlethyl-dlthiophosphate Frother
Zinc Sulfate Zinc Depressant
Sodium Cyanide Zinc Depressant
Copper Sulfate Zinc Actlvant
Sodium Dichrornate Lead Depressant
Sulfur Dioxide Lead Depressant
Starch Lead Depressant
Lime pH Adjustment
The finely ground ore slurry Is Introduced Into a series of
flotation cells, where the slurry Is agitated and air Is Introduced.
The minerals which are to be recovered have been rendered hydropho-
blc (Incapable of dissolution In water) by surface coating with
appropriate reagents. Usually, several cells are operated In a
countercurrent flow pattern, with the final concentrate being floated
off the last cell (cleaner) and the tails taken over the first
or rougher cells. In some cases, regrinding Is used on the
underflow for the cleaner cells to Improve recovery.
In many cases, more than one mineral Is recovered. In such
cases, differential flotation is practiced. The flow shown
in Figure III-8 is typical of such a differential flotation
process for recovery of lead and zinc sulfides. Chemicals
which induce hydrophilic (affinity for water) behavior by surface
interaction are added to prevent one of the minerals from floating
in the initial separation. The underflow of tailings from this
separation is then treated with a chemical which overcomes the
depressing effect and allows the flotation of the other mineral.
After the recovery of the desirable minerals, a large volume
of tailings or gangue material remains as the underflow from
the last rougher cell in the flow scheme. These tails are
typically adjusted to a slurry suitable for hydraulic trans-
port to the treatment facility, termed a tailings pond. In
some cases, the coarse tailings are separated using a cyclone
separator and pumped to the mine for backfilling.
The floated concentrates are dewatered (usually by thickening
and filtration), and the final concentrate—which contains
some residual water—is eventually shipped to a smelter for
metal recovery. The liquid overflow from the concentrate
thickeners is typically recycled in the mill.
The waste stream from a lead/zinc flotation ml]1
contains the residual solids from the original ore which
have been finely ground to allow mineral recovery. The
111-49
DRAFT
-------
DRAFT
stream also contains dissolved solids and excess mill rea-
gents. In cases where the mineral content of the ore varies,
excess reagents will undoubtedly be present when the ore
grade drops suddenly, and lead and zinc will escape with
the tails if high-grade ore creates a reagent-starved system.
Spills of the chemical used are another source of adverse
discharges from a mill.
Gold Ore
The gold ore mining and milling industry is defined for this
document as that segment of the industry involved in the
mining and/or milling of ore for the primary or byproduct/
coproduct recovery of gold. In the United States, this indus-
try is concentrated in eight states: Alaska, Montana, New
Mexico, Arizona, Utah, Colorado, Nevada, and South Dzikota.
Domestic production of gold for 1972 was 45.1 million grams
(1.45 million troy ounces). Of this, approximately 76% come
from four producers, while the 25 leading producers accounted
for 98% of production. The domestic production of gold has
been on a downward trend for the last 20 years, largely as a
result of reduction in the average grade of ore being mined,
ore depletions at some mines, and a labor strike at the major
producer during 1972. However, large increases in the free
market price of gold during recent years (approximately $70
in 1972 to nearly $200 in 1974) has stimulated a widespread
increase in prospecting and exploration activity. As a result
of this, the recovery of gold from low-grade ore may now
become economically feasible, and an increase in production
might be expected in the near future.
Mining Practices. Gold is mined from two types of deposits:
placers and lode or vein deposits. Placer mining consists of
excavating gold-bearing gravel and sands. This Is currently
done primarily by dredging but, in the past, has Included
hydraulic mining and drift mining of buried placers too deep
to strip. Lode deposits are mined by cither underground or
open-pit methods, the particular method chosen depending on
such factors as size and shape of the deposit, ore grade,
physical and mineralogical character of the ore and surround-
ing rock, and depth of the deposit.
Milling Practices. Milling pr.icliccs for Lhc recovery anil
benef id at ion of gold and goLd-i/nuld i" ing ores .iro ry.inul.i-
tion, amalgamation, flotntlun, ;mn
of ore mined from lode dcpusl Ls. I'l.urr u|>ur;iL Inns, lu'wi'vi-i .
employ only gravity methods, SOIIK-I hw. In con juncL Inn with
amalgamation.
I 1 I-SO
DRAFT
-------
DRAFT
Prior to 1970, amalgamation was the process used to recover
nearly 1/4 of the gold produced domestically. Since that
time, environmental concerns have caused restricted use of
mercury. As a result, the percent of gold produced which
was recovered by the amalgamation process dropped from 20.3%
in 1970 to 0.3% in 1972. At the same time that the use of
amalgamation was decreasing, the use of cyanidation processes
was increasing. In 1970, 36.7% of the gold produced domes-
tically was recovered by cyanidation, and this increased to
54.6% in 1972.
Current practice for the amalgamation process (as used by
a single mill in Colorado) involves crushing and grinding
of the lode ore, gravity separation of the gold-bearing black
sands by jigging, and final concentration of the gold by
batch amalgamation of the sands in a barrel amalgamator. In
the past, amalgamation of lode ore has been performed in
either the grinding mill, on plates, or in special amalga-
mators. Placer gold/silver-bearing gravels are beneficiated
by gravity methods, and, in the past, the precious metal-
bearing sands generally were batch amalgamated in barrel
amalgamators. However, amalgamation in specially designed
sluice boxes was also practiced.
There are basically four methods of cyanidation currently
being used in the United States: heap leaching, vat leaching,
agitation leaching, and the recently developed carbon-in-pulp
process. Heap leaching is a process used primarily for the
recovery of gold from low-grade ores. This is an inexpensive
process and, as a result, has also been used recently to
recover gold from old mine waste dumps. Higher grade ores
are often crushed, ground, and vat leached or agitated/leached
to recover the gold.
In vat leaching, a vat is filled with the ground ore (sands)
slurry, water is allowed to drain off, and the sands are
leached from the top with cyanide, which solubilizes the gold
(Figure III-9). Pregnant cyanide solution is collected from
the bottom of the vat and sent to a holding tank. In agita-
tion leaching, the cyanide solution is added to a ground ore
pulp in thickeners, and the mixture is agitated until solution
of the gold is achieved (Figure 111-10). The cyanide solution
is collected by decanting from the thickeners.
Cyanidation of slimes generated In the course of wee grinding
is currently being done by a recently developed process,
carbon-in-pulp (Figure 111-9). The slimes are mixed with a
cyanide solution in large tanks, and Lhe solubilized gold
cyanide is collected by adsorption onto activated charcoal.
Gold is stripped from the charcoal using a small volume of
111-51
DRAFT
-------
DRAFT
Figure 111-9. CYANIOATION OF GOLD ORE: VAT LEACHING OF SANDS
AND 'CARBON-IN-PULP' PROCESSING OF SLIMES
O HI IINIIIV
lOSMIimt
II 1-51'
DRAFT
-------
DRAFT
Figure 111-10. CYANIDATION OF GOLD ORE: AGITATION/LEACH PROCESS
ORE
CRUSHING
GRINDING
I
CONDITIONING
I
COUNTERCURRENT
LEACHING IN
THICKENERS
PRECIPITATION
OF GOLD FROM
LEACHATE WITH
ZINC DUST
*
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
I
PRECIPITATE
FILTERED AND
THICKENED
1
TO SMELTER
REAGENTS (CN)
BARREN
PULP
TAILING
POND
TAILING-POND
DECANT
RECYCLED
BARREN SOLUTION
RECYCLED
111-53
DRAFT
-------
DRAFT
hot caustic; an electrowinning process is used for final
recovery of the gold in Lhe mill. Bullion is subsequently
produced at a refinery.
Gold in the pregnant cyanide solutions from heap, vat, or
agitate leaching processes is recovered by precipitation
with zinc dust. The precipitate is collected in a filter
press and sent to a smelter for the production of bullion.
Recovery of gold by flotation processes is limited, and less
than 3% of the gold produced in 1972 was recovered in this
manner. This method employs a froth flotation process to
float and collect the gold-containing minerals (Figure III-ll)
The single operation currently using this method further
processes the tailings from the flotation circuit by the
agitation/cyanidation method to recover the residual gold
values.
Silver Ores
The silver ore mining and milling industry is defined for
this document as that segment of industry involved in the
mining and/or milling of ore for the primary or byproduct/
coproduct recovery of silver. Domestic production of silver
for 1972 was 1.158 million kilograms (37,232,922 troy ounces).
Over 38% of this production came from Idaho, and most of
this, from the rich Coeur d'Alene district in the Idaho pan-
handle. The remaining production was attributable to eleven
states: Alaska, Arizona, California, Colorado, Michigan,
Missouri, Montana, Nevada, New Mexico, South Dakota, and
Utah. The 25 leading producers contributed 85% of this
total production, and nine of these operations produced over
one million troy ounces each. During the past ten years,
the annual production of silver has varied from approximately
1 to 1.4 million kilograms (32 to 45 million troy ounces).
Prices have also varied and, during 1972, ranged from a low
of 4.41 cents per gram (137.2 cents per troy ounce) to a
high of 6.54 cents per gram (203.3 cents per troy ounce).
Average price for 1972 was 5.39 cents per gram (167.7 cents
per troy ounce) .
Current domestic production of new silver is derived almost
entirely from exploitation of low-grade and complex primary
sulfide ores. About one-fourth of this production is derived
from ores wherein silver is the chief value and lead, zinc,
and/or copper are valuable byproducts. About three-fourths
of this production is from ores in which lead, zinc, and
copper constitute the principal values, and silver is a
minor but important byproduct. The types, grade, and rela-
tive Importance of the metal sulfide ores from which domestic
111-54
DRAFT
-------
DRAFT
Figure 111-11. FLOTATION OF GOLD-CONTAINING MINERALS WITH RECOVERY OF
RESIDUAL GOLD VALUES BY CYANIDATION
ORE
CRUSHING
GRINDING
L
I
CONDITIONING
*
SELECTIVE
FROTH
FLOTATION
L
t
CONCENTRATE
FILTERED
AND THICKENED
I
TOSME
LTER
BARREN
SOLUTION
RECYCLED
WATER
I1EAACMTC
FLOTATION CIRCUIT
TAILINGS
J
\
LEACHING IN
THICKNERS
j
PRECIPITATION OF
GOLD FROM LEACHATE
WITH ZINC DUST
1
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
\
PRECIPITATE FILTERED
AND THICKENED
BARREN ^ TO TAILING
PULP POND
\
TO SMELTER
111-55
DRAFT
-------
DRAFT
silver is produced are listed in Table 111-15.
Present extractive metallurgy of silver was developed over
a period of more than 100 years. Initially, silver, as the
major product, was recovered from rich, yet simply oxidized
ores by relatively crude methods. As the ores became leaner
and more complex, an improved extractive technology was
developed. Today, silver production is predominantly as a
byproduct, and is largely related to the production of lead,
zinc, and copper from the processing of sulfide ores by froth
flotation and smelting. Free-milling—simple, easily liber-
ated—gold/silver ores, processed by amalgamation and cyani-
dation, now contribute only 1 percent of the domestic silver
produced. Primary sulfide ores, processed by flotation and
smelting, account for 99 percent (Table 111-16).
Selective froth flotation processing can effectively and effi
ciently beneficiate almost any type and grade of sulfide ore.
This process employs various well-developed reagent combina-
tions and conditions to enable the selective recovery of many
different sulfide minerals in separate concentrates of high
quality. The reagents commonly used in the process are
generally classified as collectors, promotors, modifiers,
depressants, activators, and frothing agents. Essentially,
these reagents are used in combination to cause the desired
sulfide mineral to float and be collected in a froth while
the undesired minerals and gangue sink. Practically all
the ores presently milled require fine grinding to liberate
the sulfide minerals from one another and from the gangue
minerals.
A circuit which exemplifies the current practice of froth
flotation for the primary recovery of silver from silver
and complex ores is shown in Figure 111-12. Primary recovery
of silver is largely from the mineral tetrahedrite, (Cu,Fe,
Zn,Ag)J.2Sb4S_3. A tetrahedrite concentrate contains approxi-
mately 25 to 32% copper in addition to the 25.72 to 44.58
kilograms pi-r mi-trie Inn (750 Lo 1300 troy ounce per ton)
of silver. A low-j;r.idr (i.4'5 k« per metric ton; J 00 trny
o* per Lou) .'>M v«T/pvr He i nncenLrntii Is prndufiid .it one mill.
Antimony ni.iy cuiiipr i so up to I H'X. nl the Li-Lralied r i t c conci-ii-
Lrate and ni.iy or m.iv iu»t hi- exl rncLud prlur lo shipment to
j smelte.r.
Various oilier s U viir-eunui Ininj; minerals are re-covered MS
byproducts ol primary copper, load, and/or zinc uporiiL ions.
Wherrj Lhis occur*!., Lhf usu-il pracllce is to ultimately rm-ovi-
tin? silver Irnm Lht- bfisc-muLJl flotation concentrates at Lhe
smelter or rn finery.
r I l-.r»f*
DRAFT
-------
DRAFT
TABLE 111-15. DOMESTIC SILVER PRODUCTION FROM DIFFERENT TYPES OF ORES
TYPE
SILVER
COPPER
LEAD/ZINC/
COPPER
LEAD
ZINC
OTHERS*
SILVER ORE PRODUCTION
1000 METRIC TONS
405.43
187.960.33
35.641.47
7.929.90
1.104.73
1,599.04
1000 SHORT TONS
447
207,233
39,296
8.743
1.218
1.763
GRADE OF SILVER
GRAMS PER
METRIC TON
679.0
2.06
10.29
20.57
3.53
6.86
OUNCES PER
SHORT TON
19.8
0.06
0.3
0.6
0.1
0.2
DOMESTIC
PRODUCTION
(%)
24
32
28
14
< 0.5
1.5
'DERIVED FROM GOLD AND GOLD/SILVER ORE
SOURCE: REFERENCE 2
111-57
DRAFT
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DRAFT
TABLE 111-16. SILVER PRODUCED AT AMALGAMATION AND
CYANIDATION MILLS IN THE U.S. AND
PERCENTAGE OF SILVER RECOVERABLE
FROM ALL SOURCES
YEAR
1968
1969
1970
1971
1972
YEAH
1968
1969
1970
1971
1972
SILVER BULLION AND PRECIPITATES RECOVERABLE BY
AM ALG AM AT ION
KILOGRAMS
2862.2
2605.7
2963.8
30.9
77.4
TROY OUNCES
92,021
83.775
95.287
993
2.490
CYANIDATION
KILOGRAMS
1669.2
1533.8
774.2
3321.4
3110.1
TROY OUNCES
53.666
49,312
24.892
106,785
99.992
SILVER RECOVERABLE FROM ALL SOURCES (%)
AMALGAMATION
028
0.20
0.21
t
0.01
CYANIDATION
0.16
0.11
0.05
0.26
0.27
SMELTING*
99.55
99.68
99.73
99.74
99.72
PLACERS
0.01
0.01
0.01
t
t
'Crude ores and concentrates
tLess than 1/2 unit
SOURCE: REFERENCE 2
I I »-'>»
DRAFT
-------
DRAFT
Figure 111-12. RECOVERY OF SILVER SULFIDE ORE BY FROTH FLOTATION
ORE
NO. 1
FLOTATION CIRCUIT
NO. 2
FLOTATION CIRCUIT
RETREATMENT
CIRCUIT
NO. 3
FLOTATION CIRCUIT
T
FINAL Ag
CONCENTRATE*
FINAL
TAILINGS
'CONTAINS
25.7 TO 44.6 KILOGRAMS PER
METRIC TON
(750-1300 OUNCES PER SHORT TON):
25 TO 32% COPPER
0 TO 18% ANTIMONY
PYRITE
CONCENTRATE
FINAL PYHITE
CONCENTRATE1
CONTAINS 3.43 KILOGRAMS Phil
METRIC TON (100 TROY OUNCLS
PER SHORT TON)
I LI-59
DRAFT
-------
DRAFT
Less than 1 percent of the current domestic production of
silver is recovered by amalgamation or cyanldation processes.
These processes have been described in the discussion of
gold ores of this report.
Bauxite
Bauxite mining for the eventual production of metallurgical
grade alumina occurs near Bauxite, Arkansas, where two pro-
ducers mined approximately 1,855,127 metric tons (2,045,344
tons) of ore in 1973. Both operations are associated with
bauxite refineries (SIC 2819), where purified alumina (A1203)
is produced. Characteristically, only a portion of the
bauxite mined is refined for use in metallurgical smelting,
and one operation reports only about 10 percent of its alumina
is smelted, while the remainder is destined for use as chemi-
cal and refractory grade alumina. A gallium byproduct recovery
operation occurs in association with one bauxite mining and
refining complex.
The domestic bauxite resource began to be tapped about the
turn of the century, and one operation has been mining for
about 75 years. However, the aluminum industry began to
burgeon during World War II, and, almost overnight the demands
for this lightweight metal for aircraft created the large
industry of today. Concurrent with the increase in demand
for aluminum was the startup of large-scale mining operations
by both bauxite producers.
Most bauxite is rained by open-pit methods utilizing draglines,
shovels, and haulers. Stripping ratios of as much as 10
feet of overburden to 1 foot of ore are minable, and a 15-to-l
ratio is considered feasible. Pits of 100 feet in depth are
common, and 200 feet is considered to be the economic limit
for large ore bodies. The pits stand quite well for uncon-
solidated sands and clays, but some slumping does occur.
Underground mining occurs at one Arkansas facility, and this
operation provides the low-silica ore essential to the com-
bination process of refining. Although this type of mining
is relatively costly, it is a viable alternative to the pur-
chase of foreign ores at elevated prices. However, one of
the operations utilizes imported bauxite for blending of
ore grades. Milling of the bauxite ore involves crushing,
ore blending, and grinding in preparation for refining. In
1972, less than 10 percent of the bauxite used for primary
aluminum production was of domestic origin. With the Increas-
ing demand for aluminum, it is expected that the use of
imported alumina and aluminum, as well as bauxite, will
increase. Therefore, the domestic supply of bauxite is
111-60
DRAFT
-------
DRAFT
insufficient to meet present needs of the nine domestic
refineries. Recent price increases in foreign bauxite
supplies aid in assuring the future of domestic bauxite
operations, regardless of the limited national reserves.
The search for potential economic sources of aluminum per-
sists, and many pilot projects have been designed to produce
aluminum. Currently, the most notable attempt to utilize
an alternative source of aluminum Is a 9 metric ton (10 ton)
per day pilot plant which converts alunite, K2A16^0H)12(S04)4,
to alumina through a modified Bayer process, preceded by
roasting and water leaching. The process yields byproduct
sulfuric acid and potassium sulfate as cost credits. Addi-
tionally, the processing of alunite creates no significant
"red mud" (leach residue) environmental problems. Currently
alunite mining is in the exploratory stages, with a commercial-
scale refinery is slated for construction in 1975. Full-scale
mining will entail drilling, blasting, and hauling using
bench mining techniques. From all indications, alunite may
provide an economical new source of aluminum.
Bauxite production in the United States has declined recently
from a peak year in 1970, and preliminary production figures
for 1974 Indicate a continuation of the trend. Production
figures in Table 111-17 indicate total U.S. production of
bauxite, which includes that from mines in Alabama, Georgia,
and Arkansas. These mines also produce bauxite for purposes
other than metallurgical smelting.
Ferroalloy Ores
The ferroalloy ore mining and milling category embraces the
mining and beneficiation of ores of cobalt, chromium, colum-
bium and tantalum, manganese, molybdenum, nickel, and tung-
sten Including crushing, grinding, washing, gravity concen-
tration, flotation, roasting, and leaching. The grouping
of these operations Is based on the use of a portion of their
end product in the production of ferroaJloys (e.g., ferro-
manganese, ferromolybdenum, etc.) and does not reflect any
special similarities among the ores or among the processes
for their recovery and beneficiation. SIC 1061, although
presently including few operations and relatively small
total production, covers a wide spectrum of the mining and
milling Industry as a whole. Sulfide, oxide, silicate, car-
bonate, and anionic ores all are or have been recovered for
the included metals. Open-pit and underground mines are
currently worked, and placer deposits have been mined in
past and are included in present reserves. Beneficiation
111-61
DRAFT
-------
DRAFT
TABLE 111-17. PRODUCTION OF BAUXITE IN THE UNITED STATES
YEAH
1964
1965
1966
1967
1968
1969
1970
1971
1972
1973
1000 METRIC TONS*
1626
1680
1825
1680
1692
1872
2115
2020
1930
1908
1000 SHORT TONS*
1793
1852
2012
1852
1865
2064
2332
2227
2128
2104
•Production, given in dry equivalent weight, includes bauxite mined for
purposes other than metallurgical smelting
111-62
DRAFT
-------
DRAFT
techniques include numerous gravity processes, jigging,
tabling, sink-float, Humphreys spirals; flotation, both
basic-sulfide and fatty-acid; and a variety of ore leach-
ing techniques. Operations vary widely in scale, from
very small mines and mills intermittently worked with
total annual volume measured in hundreds of tons, to two
of the largest mining and milling operations in the country
(Reference 2 ). Geographically, mines and mills in this
category are widely scattered, being found in the southeast,
southwest, northwest, north central, and Rocky Mountain regions
and operate under a wide variety of climatic and topographic
conditions.
Historically, the ferroalloy mining and milling Industry has
undergone sharp fluctuation in response to the prices of
foreign ores, government policies, and production rates of
other metals with which some of the ferroalloy metals are
recovered as byproducts (for example, tin and copper, Refer-
ence 5 )• Many deposits of ferroalloy metals in the U.S. are
of lower grade (or more difficult to concentrate) than foreign
ores and so are only marginally recoverable or uneconomic at
prevailing prices. Large numbers of mines and mills were
worked during World Wars I and II, and during government
stockpiling programs after the war, but have since been
closed. At present, ferroalloy mining and milling is at a
very low level. Increased competition from foreign ores,
the depletion of many of the richer deposits, and a shift
in government policies from stockpiling materials to selling
concentrates from stockpiles have resulted in the closure of
most of the mines and mills active In the late 1950's. For
some of the metals, there is little likelihood of further
mining and milling in the foreseeable future; for others,
increased production in the next few years is probable. Pro-
duction figures for the ferroalloy mining and milling industry
since 1945 are summarized in Table II1-18.
As Table 111-18 shows, molybdenum mining cind milJJng constitute
the largest and most stable segment of Llie ferroalloy ore
mining and milling industry in the United States. The U.S.
produces over 85% of the world's molybdenum supply, with two
mines dominating the industry. These two mines cire among tho
25 largest mining operations In the U.S. Production is
expected to Increase in the near future wirli i-.xp.mded output.
from existing facilities, and at least mio major in-.w oper.i-
tion in Colorado is expected to be in operation soon.
The only commercially important ore of molybdenum Is
molybdenite, MoS^. It is mined by both open-pit and under-
ground methods and is universally concentrated by flotation.
Commercially exploited ore currently ranges from O.I- to 0. )
Ill-bJ
DRAFT
-------
DRAFT
TABLE HI-IS. PRODUCTION OF FERROALLOYS BY
U.S. MINING AND MILLING INDUSTRY
METAL
Chromium
CoJumbJum and
Tantalum
Cobalt
Manganese
Molybdenum
Nickel
Tungsten
(60% WO3)
Vanadium*
ANNUAL PRODUCTION IN METRIC TONS (SHORT TONS)
1949*
394
(433)
0.5
(0.5)
237
(261)
103.835
(114,427)
10.222
(11,265)
0
1.314
(1,448)
N.A.
1953*
53.470
(58.817)
6.8
(7.4)
572
(629)
129,686
(142.914)
25.973
(28.622)
0
4,207
(4.636)
N.A.
1958t
—
194.7tt
(214.2)
2,202
(2,422)
—
18.634
(20,535)
-
3.437
(3,788)
2,750
(3,030)
1962t
0
-
-
-
23.250
(25.622)
-
7,649
(8,429)
4,749
(5,233)
1968**
0
0
550
(605)
43,557
(48.000)
42,423
(46.750)
13.750
(15,1501
8.908
(9.817)
5,580
(6.149)
1972T
0
0
0
16,996
(18.730)
46,368
(51.098)
15.303
(16.864)
6.716
(7,401)
4.435
(4.887)
•Reference 6
* Reference 3
••Reference 7
" Reference 5
111-64
DRAFT
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percent molybdenum content (Reference 7). Significant
quantities of molybdenite concentrate are recovered as a
byproduct In the milling of copper and tungsten ores.
Tungsten ores are mined and milled at many locations in
the U.S., but most of the production is from one operation.
In 1971, for example, the Bureau of Mines reported 66 active
tungsten mines, but total annual production from 59 of them
was less than 1000 metric tons (1102 short tons) each and,
from five others, less than 10,000 metric tons (11,023 short
tons) (Reference 2) . These small mines and mills are operated
intermittently, so it Is quite difficult to locate and contact
active plants at any given time. Tungsten production has
been strongly influenced by government policies. During
stockpiling in 1955, 750 operations produced tungsten ore at
$63 per unit in 1970 (unit = 9.07 kg (20 Ib) of 70% W con-
centrate); with the sale of some stockpiled material, only
about 50 mines operated with a price of $43 per unit (Refer-
ence 7) . Projected demand for tungsten will exceed supply
before the year 2000 at present prices, and production from
currently inactive deposits may be anticipated (Reference 7).
Commercially important ores for tungsten are scheelite (
and the wolframite series, wolframite ((Fe, Mn)WOM , ferberite
(FeWOji) , and huebnerite (MnWO^. Underground mining predomin-
ates, and concentration is by a wide variety of techniques.
Gravity concentration, by jigging, tabling, or sink float
methods, is frequently employed. Because sliming due to the
high friability of scheelite ore (most U.S. ore is scheelite)
reduces recovery by gravity techniques, fatty-acid flotation
may be used to increase recovery. Leaching may also be
employed as a major beneficiation step and is frequently
practiced to lower the phosphorus content of concentrates.
Ore generally contains about 0.6 percent tungsten, and
concentrates containing about 70 percent W03^ are produced.
A tungsten concentrate Is also produced as a byproduct of
molybdenum milling at one operation in a process involving
gravity separation, flotation, and magnetic separation.
Managanese and nickel ores are each recovered at only one
active operation In the U.S. at this time. The manganese
operation is completely dry, having no mine-water discharge
and no mill. At the nickel mine, small amounts of
conveyor wash water and scrubber water from ore milling are
mixed with effluents from an on-site smelter and with seasonal
mine-site runoff. Water-quality impact from the mining and
milling of these two metals is thus presently minimal.
111-65
DRAFT
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DRAFT
Future production of manganese and nickel, however, may be
expected to involve considerable water use.
Manganese is essential to the modern steel industry, both
as an alloying agent and as a deoxldizer, and these uses
dominate the world manganese industry (Reference 8).
Additional uses include material for battery electrodes and
agents for impurity removal in glassmaking. Domestic pro-
duction of manganese ores and concentrates has generally
accounted for a very small fraction of U.S. consumption,
the majority being supplied from foreign concentrates
(Reference 7). A number of significant plants have, however,
been operated for manganese recovery using a variety of
processing methods, and known ore reserves exist which are
economically recoverable.
The U.S. Bureau of Mines divides manganese-bearing ores into
three classes (Reference 7):
(1) manganese ores (at least 35 percent manganese
content)
(2) ferruginous manganese ore (10 to 35 percent
manganese content)
(3) manganiferous iron ore (less than 10 percent
manganese content)
The latter two classes are often grouped as manganiferous
ores and, in recent years, have accounted for nearly all
domestic production. In 1971, for example, only 5 percent
of the total production of 43,536 metric tons (48,000 short
tons) was in the form of true manganese ores (Reference 7) .
Future domestic production is likely on a significant scale
from manganiferous ores — particularly, on the Cuyuna Range
in Minnesota, where preparations for the resumption of
production are currently underway. This area, although
currently quiescent, accounted for 85 percent of domestic
production in 1971 (Reference 7).
Manganese ores have been processed by a wide variety of
techniques, ranging from dry screening to ore leaching.
Notable concentrating procedures in the recent past have
included sink-float separation, fatty-acid flotation
(References 9, 10, 11, 12), and ammonium carbanwto leach-
ing (Reference 13). It is most likely that heavy-media
separation will be practiced in the immediate future.
111-66
DRAFT
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Nickel ores are not widely available in the U.S. The
lateritic deposit which is currently being mined is the
only known domestic deposit of its kind. Some sulfide
nickel ore deposits with commercial possibilities have
been found in Alaska (Reference 2). If they are devel-
oped, processes entirely different from those in use at
the present operation will be employed. Most likely,
processing will involve selective flotation with reagent
and water usage and pollution problems quite similar to
those of Canadian nickel operations (Reference 14) and
domestic copper sulfide deposits (Reference 5).
There are no mines or mills currently active in the U.S.
producing ores or concentrates of chromium, cobalt, colum-
blum, and tantalum. Further, no operations could be
identified where they are recovered as a significant
byproduct, although the metals and their compounds are re-
covered at a number of domestic smelters and refineries.
This production is primarily from foreign ores and concen-
trates but includes some recovery from domestic concentrates
of other metals.
Chromium ore production in the U.S. has occurred only
under the Impetus of government efforts to stimulate a
domestic Industry. Production of chromite ore from the
Stillwater Complex during World War II, and from 1953
through 1961, involved gravity concentration by tabling,
and this mode of operation is likely in the event of
future production. Leaching of foreign concentrates,
currently practiced might provide an alternative method
of concentrating chromium values in domestic ores.
Domestic production by any means is unlikely, however, for
the next several years. Production costs for chromium
from domestic ores are estimated to be $110 per metric
ton ($100 per short ton), and no shortage is expected
in the near future (Reference 5).
Cobalt has been recovered in significant quantities at two
locations in the U.S., neither of which is currently active.
One of these, In the Blackbird district at Cobalt, Idaho,
has some probability of further production in the near
future. At these sites, as at essentially all sites around
the world, cobalt is a coproduct or byproduct of other metals,
and the production rates and world price of these other
metals, particularly copper and nickel, exert primary influ-
ence on the cobalt market (Reference 5). Known domestic
ore from which cobalt might be recovered is a complex copper
111-67
DRAFT
-------
DRAFT
cobalt sulfide ore which is likely to be processed by selective
flotation and roasting and leaching of the cobalt-bearing
float product (Reference 5 ).
Columbium and tantalum concentrates have in the past been
produced at as many as six sites in the U.S. (Reference 15),
and several potentially workable deposits of the ore minerals
pyrochlore and euxenite are known. Economic recovery would
require a twofold increase in price for the metals, however,
and is considered unlikely before the year 2000 (Reference 5 ).
Production, should it occur, would involve placer mining
at one of the known deposits, with the water quality impact
and treatment problems peculiar to that activity. Concentra-
tion techniques varying widely from fairly simple gravity
and hand picking techniques through magnetic and electro-
static separation and flotation have been used in the past.
Accurate prediction of the process which would be used in
future domestic production is not feasible.
Vanadium. Eighty-six percent of vanadium oxide production
has recently been used in the preparation of ferrovanadium.
Although a fair share of U.S. vanadium production is derived
as a byproduct of the mining of uranium, there are other
sources of vanadium ores. The environmental considerations
at mine/mill operations not involving radioactive constituents
are fundamentally different from those that are important at
uranium operations, and it seems appropriate to consider the
former operation separately. Vanadium is considered as part
of this industry segment: (a) because of the similarity of
non-radioactive vanadium recovery operations to the processes
used for other ferroalloy metals and (b) because, in parti-
cular, hydrometallurgical processes like those used in vana-
dium recovery are becoming more popular in SIC 1061. These
arguments are also presented in the discussion of the SIC
1094 (uranium, vanadium, and radium mining and ore dressing)
categories. Other aspects of effluent from uranium/vanadium
byproduct operations under Nuclear Regulatory Commission
(formerly AEC) license are treated further under that heading.
Vanadium is chemically similar to columbium (niobium) and
tantalum, and ores of these met.ils may be bend'tciated In
the same type of process used for vanadium. There Is also
some similarity to tungsten, moJybdenum, and chromium.
Ferroalloy Ore Beneficlatlon Processes
Ore processing in the ferroalloys industry varies widely.
and even ores bearing the same ore mineral may be coneentrdied
111-68
DRAFT
-------
DRAFT
by widely differing techniques. There is thus no scheelite
recovery process or pyrolusite concentration technique per £>e_.
On the other hand, the same fundamental processes may be used to
concentrate ores of a variety of metals with differences only
in details of flow rate, reagent dosage etc., and some func-
tions (such as crushing and grinding ore) that are common to
nearly all ore concentration procedures. Fundamental ore
beneficlation processes which require water may be grouped
into three basic classes:
1. Purely physical separation (most commonly,
by gravity)
2. Flotation
3. Ore Leaching.
Prior to using any of these processes, ore must, in general,
be crushed and ground; in their Implementation, accessory
techniques such as cycloning, classification, and thickening
may be of great importance.
Physical Ore Processing Techniques. Purely physical ore
beneficiatlon relies on physical differences between the
ore and accessory mineralization to allow concentration of
values. No reagents are used, and pollutants are limited
to mill feed components soluble in relatively pure water,
as well as to wear products of milling machinery. Physical
ore properties often exploited Include gravity, magnetic
permeability, and conductivity. In addition, friability
(or its opposite) may be exploited to allow rejection of
gangue on the basis of particle size.
Gravity concentration is effected by a variety of techniques,
ranging from the very simple to the highly sophisticated,
including Jigging, sink floatation, Humphreys spirals, and
tabling. Jigging Is applicable to fairly coarse ore, ranging
in size from 1 mm to 13 mm (approximately 0.04 to 0.50 inch),
generally the product of secondary crushing (Reference 5).
Ore is fed as a slurry to the jig, where a plunger operating
at 150 to 250 cycles per minute provides agitation. The
relatively dense ore sinks to the screen, while the lighter
gangue is kept suspended by the agitation and is removed with
the overflow. Often, a bed of coarser ore or iron shot is
used in the jig to aid in separation. Sink-float methods
rely on the buoyancy forces in a dense fluid to float the
gangue away from denser ore minerals. It is also a coarse
111-69
DRAFT
-------
DRAFT
ore separation technique generally applicable to particles
which are 2 mm to 5 mm (approximately 0.08 to 0.2 inch in
diameter) (Reference 5 ). Most commonly, the separation
medium is a suspension of very fine particles of dense materials
(ferrosilicon in the heavy media separation, and galena in
the Huntington-Heberlein process). Light gangue overflows
the separation tank, while ore is withdrawn from the bottom.
Both are generally dewatered on screens and washed, the
separation medium being reclaimed and returned to the
circuit (Reference 16).
Shaking tables and spiral separators are useful for finer
particle sizes; generally, ore must be ground before applica-
tion of these techniques. A shaking table is generally fed
at one end and slopes towards the opposite corner. Water
flows over a series of riffles or ridges which trap the heavy
ore particles and direct them at right angles to the water
flow toward the side of the table. The table vibrates,
keeping the lighter particles of gangue in suspension, and
the particles follow the feed water across the riffles.
The separation is never peffect, and the concentrate grades
into gangue at the edge of the table through a mixed product
called middlings, which is generally collected separately
from concentrate and gangue and then retabled. Frequently,
several sequential stages of tabling are required to produce
a concentrate of the desired grade. Particle size, as well
as density, affects the behavior of particles on a shaking
table, and the table feed generally must be well classified
to ensure both high ore recovery and a good concentration
ratio. Humphreys spiral separators are useful for ore
ground to between 0.1 mm and 2 mm (approximately 0.004 to
0.08 inch) (Reference 5 ). They consist of a helical conduit
about a vertical axis which is fed at the top with flow down
the spiral by gravity. Heavy minerals concentrate at the
inner edge and may be drawn off at ports along the length
of the spiral; wash water may also be added there to improve
separation. The capacity of a single spiral is generally
0.45 to 2.27 metric tons/hour (0.5 to 2.5 short tons/hour)
(Reference 17 ).
Magnetic and electrostatic separation are frequently used
for the separation of concentrates of different metals from
complex ores — for example, the separation of cassiterlr.'
columbite, and monazite (Reference 5 ) or the separation
of casslterite and wolframite (Reference 18). Although
they are both most frequently Implemented as dry processes,
wet-belt magnetic separators are used. Since ore particle:-.
are charged to 20,000 to 40,000 volts for electrostatic
111-70
DRAFT
-------
DRAFT
separation, no wet process exists. In magnetic separation,
particles of high magnetic permeability are lifted and held
to a moving belt by a strong magnetic field, while low per-
meability particles proceed with the original stream (wet-
belt separator) or belt (crossed-belt separator). In
electrostatic separation, charged nonconductive particles
adhere to a rotating conductive drum, while conductive part-
icles discharge rapidly and fall or are thrown off.
These processes may be combined with each other, and with
various grinding mills, classifiers, thickeners, cyclones,
etc., in an almost endless variety of mill flow sheets, each
particularly suited to the ore for which it has been developed.
These flow sheets may become quite complex, involving multiple
recirculating loops and a variety of processes as the examples
from the columbium and tantalum Industry shown in Figures
111-13 and 111-14 illustrate. It is believed that domestic
mills currently employing only physical separation will have
fairly simple flow sheets since they are all small processors.
Such an operation might be represented by the flow sheet
of Figure 111-15.
Water use in physical beneficiation plants may vary widely
from zero to three or more times the ore milled by weight.
However, there are no technical obstacles inherent in the
process to total reuse of water (except for the 20 to 30
percent by weight retained by tails) by recycle within the
process or from the tailings pond.
Flotation Processes. Flotation concentration has become
a mainstay of the ore milling industry. Because it is
adaptable to very fine particle sizes (less than 0.01 mm,
or 0.0004 inch), it allows high rates of recovery from
slimes which are inevitably generated in crushing and grind-
ing and are not generally amenable to physical processing.
As a physico-chemical surface phenomenon, it can often be
made highly specific, allowing production of high-grade
concentrates from very-low-grade ore (e.g., 95+ percent
MoS^. concentrate from 0.3 percent) (Reference 18 ).
Its specificity also allows separation of different ore
minerals (e.g., CuS and MoSjO where desired, and operation
with minimum reagent consumption since reagent interaction
is typically only with the particular materials to be
floated or depressed.
111-71
DRAFT
-------
DRAFT
Figure 111-13. GRAVITY-PLANT FLOWSHEET FOR NIGERIAN COLUMBITE
TO
WASTE
*- 111 Mil I
SOURCE Rtl!tl4CNCir 19
111-72
DRAFT
-------
DRAFT
Figure 111-14. EUXENITE/COLUMBITE BENEFICIATION-PLANT FLOWSHEET
TO
| DREDGE""]
HE AW MINERAL
CONCENTRATE
t
[ STORAGE |
i
1
__l
[ SCREEN | >| ROD
M.LL r*j MAGNETITE }-+. '»,„,.,
r
ro
VASTE
TO WASTE -«— SLIME -4 JCLASSIFIER [^ATTRITIONER^ SEPARATOR f*jCLASSIFIER|-^{ ORVER
RAGE-"- °OARTZ
RAGE"*- OARN"
t
INDUCED-ROLL
MAGNETIC SEPARATOI
L
1
'
1 MAGNETIC SEPARATOR
^~J (LOW INTENSITY)
'
,- .— J SCR
!_
1
MJ
>
*
*
INDUCED-ROLL
\GNETIC SEPARATOR
\ f URNACE |
i NONCONDI
H ELECTROSTATIC
SEPARATOR
NONCONDUCTORS
— . INDUCED-ROLL
v MAGNETIC SEPARATOR
T
4 CROSSBELT MAGNETIC
SEPARATOR
ELECTROSTATIC 1
SEPARATOR |
ICTORS— 1
1 1— MIDDLINGS — — »•
NONMAGNETICS
SCREEN "1
1— CONDUCTORS •,
f SCREEN |
< 28 MESH 1
}
1 CROSSBELT MAGNETIC
SEPARATOR
CROSSBELT MAGNETIC
SEPARATOR
1
' 1 ILMENITE |—
> 36 MESH < 36 MESH 1 NONMAGNETICS
tX NONMAGNETICS J.
f ALTERNATE 1 f
'
| AIR TABLE |
TE -^-TAILINGS -4-| WET TABLE | DRIED CC
CONCENTRATE
INCENTRATE » 1
ICHOSSBELI MAGNETIC 1
SEPARATOR [-
, 1
| MONAZITE ] |
FUXFNITE ] C
NONMAUNFTICS
AND MIDDLINGS •"
~l
9LUMBIII |
1
»
-^.'
TO
STORAGE
STORAGE
TO STORAGE
111-73
DRAFT
-------
DRAFT
Figure 111-15. REPRESENTATIVE FLOW SHEET FOR SIMPLE GRAVITY MILL
TO
WASTE
TAILS-
MINING
ORE
i
GRINDING AND
CRUSHING
I
SCREENING
FINE
SHAKING
TABLE
MIDDLINGS-
COARSE
- HEADS
•TAILS-
MIDDLINGS
SHAKING
TABLES
J
•CONCENTRATE
111-74
DRAFT
-------
DRAFT
Details of the flotation process — exact suite and dosage
of reagents, fineness of grinding, number of regrinds,
cleaner-flotation steps etc., — will differ at each opera-
tion where practiced; and may often vary with time at a given
mill. The complex system of reagents generally used includes
four basic types of compounds: collectors, frothers, activa-
tors, and depressants. Frequently, activators are used to
allow flotation of ore depressed at an earlier stage of the
milling process. In almost all cases, use of each reagent
in the mill is low—generally, less than 0.5 kg per metric
ton of ore (1.00 Ib per short ton)—and the bulk of the
reagent adheres to tailings or concentrates. Reagents
commonly used and observed dosage rates are shown in Table
111-19.
Sulfide minerals are all readily recovered by flotation
using similar reagents in small doses, although reagent
requirements and ease of flotation do vary through the
class. Flotation is generally carried out at an alkaline
pH, typically 8.5 for molybdenite (Reference 18). Collect-
ors are most often alkali xanthates with two to five carbon
atoms — for example, sodium ethyl xanthate (C2H5G . NaCSjZ).
Frothers are generally organics with a soluble hydroxyl group
and a "non-wettable" hydrocarbon (Reference 17 ). Pine oil
(C6H120H), for example, is widely used. Depressants vary
but are widely used to allow separate recovery of metal values
from mixed sulfide ores. Sodium cyanide is widely used as
a pyrite depressant — particularly, in molybdenite recovery.
Activators useful in sulfide ore flotation may include cuprous
sulfide and sodium sulfide.
The major operating plants in the ferroalloy industry recover
molybdenite by flotation. Vapor oil is used as the collector,
and pine oil is used as a frother. Lime is used to control
pH of the mill feed and to maintain an alkaline circuit.
In addition, Nokes reagent and sodium cyanide are used to
prevent flotation of galena and pyrite with the molybdenite.
A generalized, simplified flowsheet for an operation
recovering only molybdenite is shown in Figure I11-16.
Water use in this operation currently amounts to approxi-
mately 1.8 tons of water per ton of ore processed, essentially
all of which is process water. Reclaimed water from thick-
eners at the mill site (shown on the flowsheet) amounts to
only 10 percent of total use.
Where byproducts are recovered with molybdenite, a somcwh.it
more complex mill flowsheet results, although the
111-75
DRAFT
-------
DRAFT
TABLE 111-19. OBSERVED USAGE OF SOME FLOTATION REAGENTS
REAGENT
OBSERVED USAGE
IN KILOGRAMS
PER METRIC TON
IN POUNDS
PER SHORT TON
SULFIDE FLOTATION
Vapor oil
Pine oil
NokM reagent
MIBC (methylisobutyl carbinol)
Sodium cyanide
Sodium silicate
Starch
Butyl alcohol
Creosote
Miscellaneous xanthates
Commercial frothers
0.1 to 0.4
0.02 to 0.2
0.04
0.02
0.005 to 0.02
0.25 to 0.35
0.0005
0.08
0.45
0.0005 to 0.2
0.002 to 0.2
0.2 to 0.8
0.04 to 0.4
0.08
0.04
0.01 to 0.04
0.50 to 0.70
0.001
0.16
0.90
0.001 to 0.4
0.004 to 0.4
OTHER FLOTATION
Copper sulfate
Sodium silicate
Oleic acid
Sodium oleate
Acid dichromate
Sodium carbonate
Fuel oil
Soap skimmings
Sulfur dioxide
Long-chain aliphatic aminei
Alkylaryl sulfonate
Misc. Tradenamed Products
0.4
0.3 to 3
0.06 to 6.5
0.05 to 0.2
0.1 to 0.4
4 to 6
60 to 95*
20 to SO*
6*
—
—
0.02 to 0.4
0.8
0.6 to 6
0.12 to 13
0.1 to 0.4
0.2 to 0.8
8 to 12
120 to 190*
40 to 100*
12*
0.04 to 0.8
•IN USE AT ONLY ONE KNOWN OPERATION. NOT NOW ACTIVE
111-76
DRAFT
-------
DRAFT
Figure 111-16. SIMPLIFIED MOLYBDENUM MILL FLOWSHEET
MINING
ORE
*
CRUSHINGS.
WEIGHING. AND
SCREENING
i
BALL
MILLS
\
1
CYCLONES 1 UNDERFLOW 1
OVERFLOW
ROUGHER FLOAT
MIDDLINGS
I
-CONCENTRATE
MIDDLINGS —
SCAVENGER
FLOAT
(4 STAGES WITH
REGRIND AND
INTERNAL RECYCLE!
—CONCENTRATE -
MIDDLINGS •
»«-UNDERFLOW
-UNDERFLOW-I THICKENER
1
CLEANER
FLOAT
16 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
T
CONCENTRATE
— TAILS -
UNDERFLOW-
ORYER
MOI YBDtNUM
PHODUri
TO TAILINGS
FOND
111-77
DRAFT
-------
DRAFT
molybdenite recovery circuits themselves remain quite
similar. A very simplified flow diagram for such an opera-
tion is shown in Figure 111-17. Pyrite flotation and
monazite floation are accomplished at acid pH (4.5 and 1.5,
respectively), somewhat increasing the likelihood of solu-
bllizing heavy metals. Volumes at those points in the circuit
are low, however, and neutralization occurs upon combination
with the main mill water flows for delivery to the tailing
ponds. Water flow for this operation amounts to approximately
2.3 tons per ton of ore processed, nearly all of which is
process water in contact with ore. Essentially 100 percent
recycle of mill water from the tailing ponds at this mill
is prompted by limited water availability as well as by
environmental considerations and demonstrates its technical
and economic feasibility, even with the complications
induced by multiple flotation circuits for byproduct recovery.
Other sulfide ores in the ferroalloy cateogry which may be
recovered by flotation are those of cobalt and nickel, although
no examples of these practices are currently active in the
U.S. It is to be expected that they will be recovered as
coproducts or byproducts of other metals by selective flotat-
tion from complex ores in processes involving multiple flota-
tion steps. Some of the most likely reagents to be used in
these operations are presented in Table 111-20, although
the process cannot be accurately predicted at this point.
It is expected that, as is generally the case, in sulfide
flotation, a small total amount of reagents will be
used.
Many minerals in addition to sulfides may be and often are
recovered by flotation. Among the ferroalloys, manganese,
tungsten, columbium, and tantalum minerals are or have been
recovered by flotation. Flotation of these ores involves
a very different suite of reagents from sulfide flotation
and, in some cases, has required substantially larger reagent
dosages. Experience has indicated these flotation pro-
cesses to be, in general, somewhat more sensitive to feed-
water conditions than sulfide floats; consequently, they are
less frequently run with recycled water.
In current U.S. operations, scheelite is rccovurt-d hy
flotation using fatty acids as collectors. A Lyplral suite
of reagents includes sodium silicate (1.0 kg/metric ton or
2.0 Ib/short ton) oleic acid (0.5 kg/metric ton, or 1.0 Ib/sliorc
ton), and sodium oleate (0.1 to 0.2 kg/metric ton, or 0.2 to
0.4 Ib/short ton). In addition, materials such as copper
sulfate or acid dichromate may be used in small to moderate
111-78
DRAFT
-------
DRAFT
Figure 111-17. SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM
CRUSHING
(3 STAGES)
28% + 3 MESH
GRINDING
BALL MILLS
36% + 100 MESH
FLOTATION
•CONCENTRATE
1
FLOTATION
96% OF MILL FEED
•LIGHT TO TAILS-
GRAVITY
HUMPHREY'S SPIRALS
i
PYRITE
FLOTATION
TAILS
t
LIGHT TO TAILS
MONAZITE
CONCENTRATE
TO TAILS
-i
TABLES
MONAZITE
FLOTATION
MAGNETIC
SEPARATION
NONMAGNETIC
TIN CONCENTRATE
I
CONCENTRATE
CLEANER
FLOTATION
(4 STAGES)
I
-TAILINGS
DRYING
t
MOLYBDENUM
CONCENTRATE
(93% + MoS2)
MAGNETIC TUNGSTEN
CONCENTRATE
111-79
DRAFT
-------
DRAFT
TABLE 111-20. PROBABLE REAGENTS USED IN FLOTATION OF
NICKEL AND COBALT ORES
Lime
Amyl Xanthate
Isopropyl Xanthate
Pine Oil
Methyl liobutyl Carbinol
Triathoxybutane
Dextrin
Sodium Cyanide
Copper Sulfate
Sodium Silicate
111-80
DRAFT
-------
DRAFT
amounts as conditioners and gangue depressants. Scheelite
flotation circuits may run alkaline or acid, depending
primarily on the accessory mineralization in the ore.
Flotation of sulfides which occurs with the scheelite is also
common practice. Sulfide float products may be recovered
for sale or simply removed as undesirable contaminants for
delivery to tails. Frequently, only a portion of the ore
(generally, the slimes) is processed by flotation, the
coarser material being concentrated by gravity techniques
such as tabling. A simplified flow diagram for a small
tungsten concentrator Illustrating these features is shown
in Figure 111-18. Note that, in this operation, an acid
leach is also performed on a part of the tungsten concentrate.
This is common practice in the tungsten industry as a means
of reducing phosphorus content in the concentrates. Approxi-
mately four tons of water are used per ton of ore processed
in this operation.
The basic flotation operations for manganese ores and colum-
blum and tantalum ores are not much different from scheelite
flotation; in general, they differ in specific reagents used
and, sometimes, in reagent dosage. One past process for a
manganese ore, however, bears special mention because of its
unusually high reagent usage — which could, obviously, have
a strong effect on effluent character and treatment.
Reagents used include:
Diesel oil 80 kg/metric ton (160 Ib/short ton)
Soap skimmings 40 kg/metric ton (80 Ib/short ton)
Oronlte S (wetting agent) 5 kg/metric ton (10 Ib/short ton)
S02^ 5 kg/metric ton (10 Ib/short ton)
With the exception of reagent consumption, the plant flow
sheet is typical of a straight flotation operation (like
that shown in Figure 111-16), involving multiple cleaning
floats with recycle of tailings.
While the flotation processes are similar, columbium and
tantalum flotation plants are likely to possess an unusual
degree of complexity due to the complex nature of their ores,
which necessitates multiple processes to effectively sepa-
rate the desired concentrates. This is illustrated in the
flowsheet for a Canadian pyrochlore (NaCaCb£06_F) mill in
Figure 111-19.
111-81
DRAFT
-------
DRAFT
Figure 111-18. SIMPLIFIED FLOW DIAGRAM FOR SMALL TUNGSTEN CONCENTRATOR
ORE
SULFIDE
FLOTATION
CYCLONE
75% SANDS
i
GRAVITY
TABLES
TAILINGS
25%
SLIMES
OVERFLOW
THICKENER
SCHEELITE
FLOTATION
HCI LEACH
(15 TO 20% OF
FRACTION)
TUNGSTEN
CONCENTRA1L
111-82
DRAFT
-------
DRAFT
Figure 111-19. MILL FLOWSHEET FOR A CANADIAN COLUMBIUM OPERATION
I MINING I
ORE
PRIMARY
CRUSHING
> lA-Cffl
» i*-em r
r~<0.7B.|nJ~l
H SECONDARY I
CRUSHING |
3
SCREENING
J SECONDARY | < ' •—«» 1*-M
CRUSHING
MILL
ORE BIN
ROD MILLING
I SCREENING 1—
40 MESH
-I BALL MILLING I
PVRITE
FLOTATION
CYCLONE
OVERFLOW
• CONCENTRATE
~L
MAGNETIC
SEPARATION
TAILS
DESLIMING
J
UNDERFLOW
r
I BULK FLOTATION L
CONOENTRATE
DESLIMING
J
UNDERFLOW
*
MAGNETIC
SEPARATION
I
PRIMARY CLEANING
(FIRST STAGE)
TAILS CONCENTRATE
PRIMARY CLEANING I
(SECOND STAGE! I
CONCENTRATE
^
SECONDARY CLEANING
ITHREE STAGES)
— TAILS-
CONCENTRATE
t
_»J MAGNETITE
1
TABLING
1 - TAILS -»J
REVERSE
FLOTATION
>
TO
STOCKPILE
CONCENTRATE-*-
TAILS
_i_
TABLING
CONCENTRATE
- TAILS-
TO
' TAILING
POND
CONCENTRATE
SOURCE: REFERENCES
111-83
DRAFT
-------
DRAFT
Ore Leaching Processes. While not a predominant practice
in the ferroalloys industry, ore leaching has played a. part
in a number of operations and is likely to increase as seg-
ments of the industry process ores of lower grade or which
are less easily beneficiated. A number of leaching processes
have been developed for manganese ores in the search for
methods of exploiting plentiful, low-grade, difficult-to-
concentrate domestic ore (that from most of the state of
Maine, for example) (Reference 6 ), and one such process
has been commercially employed. As mentioned previously,
leaching of concentrates for phosphorus removal is common
practice in the tungsten industry, and the largest domestic
tungsten producer leaches scheelite concentrates with soda
ash and steam to produce a refined ammonium paratungstate
product. Leaching is also practiced on chromice concentrates
(although not as a part of the domestic mining and milling
industry). Vanadium production by leaching nonradioactive
ores will also be considered here, because of vanadium's
use as a ferroalloy, and because it provides a well-
documented example of ferroalloy beneficiation processes
not well-represented in current practice, but likely to assume
importance in the future.
Leaching processes for the various ores clearly differ
significantly in many details, but all have in common (1)
the deliberate solubilization of significant ore components
and (2) the use of large amounts of reagents (compared to
flotation, for example). These processes share pollution
problems not generally encountered elsewhere, such as ex-
tremely high levels of dissolved solids and the possibility
of establishing density gradients in receiving waters and
destroying benthic communities despite apparently adequate
dilution.
The processes for the recovery of vanadium In the presence
of uranium are discussed in the subsection on uranium.
Recovery from phosphate rocks in Idaho, Montana, Wyoming,
and Utah ~ which contain about 28% P205, 0.25% V2_05^ and
some Cr, Ni, and Mo — yields vanadium as a byproduct of
phosphate rock, silica, coke, and iron ore (if not enough
iron is present in the ore). The product separated from
the slag typically contains 60 percent iron, 25 percent
phosphorus, 3 to 5 percent chromium, and 1 percent nickel.
It is pulverized, mixed with soda ash (Na2_C03_) and salt,
and roasted at 750 to 800 degrees Celsius (1382 to 1472
degrees Fahrenheit). Phosphorus, vanadium, and chromium
are converted to water-soluble trisodium phosphate, sodium
111-84
DRAFT
-------
DRAFT
metavanadate, and sodium chromate, while the iron remains in
insoluble form and is not extracted in a water leach following
the roast.
Phosphate values are removed from the leach in three stages
of crystallization (Figure 111-20). Vanadium can be recovered
as V2_0|5_ (redcake) by acidification, and chromium is precipi-
tated as lead chromate. By this process, 85 percent of vanadium,
65 percent of chromium, and 91 percent phosphorus can be ex-
tracted.
Another, basically non-radioactive, vanadium ore, with a grade
of 1 percent VjZO^, is found in a vanldlferous, mixed-layer
montmorillonlte/illite and goethite/montroseite matrix.
This ore is opened up by salt roasting, following extrusion
of pellets, to yield sodium metavanadate, which is concen-
trated by solvent extraction. Slightly soluble ammonium
vanadate is precipitated from the stripping solution and
calcined to yield vanadium pentoxide. A flow chart for
this process is shown In Figure 111-21.
The Dean Leute ammonium carhamate process has been used
commercially for the recovery of high-purity manganese carbonate
from low-grade ore on the Cuyuna Range in Minnesota and could
be employed again (Reference 13 ). A flow sheet is shown in
Figure 111-22.
Mercury Ores
The mercury mining and milling industry is defined for this
document as that segment of Industry engaged in the mining
and/or milling of ore for the primary or byproduct/coproduct
recovery of mercury. The principal mineral source of mercury
is cinnabar (HgS). The domestic industry has been centered
in California, Nevada, and Oregon. Mercury has also been
recovered from ore in Arizona, Alaska, Idaho, Texas, and
Washington and is recovered as a byproduct from gold ore in
Nevada and zinc ore in New York.
Due to low prices and slackened demand, the mercury industry
has been in a decline during recent years (Table 111-21).
During this time, the environmental hazards and extremely
toxic nature of mercury have come under public scrutiny.
One result has been the cancellation in March 1972 of all
biocidal uses of mercury under the terms of the Federal
Insectlde, Fungicide, and Rodentlclde Act. In addition,
registration has been suspended for mercury alkyl compounds
111-85
DRAFT
-------
DRAFT
Figure 111-20. FLOWSHEET OF TR1STAGE CRYSTALLIZATION PROCESS FOR
RECOVERY OF VANADIUM, PHOSPHORUS, AND CHROMIUM
FROM WESTERN FERROPHOSPHORUS
I FERROPHOSPHORUS |
| NlCt |
ROASTING
I SETTLING AN
|
| PRIMARY CRYSTALLIZATION]
I
[ PRIMARY CENTRIFUGINC
PREGNANT CHVS
SOLUTION 1 M PO i
1 *~^ ^
| TERTIARY CRVSTALLIZATION ]
| TERTIARY CENTRIFUGING
L
•« %£??. . 1 C.O. 1 iOtUTIO
>
DOECANTATION |
1
SOLIDS WATEB WAJM
1 A
T T
| FILTRATION |
j |
RESIDUE WASH ,.
1 LIQUOR
, . ^
TALS [ WATER ]
•— ^^— DISSOLUTION *
•
[ SECONDARY CRYSTALLIZATION]
| SECONDARY CENTRIFUCING 1
] 1 1
T t
CRYSTALS SECONDARY
EGNANT I SOLUTION — *"
V 1
• TO WASTE
RED CAKE MOTHER LIQUOR
*
' FUSION •
i
I
FILTRATION
BLACK-CAKE
VANADIUM PRODUCT
CHflOMAlE
SOLUTION
TO
WASTE
| FILTRATION
J
cuanuinu 1
PRODUCT j
ZJ
1 » SOLUTION
TO WASTE
» 1U
"" STOl'KPIlE
LII-86
DRAFT
-------
DRAFT
Figure 111-21. ARKANSAS VANADIUM PROCESS FLOWSHEET
1.5 -2.0% V2O5
t
6-10%
NaCL
TERTIARY
AMINES
GRINDING
PELLETIZING
I
ROASTING
I
850°C (1562°F)
NaVO,
H2S04
LEACHING AND
ACIDIFICATION
I
pH 2.5 • 3.5
-------
DRAFT
Figure 111-22. FLOWSHEET OF DEAN-LEUTE AMMONIUM CARBAMATE PROCESS
RAW SIZED ORE - 1.9 cm (0.75 in.)
-CO,
REDUCING FURNACE
GRINDING (30 MESH)
WEAK Mn
SOLUTION
I
LEACHING
(TWO 11,356-j£(3000-GAL))
REACTION TANKS
IN SERIES
I
9.14-m (30-ft) CLARIFYING
THICKENER
7.6-m (25-ft) COUNTERCURRENT
WASHING THICKENERS
LIVE
STEAM
• NEW LEACH LIQUOR
-ill
LEACH LIQUOR
REGENERATION
TWO 11.356- X.
(3000-GAL)
PRECIPITATION
TANKS ON SERIES)
-NH,
MnCO,
SLURRY
STILL
I
NH4NH2CO2
CLARIFYING
THICKENER
MOTHER
' LIQUOR "
TAILINGS
70% SOLIDS
i
ROTARY DRYER
•NH4NH2CO2
I
PRODUCT
111-88
DRAFT
-------
DRAFT
TABLE 111-21. DOMESTIC MERCURY PRODUCTION STATISTICS
CATEGORY
No. of producing mines
Production in metric tons
(flasks)
Dollar value (thousands)
YEAR
1969
109
1,029
(29,640)
$14.969
1970
79
948
(27.296)
$11.130
1971
56
621
(17,883)
$ 5.229
1972
21
253
(7.286)
$1.590
1973
6
SOURCE: REFERENCE 2
111-89
DRAFT
-------
DRAFT
and nonalkyl uses on rice seed, in laundry products, and in
marine antifouling paint. An immediate effect of this has
been a substantial reduction in the demand for mercury for
paints and agricultural applications. However, future growth
in the consumption of mercury is anticipated for electrical
apparatus, Instruments, and dental supplies. From consider-
ation of these factors, it is anticipated that demand for
mercury in 1985 will remain at the 1972 level. Given such
variables as market prices and effects of emission standards
promulgated in April 1973, it has been predicted that pro-
duction of primary mercury will range from a high of 20,000
flasks (695 metric tons, or 765 short tons) to a low of 3,000
flasks (104 metric tons, or 115 short tons) by 1985.
Mercury ore is mined by both open-pit and underground methods.
In recent years, underground methods have accounted for about
two-thirds of the total mercury production. Ore grade has
varied greatly, ranging from 2.25 to 100 kg of mercury per
metric ton (5 to 200 pounds of mercury per short ton) .
The grade of ore currently mined averages 3.25 kg of mercury
per metric ton (6.5 pounds of mercury per short ton).
The typical practice of the industry has been to feed the
mined mercury ore directly into rotary kilns for recovery
of mercury by roasting. This is such an efficient method
that extensive beneficiation is precluded. (However, with
the depletion of high grade ores, concentration of low-grade
mercury ores is becoming more important). The ore may be
crushed — and, sometimes, screened — to provide a feed
suitable for furnacing. Gravity concentration is also done
in a few cases, but its use is limited since mercury minerals
crush more easily and more finely than gangue rock.
Flotation is the most efficient method of beneficiating
mercury ores when beneficiation is practiced. An advantage
of flotation, especially for low-grade material, is the high
ratio of concentration resulting. This permits proportionate
reductions in the size and costs of the subsequent mercury
extraction installation. Flotation of mercury ore has not
been used to date in the United States. However, an operat I-in
scheduled to begin in Nevada later in 1975 will concentrate
mercury ore by flotation. This concentrate will be furnaci-u,
and annual production of mercury from the operation is e.\p.
to reach 20,000 flasks (695 metric tons, or 765 short tons).
111-90
DRAFT
-------
UHAFT
Uranium, Radium, and Vanadium Ores
The mining and milling of uranium, vanadium, and radium
constitute one industry segment, because uranium and
vanadium are often found in the same ore and because radium,
resulting from the radioactive decay of uranium, has always
been obtained from uranium ores. In the past 20 years, the
demand for radium has vanished as radioactive isotopes
(e.g., Co 60, Pu 240) with tailored characteristics as sources
of radiation have become available. Radium is now treated
as a pollutant in the wastes. Uranium is mined primarily
for its use in generating energy and isotopes in nuclear
reactors. In the U.S., vanadium is primarily generated as a
byproduct of uranium mining for use as a ferroalloying metal
and, in the form of its oxide, as a catalyst. Vanadium used
as a ferroalloy metal has been discussed in the Ferroalloys
Section.
The ores of uranium, vanadium, and radium are found both in
the oxidized and reduced states. The uranium (IV) oxida-
tion state is easily oxidized and the resulting uranium (VI),
or uranyl, compounds are soluble in various bases and acids.
In arid regions of the western United States, the ores are
found in permeable formations (e.g., sandstones), while
uranium deposits in humid regions are normally associated
with more impervious rocks. Uranium is often found in associa-
tion with carbonaceous fossils, i.e., lignite and asphalts.
Ores with a grade in excess of a fraction of a percent uranium
are rare (80% of the industry operates with ores below 0.2%).
Because It would be uneconomical to transport low-grade
uranium ores very far, mines are closely associated with mills
that yield a concentrate containing about 90 percent uranium
oxide. This concentrate is shipped to plants that produce
compounds of natural and isotopically enriched uranium for
the nuclear industry. The processes of crushing and grinding,
conventionally associated with a mill, are intimately connected
with the hydrometallurgical processes that yield the concen-
trate, and both processes normally share a wastewater disposal
system. Mine water, when present, is often treated separately
and is sometimes used as a source of mill process water.
Mine water frequently contains a significant amount of
uranium values, and the process of cleaning up mine water
not only yields as much as one percent of the product of some
mines but is also quite profitable.
111-91
DRAFT
-------
DRAFT
The uranium oxide concentrate, whose grade is usually quoted
in percent of \J3Q^ (although that oxide figures in the assay,
rather than in the product), is generated by one of several
hydrometallurgical processes. For purposes of wastewater
categorization, they may be distinguished as follows:
(1) The ore is leached either in sulfuric acid, or
in a hot solution of sodium carbonate and sodium
bicarbonate, depending on the content of acid-
wasting limestone in the gangue.
(2) Values in the leachate are sometimes concentrated
by ion exchange, either in a solid resin (IX) or
by solvent extraction (SX). They are then precipi-
tated as the concentrate, yellowcake.
Some vanadium finds are not associated with significant uranium
concentrations. There, as in vanadium byproduct operations,
vanadium oxide (V_205) is made by water leaching following a
salt roasting operation. Some byproduct concentrate solutions
are sold to vanadium mills for purification, and not all
uranium mills separate vanadium, which appears to be in
adequate supply and could be recovered later from tailings.
Ores and Mining. Consideration of thermonuclear equilibria
suggests an initial abundance of uranium in the solar system
of 0.14 ppm (parts per million). Since uranium is radioactive,
its concentration decreases with time, and its present abundance
is estimated as 0.054 ppm. The four longest-lived isotopes
are found in the relative abundances shown in Table 111-22.
Primary deposits of uranium ore contain uraninite, the U(IV)
compound U02^ and are widely distributed in granites and
pegmatites. Pure speciments of this compound, with density
ranging to 11, are rare, but its fibrous form, pitchblende,
has been exploited in Saxony since the recognition of
uranium in 1789.
Secondary, tertiary, and higher-order deposits of uranium
ores are formed by transport of slightly water-soluble uranyL
(U(VI)) compounds, notably carbonates. Typically, a primary
deposit is weathered by oxidized water, forming hydrated
oxides of uranium with compositions intermediate between
V02_ and UOJK The composition U30£ — i.e., U0£.21)03^ — is
particularly stable. The process occasionally stops at gumniiU'
(U02^H2_0) , an orange or red, waxy mineral, but usually
involves further oxidation and reactions with alkaJine and
alkaline-earth oxides, silicates, and phosphates. The trnns-
111-92
DRAFT
-------
UHAM
TABLE 111-22. ISOTOPIC ABUNDANCE OF URANIUM
ISOTOPE
U238
U235
U234
U236
HALF-LIFE (YEARS)
451 x 109
7.13 X108
2.48 x 105
2.39 x 107
ABUNDANCE
9927%
0.72%
0.0057%
Traces Identified
(Moon-1972; Earth-1974)
111-93
DRAFT
-------
DRAFT
port leads to the surface uranium ores of arid lands, including
carnotite (K2/U02)2/V04;2.3H2p), uranophane (CaU2Si20n^7H20) ,
and autunite (Ca(U02)2_(P04)£.10-12H2_0) and, if reducing
conditions are encountered, to the redeposition of U(IV)
compounds. Vanadium is seen to follow a similar route.
Radium, with a halflife of only 1600 years, is generated
from uranium deposits in historical times.
A reducing environment is often provided by decaying biologi-
cal materials; uranium is found in association with lignite,
asphalt, and dinosaur bones. One drift at a mine in New
Mexico passes lengthwise through the ribcage of a fossil
dinosaur. Since the requisite conditions are often encoun-
tered in the sediments of lakes or streams, stratiform
uranium deposits are common, constituting 95% of U.S.
reserves. Stratiform deposits comprise sandstone, conglom-
erate , and limestone with uranium values in pores or on the
surface of sand grains or as a replacement for fossilized
organic tissue. A small fraction of steeply sloping vein
deposits, similar to those in Saxony, is found in associa-
tion with other minerals. Some sedimentary deposits extend
over many kilometers with a slight dip with respect to
modern grade that makes it profitable to mine a given deposit
by open-pit methods at one point and by underground mining
at others.
Exploration is conducted initially with airborne and surface
radiation sensors that delineate promising regions and is
followed by exploratory drilling, on a 60-m (197-ft) grid,
and development drilling, on a 15-m (49-ft) grid. Test
holes are probed with scintillation counters, and cores are
chemically analyzed. Reserves have usually been specified
in terms of ore that can yield uranium at $18 per kg (2.2 Ib),
a price paid by the government for stockpiling.
Recent increases in costs and the possibility of increased
uranium demand due to the current energy situation have
resulted in the mining, for storage, of ore below this thres-
hold and may effect an increase in reserves. Currently,
reserves are concentrated in New Mexico and Wyoming, as
shown in the tabulation below.
DISTRIBUTION OF U.S. URANIUM ORE RESERVES
New Mexico 44%
Wyoming 39%
Utah 4%
Colorado 4%
Texas 4%
Others 5%
111-94
DRAFT
-------
L/ru+r i
The number of separate deposits in the western United States
exceeds 1000, but half of the reserves lie in 15 deposits.
Four of these, in central Wyoming, on the border between
Colorado and Utah, in northwestern New Mexico, and on the
Texas gulf coast, dominate the industry. In 1970, New Mexico
provided 46 percent, Wyoming 26 percent, Colorado 12 percent,
and Utah 7 percent of uranium production, for a total of 91
percent of U.S. production.
In the eastern United States, uranium is found in conjunction
with phosphate recovery in Florida, in states throughout
the Appalachian Mountains, and in Vermont and New Hampshire
granites. The grade of these deposits is currently too low
for economic recovery of uranium, which is recovered as a
byproduct only in Florida. Vanadium, in ores that do not
contain uranium values, is mined in Arkansas and Idaho. The
humid environment of current and prospective eastern deposits
presents special problems of water management. Ocean water
contains 0.002 ppm of uranium, and its recovery with a process
akin to ion exchange using titanium compounds as a "resin"
has been explored in the United Kingdom. Uranium can be
recovered in this fashion at a cost of $150 to $300 per kg
(2.2 Ib).
Mining practice is conventional. There are 122 underground
mines as of 1 January 1974, with a typical depth of 200 m
(656 ft). Special precautions for the ventilation of under-
ground mines reduce the exposure of miners to radon, a short-
lived, gaseous decay product of radium that could leave
deposits of its daughters in miners lungs, Mine water is
occasionally recycled through the mine to recover values by
leaching and ion exchange.
Because of the small size of pockets of high-grade ore, open-
pit mines are characterized by extensive development activity.
At present, low-grade ore is stockpiled for future use.
Stockpiles on polyethylene sheets are heap leached at several
locations by percolation of dilute H2S04_ through the ore stock-
piles. On January 1974, 33 open pit mines were being worked,
and 20 other (e.g., heap-leaching) sources were in operation.
Most mines ship ore to the mill by truck. In at least one
Instance, a short (100-km, or 62-mi.) railroad run is involved.
Most mining areas share at least two mill processes, one
using acid leaching and the other, for high limestone content,
using alkaline leaching.
111-95
DRAFT
-------
DRAFT
Milling. Mills range in ore processing capacity from 450
metric tons (495 short tons) per day to 6500 metric tons
(7,150 short tons) per day, and 15 to 25 mills have been
in operation at any one time during the last 15 years.
Mill activities, listed by state, are given in Table III-
23 and are tabulated by company in Supplement B.
Blending. Crushing, and Roasting. Ore from the mine tends
to be quite variable in consistency and grade and may come
from mines owned by different companies. Fairly complex
procedures have been developed for weighing and radiometric
assay of ores, to give credit for value to the proper source
and to achieve uniform grade, and for blending to assure
uniform consistency. Sometimes, coarse material is separated
from fines before being fed to crushers that reduce it to
the 5 to 20 mm (0.2 to 0.8 in.) range. This material is
added to the fines.
Ore high in vanadium is sometimes roasted with sodium chloride
at this stage Co convert insoluble heavy-metal vanadates (vanadium
complex) and carnotlte to more soluble sodium vanadate, which is
then extracted with water. More often, this process is deferred
until after concentration of uranium/vanadium values. Ores high
in organics may be roasted to carbonize and oxidize these and
prevent clogging of hydrometallurgical processes. Clayey ores
attain improved filtering and settling characteristics by
roasting at 300 degrees Celsius (572 degrees Fahrenheit).
Grinding. Ore is ground to less than 0.6 mm (28 mesh) (0.024
in.) for acid leaching and to less than 70 micrometers (200
mesh) for alkaline leaching in rod or ball mills with water
(or, preferably, leach) added to obtain a pulp density of
about two-thirds solids. Screw classifiers, thickeners, or
cyclones are sometimes used to control size or pulp density.
Acid Leach. Ores with a calcium carbonate (CaC03_) content
of less than 12 percent are preferentially leached in sul-
furic acid, which extracts values quickly (in four hours to
a day) , and at a lower capital and energy cost than alkaline
leach for grinding, heating, and pressurizing. Any tetravalent
uranium must be oxidized to the uranyl form by the addition of an
oxidizing agent (typically, sodium chlorate or manganese dioxide),
which is believed to facilitate the oxidation of U(IV) to 'J(VI) in
conjunction with the reduction of Fe (III) to Fe (II)
at a redox (reduction/oxidation) potential of about
minus 450 mV. Free-acid concentration is held to between
111-96
DRAFT
-------
DRAFT
TABLE 111-23. URANIUM MILLING ACTIVITY BY STATE, 1972
STATE
New Mexico
Wyoming
Colorado
Utah
Texas
South Dakota
Washington
TOTAL
TOTAL MILL HANDLING CAPACITY
METRIC TONS PER DAY
12,300
8.250
4.000
1.850
3,400
600
450
30.850
SHORT TONS PER DAY
13.600
9,100
4,400
2,000
3.750
660
500
34.010
NO. OF MILLS
3
7
3
2
3
1
1
20
111-97
DRAFT
-------
DRAFT
1 and 100 grams per liter. The larger concentrations are
suitable when vanadium Is to be extracted. The reactions
taking place In acid oxidation and leaching are:
^ 0 £ — > 2U03_
2U03. + 2H2SO4 + 5H20->2 (U02S04) . 7H20
Uranyl sulfate (U02S04) forms a complex, hydrouranyl tri-
sulfuric acid (H4U02/S04)3) , in the leach, and the anions
of this acid are extracted for value.
Alkaline Leach . A solution of sodium carbonate (40 to
50 g per liter) in an oxidizing environment selectively
leaches uranium and vandium values from their ores. The
values may be precipitated directly from the leach by rais-
ing the pH with the addition of sodium hydroxide. The super-
natant can be recycled by exposure to carbon dioxide. A
controlled amount of sodium bicarbonate (10 to 20 g per
liter) is added to the leach to lower pH during leaching to
a value that prevents spontaneous precipitation.
This leaching process is slower than acid leaching since
other ore components are not attacked and shield the uranium
values. Alkaline leach is, therefore, used at elevated
temperatures of 80 to 100 degrees Celsius (176 to 212 degrees
Fahrenheit) under the hydrostatic pressure at the bottom of
a 15 to 20 m (49.2 to 65.6 ft) tall tank, agitated by a cen-
tral airlift (Figure 111-23). In some mills,
the leach tanks are pressurized with oxygen to increase the
rate of reaction, which takes on the order of one to three
days. The alkaline leach process is characterized by the
following reactions:
(oxidation)
3Na2_(C03J + UO^ + H20 — > 2NaOH
(leaching)
2NaOH + C0£ -> Na2C02 + H20
(recarbonization)
2Na(U02)(C03)2 + 6NaOH ->Na2y207^ + 6Na2C03_ + 3H2
(precipitation)
111-98
DRAFT
-------
DRAFT
Figure 111-23. PACHUCA TANK FOR ALKALINE LEACHING
LEACH
AIRLIFT
111-99
DRAFT
-------
DRAFT
The efficient utilization of water in the alkaline leach
circuit has led to the trend of recommending its expanded
application in the uranium industry. Alkaline leaching
can be applied to a greater variety of ores than in current
practice; however, the process, because of its slowness,
appears to involve greater capital expenditures per unit
production. In addition, the purification of yellow cake,
generated in a loop using sodium as the alkali element,
consumes an increment of chemicals that tend to appear in
stored or discharged wastewater but are often ignored.
Purification to remove sodium ion is necessary both to meet
.the specifications of American uranium processors and for
the preparation of natural uranium dioxide fuel. The latter
process will be used to Illustrate the problem caused by
excess sodium. Sodium diuranate may be considered as a mix-
ture of sodium and uranyl oxides — i.e., Na2.U2.07, --- Na2_0 +
The process of generating U02_ fuel pellets from yellow-cake
feed involves reduction by gaseous ammonia at a temperature
of a few hundred degrees C. At this temperature, ammonia
thermally decomposes into hydrogen, which reduces the U03^
component to U02^ and nitrogen (which acts as an inert gas
and reduces the risk of explosion in and around the reducing
furnace). With sodium diuranate as a feed, the process
results in a mix of U0£ and Na20 that is difficult to purify
(by water leaching of NaOH) without impairing the ceramic
qualities of uranium dioxide. When, in contrast, ammonium
diuranate is used as feed, all byproducts are gaseous, and
pure U02_ remains. The structural integrity of this ceramic
is immediately adequate for extended use in the popular
CANDU (Canadian deuterium-uranium) reactors. Sodium ion,
as well as vanadium values, can be removed from raw yellow
cake (sodium diuranate) produced by alkaline leaching in
two steps. In the first step, the yellow cake is roasted,
and some of the sodium ion forms water-soluble sodium vanadate,
while organics are carbonized and burned off. The roasted
product is water leached, yielding a V2_0_5. concentrate as des-
cribed below. The remaining sodium diuranate is redissolved
in sulfuric acid,
Na2jU2_0;7_ + 3H2_SCM ->Na2S(M + 3H20 + 2(U02_)S04_
and the uranium values are precipitated with ammonia and
filtered, to yield a yellow cake (ammonium diuranate or UOJ)
that is low in or free of sodium.
III-100
DRAFT
-------
DRAFT
U02SCM + H20 + 2mi_ - > (NH4) 2804^ + U03_
The reactions leading to this product are interesting for
their byproduct — namely, sodium sulfate. The latter, being
classed approximately in the same pollutant category as
sodium chloride, requires expensive treatment for its removal,
Ammonium-ion discharges which might result from a hypothe-
tical ammonium carbonate leaching circuit that would yield
the desired product immediately are viewed with more concern,
even though there is a demand for ammonium sulfate to fer-
tilize alkaline southwestern soils. Ammonium sulfate could
be generated by neutralizing the wastes of the hypothetical
ammonium loop with sulfuric acid wastes from acid leaching
wastes. Opponents of a tested ammonium process argue that
nitrites, an intermediate oxidation product of accidentally
discharged ammonium ion, present a present health hazard
more severe than that from sulfate ion.
Vanadium Recovery. Vanadium, found in carnotite (K2(U02J
(V04)2^ . 3H20) as well as in heavy metal vanadates — e.g.,
vanadinite (9PbO . 3V205^ . PbC12)— is converted to sodium
orthovanadate (Na.3V04) , which is water-soluble by roasting
with sodium chloride or soda ash (Na2C03) . After water
leaching, ammonium chloride is added, and poorly soluble
ammonium vanadates are precipitated:
Na3V(M + NH4C1 + H20 — ^ 2NaOH + NaCl + NH4V03_
(ammonium metavanadate)
Na3VO^ + 3NH4C1 — > 3NaCl + (NH4)J3V(M
(ammonium orthovanadate)
The ammonium vanadates are thermally decomposed to yield
vanadium pentoxide:
2(NH4)3V04 — > 6NH3 + 3H20
A significant fraction (86 to 87%) of V205^ is used in
the ferroalloys industry. There, ferrovanadium has been
prepared in electric furnaces by the reaction:
V205^ + FeZp^ + 8C — > SCO + 2FeV
or by alumino thermic reduction (See Glossary) in the
presence of scrap iron.
The remainder of y2_05^ production is used in the inorganic
chemical industry, and its processing is not within the
scope of these guidelines.
111-101
DRAFT
-------
DRAFT
Since the mining and beneficiation of vanadium ores not con-
taining uranium values present an excellent example of
hydrometallurgical processes in the mining and ore dressing
of ferroalloy metals (under SIC 1061), it will be explored
further under that heading. Because of the chemical similarity
of vanadium to columbium, tantalum, and other ferroalloy
metals, recovery processes for vanadium are likely to be
quite similar to hydrometallurgical processes that will be
used in the ferroalloys mining industry when it becomes
more active again.
Concentration and Precipitation. To a rough approximation,
a metric ton of ore with a grade of about 0.2% is treated
with a metric ton (or cubic meter) of leach, and the concen-
tration (s) of uranium and/or vanadium in the pregnant solution
are also of the order of 0.2%. If values were directly pre-
cipitated from this solution, a significant fraction would
remain in solution. Yellow cake is, therefore, recycled and
dissolved in pregnant solution to increase precipitation
yield. Typically, five times as much yellow cake is recycled
as is present in the pregnant solution. Direct precipitation
by raising pH is effective only with alkaline leach, which is
somewhat selective for uranium and vanadium. If it were
applied to the acid leach process, most heavy metals —
particularly, iron — would be precipitated and would severly
contaminate the product.
Uranium (or vanadium and molybdenum) in the pregnant leach
liquor can be concentrated by a factor of more than five
through ion exchange or solvent extraction. Typical concentra-
tions in the eluate of some of these processes are shown in
Table 111-24.
Precipitation of uranium from the eluates is practical without
recycling yellow cake, and the selectivity of these processes
under regulated conditions (particulary, pH) improves the
purity of the product.
All concentration processes operate best in the absence of
suspended solids, and considerable effort is made to reduce
the solids content of pregnant leach liquors (Figure III-24,i) .
A somewhat arbitrary distinction is made between quickly
settling sands that are not tolerated in any concentration
process and slimes that can be accommodated to some extent
in the resin-in-pulp process (Figure HI-24b, c). Sands
are often repulped, by the addition of some wastewater streau
or another, to facilitate flow to the tailing pond as much
III-102
DRAFT
-------
DRAFT
TABLE 111-24. URANIUM CONCENTRATION IN IX/SX ELUATES
PROCESS
U3O8 CONCENTRATION (%)
Ion exchange
Resin-in-pulp
Fixed-bed IX:
Chloride elution
Nitrate elution
Moving-bed IX:
Nitrate elution
0.8 to 1.2
0.5 to 1.0
1.0 to 2.0
1.9
Solvent extraction
Alkyl phosphates, HCI eluent
Amex process
Dapex process
Split elution minewater treatment
30.0 to 60.0
3 to 4
5.0 to 6.5
12 to 1.6
IX/SX combination
Eluex process
3.0 to 7.5
III-103
DRAFT
-------
DRAFT
Figure 111-24. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 1 of 2)
FROM
LEACH
WATER
¥»
PREGNANT
LEACH LIQUOR
»*«»o
REPULPING
CLEAR LEACH LIQUOR
TO COLUMN IX OR SX
a) LIQUID/SOLID SEPARATION
SLIMY PULP TO
RESIN-IN PULP IX
SAND
<>» Q~
TAILINGS
SLIMY,
PREGNANT-
PULP
\/
RESIN IN OSCILLATING BASKET
b) RESIN-IN-PULP PROCESS: LOADING
BARREN
PULP
TO TAILINGS
BARREN
ELUANT
PREGNANT
ELUATE TO
PRECIPITATION
c) RESIN-IN-PULP PROCESS: ELUTING
III-104
DflAFT
-------
DRAFT
Figure 111-24. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 2 of 2)
BARREN ELUANT
ELUTED (OR
REGENERATED)
RESIN
LOADED
RESIN
PREGNANT ELUATE
TO PRECIPITATION
d) FIXED-BED COLUMN ION EXCHANGE/ELUTION
PREG
LEAC
LIQU
0
H
u
o Ao«
Op°'/
W Oo
' o°
BARREN STI
* ELUANT SO
I
LOADED
i—1 l— i ORGANIC |4J— ' !-i
SOLVENT °o o
- * • •
••M^^^^M^
DIPPED
LVENT
SOLVENT
~ ~ —
HV J^ ([BARREN ^y JPREGNANT! i
^~* '| LIQUOR ^-^ ELUATE 11'
PHASE PHASE
LOADING SEPARATION STRIPPING SEPARATION
e) SOLVENT EXTRACTION
LEACH
f) E
IX
f
SX
RECYCLE
BARREN t» " ELUANT
ELUANT ' ^i f
IX IX
. I* \
PREGNANT
.ELUATE
PARTIALLY STORAGE)
STRIPPED \ . LOADED
RESIN ^ _ ' RESIN
g) SPLIT ELUTION
f PRECIPITATION
LUEX PROCESS
III-105
DRAFT
-------
DRAFT
as a few kilometers away. Consequently, there is some latitude
for the selection of the wastewater sent to the tailing pond,
and mill operators can take advantage of this fact in selecting
environmentally sound waste-disposal procedures.
Ion exchange and solvent extraction (Figure III-24b-e) are
based on the same principle: Polar organic molecules tend to
exchange a mobile ion in their structure — typically,
C1-, N03_-, HS(M-, C03_— (anions), or H+ or Na+ (cations) —
for an ion with a greater charge or a smaller ionic radius.
For example, let R be the remainder of the polar molecule (in
the case of a solvent) or polymer (for a resin), and let X
be the mobile Ion. Then, the exchange reaction for the
uranyltrisulfate complex Is
4RX + (U02SCM)3) ---- * - - (141102504 + 4X-
This reaction proceeds from left to right in the loading process.
Typical resins adsorb about ten percent of their mass in
uranium and increase by about ten percent in density. In
a concentrated solution of the mobile ion — for example,
In N-hydrochloric acid — the reaction can be reversed and the
uranium values are eluted — in this example, as hydrouranyl
trisulfuric acid. In general, the affinity of cation exchange
resins for a metallic cation increases with increasing valence
(Cr-f-H- > Mg-H- > Na+) and, because of decreasing ionic radius,
with atomic number (92U > 42 Mo > 23V) . The separation of
hexavalent 92U cations by IX or SX should prove to be easier
than that of any other naturally occurring element.
Uranium, vanadium, and molybdenum — the latter being a common
ore constituent — almost always appear in aqueous solutions
as oxidized ions (uranyl, vanadyl, or molybdate radicals),
with uranium and vanadium additionally complexed with anlonic
radicals to form trisulfates or tricarbonates in the leach.
The complexes react anionically, and the affinity of exchange
resins and solvents is not simply related to fundamental
properties of the heavy metal (uranium, vanadium, or molybdenum),
as is the case in cationic exchange reactions. Secondary
properties, Including pH and redox potential, of the pregnant
solutions influence the adsorption of heavy metals. For
example, seven times more vanadium than uranium Is adsorbed
on one resin at pH 9; at pH 11, the ratio is reversed, wirli
33 times as much uranium as vanadium being captured. Thesi-
variations In affinity, multLple columns, and control of
leaching time with respect uo breakthrough (the time when
the interface between loaded and regenerated resin, l''ij;urc
III-24d, arrives at the end of the column) are used ti> makv
III-106
DRAFT
-------
DRAFT
an IX process specific for the desired product.
In the case of solvent exchange, the type of polar solvent
and its concentration in a typically nonpolar diluent (e.g.,
kerosene) effect separation of the desired product. The
ease with which the solvent is handled (Figure III-24e)
permits the construction of multistage co-current and counter-
current SX concentrators that are useful even when each
stage effects only partial separation of a value from an
interferent. Unfortunately, the solvents are easily polluted
by slimes, and complete liquid/solid separation is necessary.
IX and SX circuits can be combined to take advantage of both
the slime resistance of resln-ln-pulp ion exchange and the
separatory efficiency of solvent exchange (Eluex process).
The uranium values are precipitated with a base or hydrogen
peroxide. Ammonia is preferred by a plurality of mills
because It results in a superior product, as mentioned in
the discussion of alkaline leaching. Sodium hydroxide,
magnesium hydroxide, or partial neutralization with calcium
hydroxide, followed by magnesium hydroxide precipitation,
are also used. The product is rinsed with water that is
recycled into the process to preserve values, filtered,
dried and packed into 200-liter (55-gal) drums. The strength
of these drums limits their capacity to 450 kg (1000 Ib) of
yellow cake, which occupies 28% of the drum volume.
Thorium. Thorium is often combined with the rare earths,
with which it is found associated in monazite sands. It
is actually an actlnide (rather than lanthanide) and chemi-
cally, as well as by nuclear structure, is closely allied
to uranium. Although it finds some use in the chemical and
electronics industry, thorium is primarily of value as a
fertile material for the breeding of fissionable reactor
fuel. In this process, thorium 232, used in a "blanket"
around the core of a nuclear reactor, captures neutrons
to form thorium 233, which decays to uranium 233 by the
emission of two beta particles with halflives of '22 minutes
and 27 days. Uranium 233 is fissile and can be used as a
fuel. The cycle is very attractive since it may be operated
in thermal-neutron, as well as fast-neutron, reactors. A
pseudo-breeding reactor (burning uranium 235 or plutonium
239 in the core and producing uranium 233 in the blanket),
with net breeding gain (quantity of fissile material bred/
quantity burned) less than one is already in commercial
operation.
Thorium is about three times as abundant as uranium in rocks,
but rich deposits are rare. Typical monazite sand ores con-
tain from 1 to 10 percent thoria (Th02). American ores from
II1-107
DRAFT
-------
DRAFT
the North and South Carolinas, Florida, and Idaho contain
1.2 to 7 percent Th02^ with a typical value of 3.4 percent.
Monazite, a phosphate of cerium and lanthanum with some
thorium and some uranium and other rare earths, is found
in granites and other igneous rocks, where its concentra-
tion is not economically extractable. Erosion of such rocks
concentrates the monazite sands, which constitute about 0.1
percent of the host rock, in beach and stream deposits.
Mining often is combined with the recovery of ilmenite, rutile,
gold, zircon, casslterite, or other materials that concen-
trate in a similar way. Monazite is brittle, radioactive,
and magnetic, permitting concentration by magnetic means.
There are some deposits of consolidated monazite sands in
Wyoming.
Hydrometallurglcal processes are used to separate a thorium
and rare-earth concentrate from magnetically and gravity
concentrated sands (Figures 111-25 and 111-26). Either acid
or alkaline leach processes may be used, but cationic rather
than anionic species predominate in the leach, in contrast
with otherwise analogous uranium processes. Thorium preci-
pitates from sulfuric acid solution at a pH below one
(Figure 111-27), in contrast to rare earths and uranium;
this fact, as well as its reduced solubility in dilute mona-
zite sulfate solution, is utilized for thorium concentration.
The latter process, when used alone, requires as much as
300 liters (318 qt) of water per kilogram (2.2 Ib) of mona-
zite sulfate and is not very economical. When used in con-
junction with neutralizing agents as a fine control on pH,
it is very effective.
Recycle of leachant should be possible with an alkaline
leach process that has been evaluated in pilot-plant scale.
The process consumes caustic soda in the formation of tri-
sodium phosphate, which can be separated to some extent by
cooling the hot (110 to 137 degrees Celsius) (230 to 279
degrees Fahrenheit) leach to about 60 degrees Celsius (140
degrees Fahrenheit) and filtering. Uranium is precipitated
with the phosphate if NaOH concentration is too low during
the crystallization step, and NaOH concentration should be
raised to more than ION before cooling. The cyclic cooling
and heating of leach to separate phosphate values represents
an energy expenditure that must be weighed against the environ-
mental benefits of the process.
The alkaline leach process is unusual in that the leaching
action removes the gangue in the solute, as sodium silicate.
and leaves the values as rare-earth oxides, thorium, and
III-108
DRAFT
-------
DRAFT
Figure 111-25. SIMPLIFIED SCHEMATIC DIAGRAM OF SULFURIC ACID DIGESTION
OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
AND RARE EARTHS
TO
STOCKPILE
Fll
IU. A
SE
PREC
Al
*
FILTRATE
1
TO WASTE
— MAIN STREAM
I
RESIDUE
._ (UNDIGESTED MONAZITE SAND.
SILICA. ZIRCON. AND RUTILEI
FILTRATE
(R.E , U. AND P2O6)
|
SELE
PRECIP
AT(
CTIVE
IAIIUN •*
H23
I
t
-TRATE
ND P20g)
\
.ECTIVE
pH 60
t
*
PRECIPITATE OF
R E AND P2O5
}
PRECIPITATE OF
U AND PjOB
(BYPRODUCT)
TO SHIPPING
MONAZITE
SAND
*
GRINDING
OPERATION
1
DIGE!
1
DISSOl
1
1
1
Th. R E . U. AND PjOj 1 HjO
\
SELE
PRECIPI
ATp
(
TATION
NH
PRECIPITATE
ITh, R E . AND PjOjl
*
PURIFICATION BY SOLVENT EXTRACTION.
SELECTIVE PRECIPITATION. OR FRAC
TIONAL CRYSTALLIZATION
l
1
IO°C
i»FI
,OH
CONCENTRATES
TO SHIPPING
SOURCE- REFERENCE 20
III-109
DRAFT
-------
DRAFT
Figure 111-26. SIMPLIFIED SCHEMATIC DIAGRAM OF CAUSTIC SODA DIGESTION
OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
AND RARE EARTHS
MONAZITE
SAND
^^•^•B MAIN STREAM l •
IMSOH OPERATION
1 f
i
^___^ DIGESTION
f HYDROUS METAL-OXI
CRYSTALL.ZATION 1 Hh. U. AMU HA
* \
±Y DISSOLUTION
1 NaOH (Na3PO4) 1
.1
' ' fc SELECTIVE
^ PRECIPITATION
+
FILTRATE
(RARE EARTHS)
1 PRECIPITATE
(Th AND U)
^ SELECTIVP
"~ PRECIPITATION
1
1 *
FILTRATE! PRECIPITATE OF PURIFICATION
1 RARE EARTHS SOLVENT EXTRAC
JL (BY-PKODUCI) I
138°C
(280°F)
DE CAKE
.)
BY
;TIOIM
TO STOCKPILE CONCENTRATES h^- STOCKPILE
tllRrp- REFERENCE 2D . 1
III-110
DRAFT
-------
DRAFT
Figure 111-27. EFFECT OF ACIDITY ON PRECIPITATION OF THORIUM, RARE
EARTHS AND URANIUM FROM A MONAZITE/SULFURIC ACID
SOLUTION OF IDAHO AND INDIAN MONAZITE SANDS
100
o
LU
DC
SI
O
IDAHO MONAZITE SAND
a INDIAN MONAZITE SAND
A
I I 1 1
40 -
20 -
ACIDITY (pH)
AGITATION TIME:
DILUTION RATIO:
DIGESTION RATIO:
NEUTRALIZING AGENT:
SOURCE: REFERENCE 20
5 MINUTES
H2O: SAND - 45:1 TO 50:1
93% H2S04: SAND - 1.77
3.1% NH4OH
Ill-Ill
DRAFT
-------
DRAFT
uranium diuranate in Che residue. They are preserved as a
slurry or filter cake, which is then dissolved in sulfuric/
nitric acid and subjected to fractional precipitation, as
in the acid leach process.
The methods for recovering thorium and uranium from monazite
sands are almost identical to 'those used in the acid and
alkaline leach processes for recovering uranium from its
primary ores. Although thorium production in the U.S. is
currently not sufficient to characterize exemplary operations,
guidelines developed for the uranium mining and ore dressing
industry and its analogous subcategories should adequately
cover future operations for thorium.
Radiation parameters of thorium and uranium daughters are
somewhat different. The two decay series are compared in
Table 111-25. The uranium series is dominated by radium,
which—with a halflife of over 1,600 years and chemical
characteristics that are distinctly different from those of
the actinides and lanthanldes—can be separately concentrated
in minerals and mining processes. It then forms a noteworthy
pollutant entity that is discussed further in Section V.
Thorium, by contrast, decays via a series of daughters with
short halflives; the longest, Ra228, at 6.7 years, finds
use in luminous watches but lacks the long-term potency of
Ra226. In addition, by equilibrium rules, it can only attain
a concentration of less than one part in 2 billion. Radium
226 occurs at one part in 3 million in an equilibrium of
uranium and its daughters. Thus, radiation problems in
thorium mining are present but less critical than in the
case of uranium.
Industry Flow Charts. Of the sixteen mills operating in
1967 (Table 111-26) , no two used identical leaching concen-
tration, and precipitation steps. The same was probably
true of the 20 mills operating in 1972 (Table 111-23, also
Supplement B) . A general flow chart, to be used in con-
junction with Table 111-26, is presented in Figure 111-28.
Detailed flow charts of exemplary mills are presented in
Section VII.
Production Data. Recent uranium, vanadium, and radium pro-
duction data (U.S. Bureau of Mines 1972 data published in
1974) show that uranium production increased slightly in
anticipation of uranium use in commercial reactors. This
trend is expected to intensify until about 1975—fspecia]Jy,
in response to the energy situation of 1974. Thereafter,
delays in reactor licensing, the shortage of waste-fuel
storage space, and the difficulties involved in nuclear-waste
III-112
DRAFT
-------
DRAFT
TABLE 111-25. DECAY SERIES OF THORIUM AND URANIUM
ELEMENT OR
NAME
SYMBOL (S)
Thorium
Mesothorlum 1
Mesotttorlum 2
Radiothorlum
Thorium X
Thoron
Thorium A
Thorium B
Thorium C
Thorium C'
Thorium C"
Thorium D
Th232
90Th
ggRa228 (MsTh,)
ggAe^^ (MsThg)
gjjTh"^ (RrfTh)
B8Ra224 (ThX>
oaRn (Tn)
MPo218 (ThA)
82Pb212 (ThB)
83Bi212 (ThC)
84Po212 (ThC'J
g,™208 (ThC"l
gjPb208 (ThD)
HALF-LIFE
ENERGY OF RADIATION
(MeV)
* \ f
r
Thorium Series
1.34 x 1010 years
6.7 years
6.13 hours
1.90 years
3.64 days
64.5 seconds
0.158 seconds
10.6 hours
60.5 min
3 x 10* second
3.1 minutes
Stable
4.20
-
4.5
BM
6.68
6.28
6.77
-
6.05
8.77
-
-
—
0.053
1.55
-
-
—
^
0.36
2.20
-
1.82
-
-
-
-
r
-
-
-
-
r
-
2.62
-
Uranium Series
Uranium
Thorium
Protactinium
Uranium
Thorium
Radium
Radon
Polonium
Lead
Bismuth
Polonium
Thallium
Lead
Bismuth
Polonium
Lead
gjU238 (Ul)
gOTh^lUX,)
91Pe234 (UX2I
gjU234 <•*"*>
82Pb206
-------
DRAFT
TABLE 111-26. URANIUM MILLING PROCESSES
(a) 1967 Uranium Mills by Process
MILL
American Metal Climax
Anaconda
Atlas (Acid)
Atlas (Alkaline)
Cotter
Federal/American
Foote Mineral
United Nuclear/Homestake
Karr-McGee
Mines Development
Petrotomics
Susquehanna Western
UCC Uravan
UCC Gas Hills
Utah Construction & Mining
Western Nuclear
LEACH
Acid
Acid
Acid
Alkaline
Alkaline
Acid
Acid
Alkaline
Acid
Acid
Acid
Acid
Acid
Acid
Acid
Acid
CONCENTRATION
SX
RIP. IX
SX
RIP. IX
-
RIP. IX & SX
SX
-
SX
RIP. ix a sx
SX
sx
IX
RIP. IX
IX&SX
RIP. IX & SX
PRECIPITATION
H2°2
Lime/MgO
Ammonia
Ammonia
NaOH
Ammonia
MgO
NaOH
Ammonia
Ammonia
MgO
NaOH
Ammonia
Ammonia
Ammonia
Ammonia
VANADIUM
Salt roast
SX
-
-
-
SX
-
-
Na2 CO3 roast
IX
-
-
-
(b) Process by Number of Operations (1967)
ORE TREATMENT
Salt Roasting
Flotation
Pre-leach Density Control
LEACHING
Acid
Alkaline
2-Stage
LIQUID-SOLID SEPARATION
Countercurrent Decantation
Staged Filtration
Sand/Slime Separation
RESIN ION EXCHANGE (IX)
Basket Resin In
Pulp (Acid)
Basket RIP (Alkaline)
Continuous RIP
Fix Bed IX
Moving Bed IX
1
2
3
3
3
4
9
3
7
2
1
3
1
1
SOLVENT EXTRACTION (SX)
Amine
Alkyl Phosphoric
Eluex
PRECIPITATION
Lime/MgO
MgO
Caustic Soda (NaOH)
Ammonia )NH4OH)
Peroxide (HjOj)
VANADIUM RECOVERY
7
3
4
1
3
3
8
1
6
SOURCE. REFERENCE 21
III-114
DRAFT
-------
DRAFT
Figure 111-28. GENERALIZED FLOW DIAGRAM FOR PRODUCTION OF URANIUM,
VANADIUM. AND RADIUM
MINING
I
ORE TREATMENT
I
LEACHING
1
LIQUID/SOLID
SEPARATION
ION EXCHANGE
SOLVENT EXTRACTION
I
PATH!
I
PATHUI
PATHH
\
1
1
' T '
DDCfMDITATiriK
TO
STOCKPILE
I
URANIUM
CONCENTRATE
I
VANADIUM
BYPRODUCT
RECOVERY
TO
STOCKPILE
III-115
DRAFT
-------
DRAFT
disposal technology may cause some reduction in demand.
Table 111-27 shows uranium production for the period 1968
through 1972, expressed in terms of both ore movement and
UJto8 production and reserves. The reserves are estimated
to be recoverable at the traditional AEC stockpiling price
of $18/kg ($8/lb); with inflation, this price figure should
be revised upward. Reserves were seen to be increasing even
before this adjustment. They are presumably expanding even
faster when measured in terms of the energy to be extracted
from uranium. Additional uranium (and its derivative, plu-
tonlum) will become available if and when environmental
problems of fuel recycling are resolved—partlcuarly, when
breeder reactors become practical. The latter step alone
should Increase the economic ($18/kg) reserves, estimated
to last for about 20 years, to about 500 years.
Vanadium production, Table 111-28, is treated somewhat differ-
ently, since vandium is often an unwanted byproduct of uranium
mining and is only concentrated (recovered) when needed.
Value of the product fluctuates with demand, unlike uranium,
as indicated in the table. World production is also shown,
to indicate that U.S. production presents a fair fraction of
the world supply. The applications of vanadium are illus-
trated in Table 111-29.
Radium is traded from foreign sources, but not mined, in
quantities of about 40 grams (or curies) (0.14 ounce), at
a price of about $20,000/gram ($567,000/ounce) each year.
The high price is set by the historically determined cost
of refining and not by current demand. Reserves of radium
in uranium tailings are plentiful at this price. It has been
estimated that concentration of radium to prevent its discharge
to uranium tailings would approximately double the cost of
uranium concentrate.
Thorium production in the U.S. during 1968 was 100 metric
tons (110 short tons) as was demand, mostly for the chemical
and electronic uses. The U.S. imported 210 metric tons (231
short tons) to increase privately held stocks from 560 to
770 metric tons (616 to 847 short tons). Ihe General Services
Administration also held a stockpile of 1465 metric tons
(1612 short tons) which was intended to contain only 32
metric tons (35 short tons)—i.e., was in surplus by 1433
metric tons (1577 short tons).
III-116
DRAFT
-------
DRAFT
TABLE 111-27. URANIUM PRODUCTION
YEAR
1968
1969
1970
1971
1972
1973
ORE MOVEMENT
1000
METRIC TONS
5.861
6.367
6.749
5.708
6.834
6.162
1000
SHORT TONS
6,461
6.916
6,337
6,292
6,431
6.781
U3Og PRODUCTION
1000
METRIC TONS
11.244
10.664
11.732
11.167
11.727
12.032
1000
SHORT TONS
12.394
11.634
12.932
12.298
12.927
13.263
U3Og RESERVES*
1000
METRIC TONS
146
186
224
248
248
261
1000
SHORT TONS
161
204
247
273
273
277
•At $18.000 per metric ton ($16,340 par diort ton).
TABLE III-28. VANADIUM PRODUCTION
YEAR
1968
1969
1970
1971
1972
u.s.v2os
PRODUCTION
1000
METRIC
TONS
6,590
5.369
5.085
4,812
4,771
1000
SHORT
TONS
6,192
5.918
6,605
5,304
5,259
XOF
WORLD
46
31
27
28
26
WORLD V20g
PRODUCTION
1000
METRIC
TONS
12.119
16.892
18.337
16,883
18,136
1000
SHORT
TONS
13,359
18,620
20.213
18.610
19,990
V2OB VALUE
PER
METRIC
TON
$3,910
$5.190
$7.216
$7.887
$6.941
PER
SHORT
TON
$3.547
$4.708
$6.546
$7.156
$6.297
TABLE 111-29. VANADIUM USE
CATEGORY
Ferrovanadium
Vanadium Oxide
Ammonium Metavanadate
Vanadium Metal/alloys
1971
METRIC
TONS
3.792
130
32
412
SHORT
TONS
4.180
143
35
454
%
87
3
1
9
1972
METRIC
TONS
4,084
172
43
453
SHORT
TONS
4,502
190
47
499
%
86
4
1
9
III-117
DRAFT
-------
DRAFT
Metal Ores, Not Elsewhere Classified
This category Includes ores of metals which vary widely
in their mode of occurrence, extraction methods, and nature
of associated effluents. The discussion of metals ores
under this category which follows treats antimony, beryllium,
platinum, tin, titanium, rare-earth, and zirconium ores.
Thorium ores (monazite) have been previously discussed under
the Uranium, Radium, Vanadium category because of the similarity
of their extractive methods and radioactivity.
Antimony Ores
The antimony ore mining and milling industry is defined for
this document as that segment of industry involved in the
mining and/or milling of ore for the primary or byproduct/
coproduct recovery of antimony. In the United States, this
industry is concentrated in two states: Idaho and Montana.
A small amount of antimony also comes from a mine in Nevada.
Table 111-30 summarizes the sources and amounts of antimony
production for 1968 through 1972. The decrease in domestic
production during 1972 Indicated in Table 111-30 was largely
due to a fire which forced the major byproduct producer of
antimony to close in May of that year.
Antimony is recovered from antimony ore and as a byproduct
from silver and lead concentrates.
Only slightly more than 13 percent of the antimony produced
in 1972 was recovered from ore being mined primarily for its
antimony content. Nearly all of this production can be
attributed to a single operation which is using a froth
flotation process to concentrate stlbnite (Sb2£3_) (Figure
111-29).
The bulk of domestic production of antimony is recovered as
a byproduct of silver mining operations in the Coeur d'Alene
district of Idaho. Antimony is present in the silver-con-
taining mineral tetrahedrite and is recovered from tetra-
hedrite concentrates in an electrolytic antimony extraction
plant owned and operated by one of the silver mining companies
in the Coeur d'Alene district. Mills are usually penalized
for the antimony content in their concentrates. Therefore,
the removal of antimony from the tetrahedrite concentrates
not only Increases their value, but the antimony itself then
becomes a marketable item. In 1972, the price for antimony
was $1.25 per kilogram ($0.57 per pound).
III-118
DRAFT
-------
DRAFT
TABLE 111-30. PRODUCTION OF ANTIMONY FROM DOMESTIC SOURCES
YEAR
1968
1969
1970
1971
1972
ANTIMONY CONCENTRATE
METRIC TONS
4.774
5.176
6.060
4.282
1.879
SHORT TONS
S.263
5.707
6.681
4.721
2,072
ANTIMONY*
METRIC TONS
776
851
1.025
930
444
SHORT TONS
856
938
1.130
1,025
489
ANTIMONIAL LEAOt
(ANTIMONY CONTENT)
METRIC TONS
1.179
1.065
542
751
468
SHORT TONS
1.300
1.174
598
828
516
'Include! production from antimony ores and concentrate! and byproduct recovery from silver concentrate!
tByproduct produced at lead relmeriei in the United State*
III-119
DRAFT
-------
DRAFT
Figure 111-29. BENEFICIATION OF ANTIMONY SULFIDE ORE BY FLOTATION
ROUGHER
FLOTATION
FROTH
I
-TAILS'
MINING
ORE
J_
CRUSHING
GRINDING
I
CLASSIFICATION
•TAILS
^—
SCAVENGER
FLOTATION
FINAL
TAILINGS
TO
WASTE
CLEANER
FLOTATION
FROTH
I
FROTH
FILTER
I
THICKENER
FINAL
CONCENTRATE
F
TO SHIPPING
III-120
DRAFT
-------
DRAFT
Antimony Is also contained in lead concentrates and is ulti-
mately recovered as a byproduct at lead smelters. This
source of antimony represents about 30 to 50 percent of
domestic production In recent years.
Beryllium Ores
The beryllium ore mining and milling industry is defined for
this document as that segment of industry involved in the
mining and/or milling of ore for the primary or byproduct/
coproduct recovery of beryllium. Domestic beryllium produc-
tion data are withheld to avoid disclosing individual company
confidential data. During 1972, some beryl (Be3A12.(S16pl8))
was produced in Colorado and South Dakota. The largest
domestic source of beryllium ore is a bertrandite (Be4S12p_7_
(OH)2} mine in the Spor Mountain district of Utah. Domestic
beryl prices were negotiated between producers and buyers and
were not quoted in the trade press.
Mining and milling techniques for beryl are unsophisticated.
Some pegmatite deposits are mined on a small scale—usually,
by crude opencut methods. Mining is begun on an outcrop,
where the minerals of value can readily be seen, and cuts
are made or pits are sunk by drilling and blasting the rock.
The blasted rock is hand-cobbed, by which procedure as much
barren rock as practicable Is broken off with hand hammers
to recover the beryl. Beryl and the minerals it is commonly
associated with have densities so nearly the same that it is
difficult to separate beryl by mechanical means. Consequently,
beryl is recovered by hand cobbing.
A sulfuric acid leach process is employed to recover beryllium
from the Spor Mountain bertrandrlte. This Is a proprietary
process, however, and further details are withheld. No
effluent results from this operation.
Platinum-Group Metal Ores
The platinum-group metal ore mining and milling industry is
defined for this document as those operations which are
Involved in the mining and/or milling of ore for the primary
or byproduct/coproduct recovery of platinum, palladium,
Iridlum, osmium, rhodium, and ruthenium. These metals are
characterized by their superior resistance to corrosion
and oxidation. The industrial applications for platinum
and palladium are diverse, and the metals are used in the
production of high-octane fuels, catalysts, vitamins and
drugs, and electrical components. Domestic production of
111-121
DRAFT
-------
DRAFT
platinum-group metals is principally as a byproduct of copper
smelting, with production also from platinum placers.
Table 111-31 lists annual U.S. mine production and value
for the period 1968 through 1972.
The geologic occurrence of the platinum-group metals as
lodes or placers dictates that copper nickel, gold, silver,
and chromium will be either byproducts or coproducts in
the recovery of platinum metals, and that platinum will
be largely a byproduct. With the exception of occurrences
in the Stillwater Complex, Montana, and production as a
byproduct of copper smelting, virtually all the known
platinum-group minerals in the United States come from
placers. Platinum placers consist of unconsolidated alluvial
deposts in present or ancient stream valleys, terraces,
beaches, deltas, and glaciofluvlal outwash. The other
domestic source of platinum is as a byproduct of refining
copper from porphyry and other copper deposits and from lode
and placer gold deposits, although the grade is extremely
low.
Platinum-group metals occur in many placers within the United
States. Minor amounts have been recovered from gold placers
in California, Oregon, Washington, Montana, Idaho, and
Alaska, but significant amounts have been produced only from
the placers of the Goodnews Bay District, Alaska. Production
over the past several years from this district has remained
fairly constant, although domestic mine production declined
5 percent in quantity and 7 percent in value in 1972 (Refer-
ence 2) .
Beneficiation of Ores. The mining and processing techniques
for recovering crude platinum from placers in the U.S. are
similar to those used for recovering gold. The bulk of the
crude placer platinum is recovered by large-scale bucket-line
dredging, but small-scale hand methods are also used in
Columbia, Ethiopia, and (probably) the U.S.S.R. A flow
diagram for a typical dredging operation is presented as
Figure 111-30.
In the Republic of South Africa, milling and beneficiatlon
of platinum-bearing nickel ores consist essentially of
gravity concentration, flotation, and smelting to produce
a high-grade table concentrate called "metallic" for direct
chemical refining and a nickel-copper matte for subsequent
smelting and refining.
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TABLE 111-31. DOMESTIC PLATINUM-GROUP MINE PRODUCTION AND VALUE
YEAR
1968
1969
1970
1971
1972
MINE PRODUCTION
KILOGRAMS
460.1
671.4
638.6
660.8
532.2
TROY OUNCES
14.793
21,588
17,316
18,029
17,112
VALUE
$1,600,603
$2.094,607
$1.429.621
$1,369,676
$1,267,298
SOURCE: REFERENCE 2
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Figure 111-30. GRAVITY CONCENTRATION OF PLATINUM-GROUP METALS
DREDGE
(SCREENING,
JIGGING, AND
TABLING)
TABLING
MAGNETIC
SEPARATION
CHROMITE/
(MAGNETITE
DRYING
SCREENING
IT
SIZING
BLOWER
90% CONCENTRATE
(PLATINUM GROUP AND GOLD)
TO SHIPPING
TO
WASTE
TO
' SHIPPING
,TO
WASTE
III-124
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Byproduct platinum-group metals from gold or copper ores are
sometimes refined by electrolysis and chemical means. In
the Sudbury District of Canada, sulfide ore is processed
by magnetic flotation techniques to yield concentrates of
copper and nickel sulfides. The nickel flotation concentrate
is roasted with a flux and melted into a matte, which is cast
into anodes for electrolytic refining, from which the precious
metal concentrate is recovered.
In the U.S., the major part of output of platinum is recovered
as a byproduct of copper refining in Maryland, New Jersey,
Texas, Utah, and Washington. Byproduct platinum-group metals
from gold or copper ores are sometimes refined by electrolysis
and by chemical means. Metal recovery in refining is over
99 percent.
Rare-Earth Ores
The rare-earth minerals mining and milling Industry is defined
for this document as that segment of industry engaged in
the mining and/or milling of rare-earth minerals for their
primary or byproduct/coproduct recovery. The rare-earth elements,
sometimes known as the lanthanides, consist of the series
of 15 chemically similar elements with atomic numbers 57
through 71. Yttrium, with atomic number 39, is often Included
in the group, because its properties are similar, and it
more often than not occurs in association with the lanthanides.
The principal mineral sources of rare-earth metals are
bastnaesite (CeFC03_) and monazlte (Ce, La, Th, Y)P04_. The
bulk of the domestic production of rare-earth metals is
from a bastnaesite deposit in Southern California which is
also the world's largest known single commercial source of
rate-earth elements. In 1972, approximately 10,703 metric
tons (11,800 short tons) of rare-earth oxides were obtained
in flotation concentrate from 207,239 metric tons (approxi-
mately 228.488 short tons) of bastnaesite ore mined and milled
(Reference 2 ). Monazlte is domestically recovered as a
byproduct of titanium mining and milling operations in Georgia
and Florida. A company which recently began a heavy-mineral
(principally, titanium) sand operation in Florida is expected
to produce over 1.8 metric tons (2.0 short tons) of byproduct
monazlte annually.
At the Southern California operation, bastnaestite is mined
by open-pit methods. The ore, containing 7 to 10 percent
rare-earth oxides (REO) is upgraded by flotation techniques
to a mineral concentrate containing 63 percent REO. Calcite
is removed by leaching with 10 percent hydrochloric acid
III-125
DRAFT
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and countercurrent decantation. The bastnaesite is not
dissolved by this treatment, and the concentrate is further
upgraded to 72 percent REO. Finally, the leached product
is usually roasted to remove the carbon dioxide from the
carbonate, resulting in a product with over 90 percent REO.
Monzazite is recovered from heavy-mineral sands mined primarily
for their titanium content. Beneficiatlon of monazite is by
the wet-gravity, electrostatic, and magnetic techniques dis-
cussed in the titanium portion of this document. Monazite,
an important source of thorium, is also discussed under SIC
1094 (Uranium, Radium, and Vanadium). Extraction of the
thorium is largely by chemical techniques.
Tin Ores
The tin mining and milling industry is defined for this
document as that segment of industry engaged in the mining
and/or milling of ore for the byproduct/coproduct recovery
of tin.
There are no known tin deposits of economic grade or size
in the United States. Most of the domestic tin production
in 1972, less than 102 metric tons (112 short tons), came
from Colorado as a byproduct of molybdenum mining. In addition,
some tin concentrate was produced at dredging operations and
as a byproduct of placer gold mining operations in Alaska.
Feasability studies continue for mining and milling facilities
for a 4,065-metric-ton-per-day (4,472-short-ton-per-day)
open-pit fluorite tin/tungsten/beryllium mine in Alaska's
Seward Peninsula which is to open by 1976. Reserves at the
prospect area represent at least a 20-year supply. As tech-
nological improvements in beneficiation are made and demands
for tin increase, large deposits considered only submarginal
resources, in which tin in only one of several valuable
commodities, are expected to be brought into production.
In general, crude casslterlte concentrate from placer mining
is upgraded by washing, tabling,and magnetic or electrostatic
separation. Tin ore from lode deposits is concentrated by
gravity methods involving screening, classification, Jigging,
and tabling. The concentrate is usually a lower grade than
placer concentrate, owing to associated sulfide minerals.
The sulfide minerals are removed by flotation or magnetic
separation, with or without magnetic roasting. The majority
of tin production in the United States is the result of
beneficiation as a byproduct. Cassiterlte concentrate
recovery takes place after flotation of molybdenum ore by
III-126
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magnetic separation of the dewatered and dried tailings.
Despite considerable research, successful flotation of tin ore
has never been completely achieved.
Titanium Ores
The titanium ore mining and milling industry is defined for
this document as that segment of Industry engaged in the mining
and/or milling of titanium ore for Its primary or byproduct/
coproduct recovery. The principal mineral sources of titanium
are ilmenite (FeT102) and rutile (T102) . The United States
is a major source of ilmenite but not of rutile. Since 1972,
however, a new operation in Florida has been producing a
small amount (less than approximately 2 metric tons, or 2.2
short tons, per year) of rutile. About 85 percent of the
ilmenite produced in the United States during 1972 came from
two mines in New York and Florida. The remainder of the
production came from New Jersey, Georgia, and a second opera-
tion in Florida. A plant with a planned production of 168,000
metric tons (185,000 short tons) per year opened in New Jersey
during 1973. This plant and another which opened during 1972
in Florida are not yet at full production capability but are
expected to contribute significantly to the domestic production
of titanium in the future. Domestic production data are
presented in Table 111-32.
Two types of deposits contain titanium minerals of economic
importance: rock and sand deposits. The ilmenite from rock
deposits and some sand deposits commonly contains 35 to 55
percent TiO_2_; however, some sand deposits yield altered
ilmenite (leucoxene) containing 60 percent or more T102^ as
well as rutile containing 90 percent or more
The method of mining and beneficiatlng titanium minerals
depends upon whether the ore to be mined is a sand or rock
deposit. Sand deposits occurring in Florida, Georgia, and
New Jersey, contain 1 to 5 percent TiO£ and are mined with
floating suction or bucket-line dredges handling up to 1,088
metric tons (1,200 short tons) of material per hour. The
sand is treated by wet gravity methods using spirals, cones,
sluices, or jigs to produce a bulk, mixed, heavy-mineral
concentrate. As many as five individual marketable minerals
are then separated from the bulk concentrate by a combination
of dry separation techniques using magnetic and electrostatic
(high-tension) separators, sometimes in conjunction with dry
and wet gravity concentrating equipment.
Ill-127
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TABLE 111-32. PRODUCTION AND MINE SHIPMENTS OF TITANIUM
CONCENTRATES FROM DOMESTIC ORES IN THE U.S.
YEAR
1968
1969
1970
1971
1972
PRODUCTION*
METRIC TONS
887,508
884,641
787,235
619.549
618,251
SHORT TONS
978,509
931,247
867,955
683,075
681,644
SHIPMENTS*
METRIC TONS
870,827
809,981
835,314
647,244
661.591
SHORT TONS
960,118
893,034
920.964
713.610
729.428
•Includes a mixed product containing rutile, leucoxene. and altered ilmenite.
SOURCE: REFERENCE 2
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High-tension (HT) electrostatic separators are employed to
separate the titanium minerals from the silicate minerals.
In this type of separation, the minerals are fed onto a high-
speed spinning rotor, and a heavy corona (glow given off by high-
voltage charge) discharge is aimed toward the minerals at the
point where they would normally leave the rotor. The minerals of
relatively poor electrical conductance are pinned to the rotor by
the high surface charge they receive on passing through the high-
voltage corona. The minerals of relatively high conductivity
do not as readily hold this surface charge and so leave the rotor
in their normal trajectory. Titanium minerals are the only ones
present of relatively high electrical conductivity and are, there-
fore, thrown off the rotor. The silicates are pinned to the rotor
and are removed by a fixed brush.
Titanium minerals undergo final separation in induced-roll
magnetic separators to produce three products: ilmenite,
leucoxine, and rutile. The separation of these minerals is
based on their relative magnetic propertities which, in turn,
are based on their relative iron content: ilmenite has 37
to 65 percent iron, leucoxine has 30 to 40 percent iron, and
rutile has 4 to 10 percent iron.
Tailings from the HT separators (nonconductors) may contain
zircon and monazite (a rare-earth mineral). These heavy
minerals are separated from the other nonconductors (silicates)
by various wet gravity methods (i.e., spirals or tables).
The zircon (nonmagnetic) and monazite (slightly magnetic)
are separated from one another in induced-roll magnetic
separators.
Beneficiation of titanium minerals from beach-sand deposits
is illustrated in Figure 111-31.
Ilmenite is also currently mined from a rock deposit in New
York by conventional open-pit methods. This ilmenite/
magnetite ore, averaging 18 percent Ti02_, is crushed and ground
to a small particle size. The ilmenite and magnetite fractions
are separated in a magnetic separator, the magnetite being
more magnetic due to its greater iron content. The Ilmenite
sands are further upgraded in a flotation circuit. Beneficia-
tion of titanium from a rock deposit is illustrated in Figure
111-32.
Zirconium Ore
The zirconium ore mining and milling industry is defined for
this document as that segment of industry engaged in the
III-129
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Figure 111-31. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS
ORE FED
FROM DREDGE
\ |
VIBRATING
SCREENS
+
SPIRALS OR LAMINA
FLOWS (ROUGHERS AND CL
WET MILL
DRV MILL 1
SCRUBBER PLANT
*
DRIER
*
ELECTROSTATIC
SEPARATORS
|
SPIRALS AND/OR
TABLES
1
MAGNETIC
— -T— 1 ""
*
• MAGNETICS ' NONMAGNETICS— 1
1 MONAZITE | ( ZIRCON
___[ ' 1
TO TO 1
SHIPPING SHIPPING SHIf
TO
POND
"• SODIUM
^ HYDROXIDE
MAGNETIC
SEPARATOR
wnuunriurTirT. — .. — ™ MAGNETICS 1
TII c I 1 ILMENITE 1
J 1 '
, , ,
) TO
>P|NG SHIPPING
III-130
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Figure 111-32. BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
MINING
ORE
t
CRUSHING
GRINDING
±
CLASSIFICATION
I
MAGNETIC
SEPARATION
•MAGNETICS-
I
NONMAGNETIC*
MAGNETITE
i
ILMENITE
AND GANGUE
DEWATERER
i
FLOTATION
CIRCUIT
TAILINGS
T
i
THICKENER
TO
WASTE
FILTER
±
DRIER
1
CONCENTRATE
TO SHIPPING
III-131
DRAFT
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mining and/or milling of zirconium or for its primary or
byproduct/coproduct recovery.
The principal mineral source of zirconium (ZrS104_), is zircon
which is recovered as a byproduct in the mining of titanium
minerals from ancient beach-sand deposits, which are mined
by floating suction or bucket-line dredges. The sand is
treated by wet gravity methods to produce a heavy-mineral
concentrate. This concentrate contains a number of minerals
(zircon, ilmenite, rutile, and monazlte) which are separated
from one another by a cominatlon of electrostatic and mag-
netic separation techniques, sometimes used in conjunction
with wet gravity methods. (Refer to the titanium section
of this document.) Domestic production of zircon is currently
from three operations: two in Florida and one in Georgia.
The combined zircon capacity of these three plants is esti-
mated to be about 113,400 metric tons (125,000 short tons).
The price of zircon in 1972 was $59.50 to $60.50 per metric
ton ($54.00 to $55.00 per short ton). Zircon occurs with
titanium minerals In beach-sand deposits.
III-132
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SECTION IV
INDUSTRY CATEGORIZATION
INTRODUCTION
In the development of effluent limitations and recommended
standards of performance for new sources in a particular
industry, consideration should be given to whether the industry
can be treated as a whole in the establishment of uniform
and equitable guidelines for the entire industry or whether
there are sufficient differences within the industry to justify
its division into categories. For the ore mining and dressing
industry, which contains nine major ore categories by SIC
code (many of which contains more than one metal ore), many
factors were considered as possible justification for Industry
categorization and subcategorlzation as follows:
(1) Designation as a mine or mill;
(2) Type of mine;
(3) Type of processing (beneficiation, extraction
process);
(4) Mineralogy of the ore;
(5) End product (type of product produced);
(6) Climate, rainfall, and location;
(7) Production and size;
(8) Reagent use;
(9) Wastes or treatability of wastes generated;
(10) Water use or water balance;
(11) Treatment technologies employed;
(12) General geologic setting;
(13) Topography;
(14) Facility age;
(15) Land availability.
IV-1
DRAFT
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Because of their frequent use in this document, the defini-
tions of a mine and mill are included here for purposes of
recommending subcategorization and effluent limitation
guidelines and standards:
MINE
"A mine is an area of land upon which or under which
minerals or metal ores are extracted from natural deposits
In the earth by any means or methods. A mine includes the
total area upon which such activities occur or where such
activities disturb the natural land surface. A mine shall
also include land affected by such ancillary operations
which disturb the natural land surface, and any adjacent
land the use of which is incidental to any such activities;
all lands affected by the construction of new roads or the
improvement or use or existing roads to gain access to the
site of such activities and for haulage and excavations,
workings, impoundments, dams, ventilation shafts, entryways,
refuse banks, dumps, stockpiles, overburden piles, spoil
banks, culm banks, tailings, holes or depressions, repair
areas, storage areas, and other areas upon which are sited
structures, facilities, or other property or materials on
the surface, resulting from or incident to such activities."
MILL
"A mill is a preparation facility within which the
mineral or metal ore is cleaned, concentrated or otherwise
processed prior to shipping to the consumer, refiner, smelter
or manufacturer. This includes such operations as crushing,
grinding, washing, drying, sintering, briquetting, pelletiz-
ing, nodulizing, leaching, and/or concentration by gravity
separation, magnetic separation, flotation or other means.
A mill includes all ancillary operations and structures
necessary for the cleaning, concentrating or other process-
ing of the mineral or metal ore such as ore and gangue storage
areas, loading and shipping facilities."
IV-2
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Examination of the metal ore categories covered in this
document indicates that ores of 23 separate metals
(counting the rare earths as a single metal) are repre-
sented. Two materials are treated in two places in
this document: (1) vanadium ore is considered as a
source of ferroalloy metals (SIC 1061) and also in
conjunction with uranium/vanadium extraction under NRC
licensing surveillance (SIC 1094); and (2) monazite, listed
as a SIC 1099 mineral because it is a source of rare-earth
elements, also serves as an ore of a radioactive material
(thorium) and, therefore, is also treated in SIC 1094.
The discussion that follows is organized into five major areas
which illustrate the procedures and final selection of sub-
categories which have been made as part of these recommenda-
tions :
(1) The factors considered in general for all categories.
(Rationale for selection or rejection of each as
a pertinent criterion for the entire industry is
Included.)
(2) The factors which determined the subcategorization
within each specific ore category.
(3) The procedures which led to the designation of
tentative and, then, final subcategories within
each SIC code group.
(4) The final recommended subcategories for each ore
category.
(5) Important factors and particular problems pertinent
to subcategorization in each major category.
FACTORS INFLUENCING SELECTION OF SUBCATEGORIES IN ALL METAL
ORE CATEGORIES
The first categorization step was to examine the ore categories
and determine the factors influencing subcategorization for
the Industry as a whole. This examination evolved a list of
15 factors considered important in subcategorization of the
industry segments (as tabulated above). The discussion which
follows describes the factors considered in general for all
categories and subcategories. Rationale for selection or
rejection of each as a pertinent subcategorization criterion
is included.
IV-3
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Designation as a_ Mine or Mill
There are many reasons for recommending mine water discharge
limits different from mill effluents limits. There are
many mining operations which do not have an associated mill
or in which many mines deliver ore to a single mill, located
some distance away. In some instances where mine water is
used in a process, or is discharged to a treatment system
associated with a mill, the effluent limitations of the mill
will prevail. In many instances, it may be desirable to seg-
regate mine water from mill process water because of differing
water quality or ease of treatment with respect to water flow
volume, because of the parameters contained, or because of
parameter concentration. In general, levels of pollutants
in mine waters are lower or less complex than those in mill
process waters. Contact with finely divided ores (especially,
oxidized ores) is minimal, and mine water is not exposed to
the suite of process water reagents often added in milling.
The much smaller suspended-solid loads present in mine water
may be effectively removed in thickeners and later transported
to tailing ponds or other disposal sites, or may be settled
or treated in relatively permanent Impoundments without the
extensive dam raising necessary in.tailing disposal practice.
In addition, the volume of mine water is almost totally beyond
the operator's control and may greatly exceed mean water
flow. Effluent volume reduction is often not a viable option
in mine water treatment.
Type £f_ Mine
The choice of mining method is determined by the ore grade, size,
configuration, depth, and associated overburden of the orebody to
be exploited rather than by the chemical characteristics or
mineralogy of the deposit. Although interception of aquifers
occurs with both open-pit and underground mines and can create
difficulties associated with mine dewatering, it is largely
uncontrollable with respect to mine-operator choice of loca-
tion. Categorization as an open-pit or underground mine,
although the two types differ in the methods used for extraction
of ore, did not result in a usable scheme — that is, one which
reflects differences in water quality, ore mineralogy, or
other pertinent factors. Placer mines exploit unconsolidated
deposits for metal ores that can be concentrated largely by
gravity methods. Therefore, the mineralogy and extraction
methods become the dominant factors in determining subcategoi.-
ization. In addition, high suspended-solid loads can be
controlled by existing technology to levels equivalent to
existing underground and open-pit mine treated effluents.
IV-4
DRAFT
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Because the general geology is the determining factor in
selection of the mining method, and because no significant
differences resulted from application of control and treat-
ment technologies for mine waters from the above sources,
designation of the type of mine was not selected as a suitable
basis for subcategorization in the Industry.
Type of_ Processing (Beneficlation, Extraction Process)
The processing or beneficiation of ores in the ore mining
and dressing industry varies from crude hand methods to gravity
separation methods, froth flotation with extensive reagent
use, chemical extraction, and hydrometallurgy. Purely
physical processing using water provides the minimal pollu-
tion potential consistent with recovery of values from an ore.
All mills falling in this group are expected to share the
same major pollution problem—namely, suspended solids gener-
ated either from washing, dredging, crushing, or grinding.
The exposure to water of finely divided ore and gangue also
leads to solution of some material but, in general, treatment
required is relatively simple. The dissolved material will vary
with the ore being processed, but treatment is expected to be
essentially similar, with resultant effluent levels for impor-
tant parameters being nearly identical for many subcategories.
The practice of flotation significantly changes the character
of mill effluent in several ways. Generally, mill water pH
is altered or controlled to increase flotation efficiency.
This, together with the fact that ore grind is generally finer
than for physical processing, may have the secondary effect
of substantially increasing the solubility of ore components.
Reagents added to effect the flotation may include major
pollutants. Cyanide, for example, is used in several sub-
categories. Although usage is usually low, its presence in
effluent streams has potentially harmful effects. Oils are
also common flotation reagents which are undesirable in
effluent streams. The added reagents may have secondary
effects on the wastewater as well, such as in the formation of
cyanide complexes. The result may be to increase solubility
of some metals and decrease treatment effectiveness. Some
flotation operations may also differ from physical processors
in the extent to which water may be recycled without major
process changes or serious recovery losses.
IV-5
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Ore leaching operations differ substantially from physical
processors and flotation plants In wastewater character and
treatment requirements. The use of large quantities (in
relation to ore handled) of reagents, and the deliberate
solubillzation of ore components characterizes these opera-
tions. Wide diversity of leaching and chemical extraction
processes, therefore, affects the character and quantities
of water quality constituents, as well as the treatment and
control technologies employed.
To a large extent, mineralogy and extractive processes are
inextricable, because mineralogy and mineralogical variations
are responsible for the variations in processing technologies.
Both factors influence the treatability of wastes and efficiency
of removal of pollutants by treatment and control technologies.
Therefore, processing methods were a major factor in subcate-
gorizing each major ore category.
Mineralogy of the Ore
The mineralogy and host rock present greatly determine the
beneflciation of ores. Ore mineralogy and variations in
mineralogy affect the components present in effluent
streams and thus the treatability of the wastes and
treatment and control technology used. Some metal ores
contain byproducts and other associated materials, and some
do not. The specific beneficlatlon process adopted is based
upon the mineralogical characteristics of the ore; therefore,
the waste characteristics of the mine or mill reflect both
the ores mined and the extraction process used. For these
reasons, ore mineralogy was determined to be a primary
factor affecting subcategorization In all categories.
End Product
The end product shipped is closely allied to the mineralogy
of the ores exploited; therefore, mineralogy and processing
were found to be more advantageous methods of subcategorization.
Two ores, vanadium ores and monazite ores, are the exceptions
treated here which were based upon considerations of end
product or end use. Therefore, end product was not found
to be a suitable basis for categorization of the Industry
as a whole.
IV-6
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Climate, Rainfall, and Location
These factors directly Influenced subcategorization consid-
eration because of the wide diversity of yearly climatic
variations prevalent in the United States. Mining and
associated milling operations cannot locate in areas which
have desirable characteristics, such as many other industry
sedments. Therefore, climate and rainfall variations must
De accommodated or designed for. Some mills and mines are
located in arid regions of the country, allowing the use
of evaporation to aid in reduction of effluent discharge
quantity or attainment of zero discharge. Other facilities
are located in areas of net positive precipitation and high
runoff conditions. Two ore categories (i.e., the uranium
and copper ore industries) make primary use of the process-
ing method, followed by the secondary factors of climatic
conditions or rainfall as the basis for subcategorization.
Treatment of large volumes of water by evaporation in many
areas of the United States cannot be utilized where topo-
graphic conditions limit space and provide excess surface
drainage water. A climate which provides icing conditions
on ponds will also make control of excess water more diffi-
cult than in a semi-arid area. Therefore, limited use was
made of climate and rainfall as secondary subcategorization
factors.
Production and Size
The variation of size and production of operations in the
industry ranges from small hand cobbing operations to those
mining and processing millions of tons of ore per year.
The size or production of a facility has little to do with
the quality of the water or treatment technology employed,
but have considerable influence on the water volume and costs
incurred in attainment of a treatment level in specific
cases. Mills processing less than 5,000 short tons (4,535
metric tons) of ore per year in the ferroalloys industry
(most notably, tungsten) are typically intermittent in opera-
tion, have little or no discharge, and are economically mar-
ginal. Pollution potential for such operations is relatively
low due to the small volume of material handled if deliberate
solution of ores Is not attempted. Few of the operations are
covered by NPDES permits. Accordingly, size or production
was used in a limited sense for subcategorization in the
ferroalloys categories but was not found to be suitable for
the industry as a whole.
IV-7
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Reagent Use
The use of reagents in many segments of the industry, such
as different types of froth flotation separation processes,
can potentially affect the quality of wastewater. However,
the types and quantities of reagents used are a function of
the mineralogy of the ore and extraction processes employed.
Reagent use, therefore, was not a suitable basis for subcate-
gcrizatlon of any of the metals ores examined in this program.
Wastes or Treatability of Wastes Generated
The wastes generated as part of mining and beneficiating
metals ores are highly dependent upon mineralogy and pro-
cesses employed. In mines, however, mineralogy influences the
chemical nature of the wastewater. This characteristic was
used in subcategorization of lead and zinc mines.
Water Use and/or Water Balance
Water use or water balance is highly dependent upon choice
of process employed or process requirements, routing of mine
waters to a mill treatment system or discharge, and potential
for utilization of water for recycle in a process. Processes
employed play a determining role in mill water balance and,
thus, are a more suitable basis for subcategorization.
Treatment Technologies Employed
Many mining and milling establishments currently use a single
type of effluent treatment method today. While treatment
procedures do vary within the industry, widespread adoption
of these technologies is not prevalent. Since process and
mineralogy control treatability of wastes and, therefore,
treatment technology employed, treatment technology was not
used as a basis for subcategorization.
General Geologic Setting
The general geologic setting determines the type of mine—
i.e., underground, surface or open-pit, placer, etc.
Significant differences which could be used for subcategori-
zation with respect to geology could not be determined.
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Topography
Topographic differences between areas are beyond the control
of mine or mill operators and largely place constraints on
treatment technologies employed, such as tailing pond loca-
tion. Topographic variations can cause serious problems with
respect to rainfall accumulation and runoff from steep slopes.
Topographic differences were not found to be a practical basis
or which subcategorization could be based, but topography is
known to influence the treatment and control technologies
employed and the water flow within the mine/mill complex.
Facility Age
Many mines and mills are currently operating which have oper-
ated for the past 100 years. In virtually every operation
involving extractive processing, continuous modification of
the plant by Installation of new or replacement equipment
results in minimal differences for use in subcategorization
within a metal ore category. Many basic processes for con-
centrating ores in the Industry have not changed considerably
(e.g., froth flotation, gravity separation, grinding and
crushing), but improvements in reagent use and continuous
monitoring and control have resulted in Improved recovery
or the extraction of values from lower grade ores. New and
innovative technologies have resulted in changes of the
character of the wastes, but this is not a function of age
of the facilities, but rather of extractive metallurgy and
process changes. Virtually every facility continuously
updates in-plant processing and flow schemes, even though
basic processing may remain the same. Age of the facility,
therefore, is not a useful factor for subcategorization in
the Industry.
DISCUSSION OF PRIMARY FACTORS INFLUENCING SUBCATEGORIZATION
BY ORE CATEGORY
The purpose of the effluent limitation guidelines can be
realized only by categorizing the industry into the minimum
number of groups for which separate effluent limitation
guidelines and new source performance standards must be
developed. The categorization presented here is believed to
be the least number of groups having significantly different
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water-treatment problems and water-pollution potentials.
The best technology, from a water-quality standpoint, would
be a completed closed system with all mining and processing
water being recycled. This technology should be considered
for any new sources and for existing operations which have
the favorable processes and mineralogies, water flows and
quality requirements, location, and land availability that
will allow economic practice.
This section outlines and discusses briefly the factors which
were used to determine the subcategories within each ore
category. A presentation of the procedures leading to the
tentative and then final subcategories, together with a
listing of the final recommended subcategories, is included.
The treatment by ore category also includes a brief dis-
cussion, where applicable, of important factors and pertinent
problems which affect each category.
Iron Ore
In developing a categorization of the iron ore industry, the
following factors were considered to be significant in providing
a basis for categorization. These factors include character-
istics of individual mines, processing plants, and water uses.
1. Type of Mining
a. Open-Pit
b. Underground
2. Type of Processing
a. Physical
b. Physical - Chemical
3. Mineralogy of the Ore
4. General Geologic Setting, Topography, and Climate
(also Rainfall and Location)
Information for the characterization was developed from pub-
lished literature, operating company data, and other informa-
tion sources discussed in Section III.
As a result of the above, the first categorization developed
for the iron mining and beneficiation industry was based on
whether or not a mine or mill produces an effluent. This
Initial categorization considered both the mining and milling
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water circuits separately, as well as a category where mines
and mills were in a closed water system. The resulting
tentative subcategories which resulted are presented in the
listing given below:
I. Mine producing effluent - processing plant with
a closed water circuit.
Ha. Mine producing effluent - processing plant producing
an effluent - physical processing.
lib. Mine producing effluent - processing plant producing
an effluent - physical and chemical processing.
III. Mine and processing plant with a closed water circuit.
Examination of the preliminary subcategorlzation and further
compilation of information relative to iron mining and processing
methods resulted in a classification of the mines and mills
into the following order by production:
Open-Pit Mining, Iron Formation, Physical Processing
Open-Pit Mining, Iron Formation, Physical and Chemical
Processing
Open-Pit Mining, Natural Ores, Physical Processing
Underground Mining, Iron Formation, Physical Processing
Underground Mining, Iron Formation, Physical and Chemical
Processing
Underground Mining, Natural Ores, Physical Processing
In preparation for selection of sites for visitation and
sampling, the operations were further classified on the basis
of size, relative age, and whether they had closed water
systems or produced an effluent from either the mining or
processing operation:
Operation A
High tonnage Older plant (1957)
Open-pit Mine produces effluent
Iron formation Processing plant has closed water
system
Physical processing
Operation B
Medium tonnage Medium age plant (1965)
Open-pit Mine produces effluent
Iron formation Processing plant has closed water
system
Physical processing
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Operation C
Medium tonnage
Open-pit
Natural ore
Physical processing
Operation D
Low tonnage
Open-pit
Natural ore
Physical processing
Operation E
High tonnage
Open-pit
Iron formation
Physical processing
Operation G
Low tonnage
Open-pit
Iron formation
Physical and chemical
processing
Operation H
Medium tonnage
Open-pit
Iron formation
Physical and chemical
processing
Operation I
Medium tonnage
Open-pit
Iron formation
Physical and chemical
processing
Operation J
Low tonnage
Underground
Iron formation
Physical and chemical
processing
Older plant (1948)
No effluent
Older plant (1953)
Mine produces effluent
Processing plant produces effluent
Medium age plant (1967)
Mine produces effluent
Processing plant has closed
water system
Older plant (1959)
Mine produces effluent
Processing plant produces
effluent
Older plant (1956)
Mine produces effluent
Processing plant produces effluent
Medium age plant (1964)
Mine produces effluent
Processing plant produces effluent
Older plant (1958)
Mine produces effluent
Processing plant produces effluent
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The mines visited and sampled had a 1973 production of approxi-
mately 43,853,450 metric tons (48,350,000 short tons), or 47.5
percent of the total United States production of iron ore.
One of the Initial goals of this study was determination
of the validity of the Initial categorization. The primary
source of the data utilized for this evaluation was information
obtained during this study, plant visits, and sampling program.
This Information was supplemented with data obtained through
personal interviews and literature review and with historical
effluent quality data from NPDES permit applications and monitor-
ing data supplied by the iron mining and beneficiating industry.
Based on this exhaustive review, the preliminary industrial
categorization was substantially altered.
The data review revealed two distinct effluents from the
mining and milling of iron. The first (I) coming from the
mines and second (II) coming from the mills. It was also
determined that all mills in general could not be classed
together. This is primarily because a large number of milling
operations achieve zero discharge without major upset to pre-
sently used concentrating technology.
The milling categorized into four distinct classes based
on the type of ore, geographical location, and the type of
processing.
Category Ha. Mills using physical separation techniques,
exclusive of magnetic separation (washing,
Jigging, cyclones, spirals, heavy media).
Category lib. Mills using flotation processes and using
the addition of chemical reagents.
Category lie. Mills using magnetic separation and not
beneficiating ores of the Minnesota
Biwablk formation.
Category lid. Mills using magnetic separation for the
benefication of iron formations in the
Minnesota Biwabik formation.
Final Iron-Ore Subcategorization. Based on the types of
discharges found from all mills, the first three subcategories
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can be grouped into a single segment. Mills in the Minnesota
Mesabi Range have demonstrated that a distinct subcategory
can be made because of hydrological characteristics of the
area, type of ore, and the mode of beneficlation.
I. Mines
Open-pit or underground, removing natural ores
or iron formations.
II. Iron ore mills employing physical and/or chemical
separation
III. Iron ore mills employing magnetic and physical separa-
tion (Mesabi Range)
Copper Ores
The copper-ore subcategorization consideration began with
the approach that mineralization and ore beneficiating or
process method were intimately related to one another.
This relationship together with a basic division into mining,
milling and hydrometallurgical processing resulted in a
preliminary subcategorization scheme based primarily on
division into mine or concentrating facility and then further
based the method of concentrating or extraction of values
from the ore. Examination of water quality data supplied
by the Industry and other sources indicated that division of
mills into further subcategorles based upon process resulted
in grouping operations with similar water quality character-
istics. Other factors such as climate and rainfall presented
problems of subcategorization particularly with respect to
conditions prevalent in certain areas during approximately
two months of the year.
Final Copper-pre Subcategorization
Based on data collected from existing sources In addition
to visits and sampling of copper mines and extraction
facilities, the following final subcategories have been
established based primarily on designation as a mine or con-
centrating or chemical extraction facility:
I. Mines
Open-pit or underground, removing sulfide,
oxide, mixed sulfide oxide ores, or native
copper.
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II. Copper mines employing hydrometallurgical processes
III. Copper mills employing the vat-leaching process
IV. Copper mills employing froth flotation (areas
where net evaporation equals or exceeds 76.2
centimeters (30 inches) per year
V. Copper mills employing froth flotation (areas
where net evaporation is less than 76.2 centi-
meters (30 inches) per year.
Problems in Subcategorizing the Copper Industry. Copper is
produced in many areas of the United States which vary in
mineralization, climate, topography, and-process-water source.
The processes are outlined in Section V, but the froth flota-
tion of copper sulfide is adjusted to conditions at each
plant and will also vary from day to day with the mill feed.
Excess runoff from rainfall and snow melt do alter the sub-
categorization, but they can be controlled by enlargement of
tailing ponds and construction of diversion ditching. Pre-
sently, mine drainage is sent to tailing lagoons, although
a decrease in excess water problems can be realized in many
cases if mine water,is treated separately from mill process
water. For this reason, the mining subcategory remains inde-
pendent of mill and hydrometallurgical beneficiating processes.
Dissolved salt buildup may cause problems in the recycling
of mill process waters, when the makeup water source and/or
ore body contain a high content of dissolved salts. Additional
treatment of the process water for removal of some of the waste
constituents may be necessary for recycle of process water and
may produce a zero effluent from many plants where buildup of
materials may adversely affect recovery.
Lead and Zinc Ores
As a result of an initial review of the lead/zinc mining and
milling Industry which considered such factors as mineralogy of
ore, type of processing, size and age of facility, wastes and
treatability of waste, water balance associated with the facilities*,
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land availability, and topography, a preliminary scheme for
subcategorlzation of the lead/zinc industry was developed.
The preliminary analysis disclosed that size and age of a
facility should have little to do with the characteristics
of the wastes from these operations in that the basic flota-
tion cells have not changed significantly in a decade.
The reagents used, even in very old facilities, can be utilized
the same as in the newest. These factors, in addition to
life cf an ore body, and such factors as land availability,
topography, and, perhaps, volume of water which must be removed
from a mine have little to do with technology of treatment
but can have considerable effect on the cost of a treatment
technology employed in a specific case. These factors which
effect the economics of treatment at specific facilities, if
considered as a basis for subcategorizatlon, would, if carried
to their logical completion, result in individual considera-
tion for each facility.
The preliminary subcategorization scheme utilized was selected
to provide subcategorlzation on basic technological factors
where possible. The factors considered in the preliminary
scheme were:
I. End Product Recovered:
(a) Lead/zinc
(b) Zinc
(c) Lead
(c) Others with lead/zinc byproducts
II. Designation as a Mine or Mill:
(a) Mine
(b) Mill
(c) Mine/mill complex
III. Type of Processing:
(a) Gravity separation (no reagents)
(b) Flotation
IV. Wastes or Treatability of Wastes Generated:
(a) Potential for development of conditions
with soluble undesirable metals or salts
(b) No potential for solubilization
V. Water Balance:
(a) Total recycle possible
(b) Total recycle not possible
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The plant visits and subsequent compilation of data and
literature review were aimed at establishing which factors
were really significant in determining what effluent quality
could be achieved with respect to the tentative subcategori-
zation.
An analysis of the data compiled indicated that subcategori-
zation within the lead/zinc industry could be simplified
cont>iderably. No basic differences In raw waste characteristics
or treatability were found to be associated with the type of
concentrates obtained from a facility.
The proposed subcategorization based on what facility is
discharging—that is, a mine or a mill—is justified because
effluents from a mine dewaterlng operation and those
from a milling operation, into which various chemicals may
be introduced, are different. In the case of a mine dis-
charging only into the water supply of the mill, the only
applicable guideline would be that of the mill.
No evidence of current practice of strictly physical concen-
tration by gravity separation was found. The recovery of
desirable minerals from known deposits utilizing only such
physical separations is likely to be so poor as to result in
discharge of significant quantities of heavy-metal sulflde
to the tailing retention area. The only ore concentration
process currently practiced in the lead/zinc industry is froth
flotation. Subcategorization based on milling process is,
therefore, not necessary.
The treatability of mine wastewater is significantly affected
by the occurrence of local geological conditions which cause
solubilizatlon of undesirable metals or salts. A common,
and well-understood, example is acid mine drainage caused
by the oxidation of pyrite (FeS2) to ferrous sulfate and
sulfuric acid. This oxidation requires both moisture and
air (oxygen source) to occur. The acid generated then leaches
heavy metals from the exposed rock as particle surfaces.
Heavy metals may also enter solution as a result of oxidation
over a period of time through fissured ore bodies to form
more soluble oxides of heavy metals (such as zinc) in mines
which do not exhibit acidic mine drainages. Another route
which may result in solubilized heavy metals involves the
formation of acid and subsequent leaching in very local
areas in an ore body. The resultant acid may be neutralized
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by later contact with limestone or dolomitic limestone, but
the pH level attained may not be high enough to cause pre-
cipitation of the solubilized metals. The important aspect
of all of these situations is that the mine water encountered
is much more difficult to treat than those where solubiliza-
tion conditions do not occur. The treated effluents from
mines in this subcategory often exhibit higher levels of
heavy metals in solution than untreated mine waters from
mines where solubilization conditions do not occur.
The water-balance parameter, of course, does not apply to mine-
only operations. In the case of milling operations, system design
and alteration of process flows can have considerable effect on
the water balance of a milling operation. No justification was
found for substantiation of subcategorlzation on this basis.
The final recommended subcategorization for the lead/zinc
mining and milling industry is, therefore, condensed to:
1. Lead and/or zinc mines having no solubilization potential
II. Lead and/or zinc mines having solubilization potential
III. Lead and/or zinc mills
For purposes of the subcategorization recommended here, solu-
bilization potential in the lead/zinc mining industry is defined
as total heavy metal concentration in untreated mine wastewater
equal to or exceeding 2 mg/1 for the sum of the concentrations
of Cd, Cr, Cu, Pb and Zn.
Problems Affecting Subcategorization. The separation of a
specific discharge into the mine or mill subcategory is simple
and straightforward. The further subcategorization of mine
water based on the local geologic conditions which lead to
solubilization is, clearly, not so straightforward. It is
necessary, however, because a treatment practice appropriate
for mine waters not affected by this condition is relatively
ineffective in controlling wastes from mines in this subcate-
gory. Indeed, treatment consisting of chemical precipitation,
followed by sedimentation, is required to attain effluent
quality approaching the raw waste chemical composition dis-
charged from the typical non-affected mine. Assignment
of a mine to this subcategory will be straightforward
if the pH of the mine water is significantly below neutral
(i.e., if it is 6.0 or below). Those situations where the
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mine water is alkaline are not so easily classified. The
manifestation of a mine falling in this category is high
soluble metals (particularly, zinc) in the effluent. The
cause, as previously discussed, is related to the geology
of the formation being mined.
In general, the type of treatment practices currently utilized
and those envisioned are not energy-intensive technologies;
therefore, energy consideration on an overall basis has
minor impact. In specific cases, however, topographical con-
sideration may create instances where significant energy
would be expended in pumping water for treatment.
In general, it is prudent to minimize the volume of the waste
stream which must be treated. It is, therefore, good practice
to segregate runoff from non-contaminated areas from that of
areas which will require treatment.
Gold Ores
The most important factors considered in determining whether
subcategorization was necessary for the gold ore category
were ore mineralogy, general geologic setting, type of
processing, wastes and waste treatability, water balance, and
final product. Upon intensive background data compilation
(as discussed in Section III), mill inspections, and communi-
cations with the industry, most of the factors were found to
reduce to mineralogy of the ore (and, thus, product) and mill-
ing process employed. The initial subcategorization was found
to differ little from final subcategorization selection after
site visitation and sampling data were obtained.
The most effective means of categorizing the gold industry
is based upon relative differences among existing sources
of discharge (mine or mill/mine-mill complexes) and on
characteristics of the beneficiation process. The rationale
for this is based on several considerations:
(1) Apart from milling processing, the charac-
teristic difference between mine effluents and
mill/mine-mill effluents is their quantitative
and qualitative pollutant loadings. This differ-
ence between mines and mills makes necessary the
application of differing waste-treatment tech-
nologies and/or the segregation of sources for
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purposes of treatment. A mill effluent normally
contains a greater quantity of total solids—up
to 40 to 50 percent more than a mine effluent.
Much of these solids are suspended solids, and
treatment involves removal by settling. This is
usually treated in tailing ponds. Where mines
occur alone, or where their effluents are treated
separately from the mill, these effluents may be
treated on a smaller scale by a different tech-
nology.
(2) The specific beneficiation process adapted is
based on the geology and mineralogy of the ore.
The waste characteristics and treatabllity of
the mill effluent are a function of the particu-
lar beneficiation process employed. This takes
into account the reagents used and the general
mineralization of the ore by each particular
process as these factors affect differing waste
characteristics. The waste characteristics affect
treatabllity; for example, cyanide removal requires
different technology than that used for metal
removal.
Consideration was also given to the regional availability
of water, as this factor is relevant to water conservation
and "no discharge" and waste-control feasibility. Since it
is common engineering practice to design tailing ponds to
accommodate excesses of water, and also since pond design
can include systems to divert surface runoff away from the
pond, regional availability of water was judged not to be
a limiting factor with respect to the feasibility of a no-
discharge system.
Final GoId-Ore Subcategorization
On the basis of the rationale developed above and previously
discussed in the introductory portion of this section, six
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subcategories were identified for the gold mining and milling
industry:
I. Hlne(s) alone.
II. Mill(s) or mine/mill complex(ea) using the process
of cyanidation for primary or byproduct
recovery of gold.
III. Mill(s) or mine/mill complex(es) using process of
amalgamation (includes dredging operations,
if amalgamation is used).
IV. Mill(s) or mine/mill complex(es) using the process
of flotation.
V. Mill(s) or mine/mill complex(es) using gravity
separation (includes dredging or hydraulic
mining operation).
VI. Mill(s) where gold is a byproduct of a base-metal
operation. (Gold values are present in the
base metal concentrate and are recovered at
the smelter.)
Silver Ores
The development of subcategorization in the silver industry
was essentially identical to that of the gold industry
previously discussed. The primary basis for division into
subcategories was mineralogy of the ore and type of process-
ing. Since mineralogy and type of extraction processing are
intimately related, these factors served, Just as in the gold
industry, to divide the industry into mine and mill categories,
and then further into milling categories based upon type of
processing. Also note that, in many places, gold and silver
are exploited as coproducts or, together, as byproducts of
other base metals (such as copper).
Final Silver-Ore Subcategorization
Based upon the previous rationale developed in the intro-
ductory portion of this section (and also discussed in con-
nection with gold ores), tentative subcategorization was
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developed and then verified by field sampling and site visits.
Based upon field confirmation, the tentative subcategories,
found to be unchanged, are:
I. Mine(s) alone
II. Mill(s) or mine/mill complex(es) using flotation
for primary or byproduct recovery of silver.
III. Mill(s) or mine/mill complex(es) using cyanidation
for primary or byproduct recovery of silver.
IV. Mill(s) using amalgamation process for primary
or byproduct recovery of silver.
V. Mill(s) using gravity separation process for primary
or byproduct recovery of silver.
VI. Mill(s) or mine/mill complex(es) where silver is
recovered as a byproduct of a base-metal
operation. (Silver is present in the base-
metal concentrate(s) and is recovered at the
smelter.)
Bauxite Ore
In the bauxite mining industry, most criteria for subcate-
gorizatlon bear directly or indirectly upon two basic factors:
(1) nature of raw mine drainage, which is a function of the
mineralogy and general geological setting related to percolating
waters; and (2) treatability of waste generated, based upon
the quality of the effluent concentrations. Initially, general
factors, such as end products, type of processing, climate,
rainfall, and location, proved to be of minor importance as
criteria for subcategorization. The two existing bauxite min-
ing operations are located adjacent to one another in Arkansas
and share similar rainfall and evaporation rates, 122 cm (48 in.)
and 109 cm (43 in.). Both operations produce bauxite, though
slightly different in grade, which is milled by a process emit-
ting no wastewater.
After the site visits to both operating mines, it was evident
that the mining technique is closely associated with the
characteristics of the mine drainage, and that mineralization
is directly responsible for mining-technique and raw mine-
drainage characteristics. In addition, an evaluation of
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removal efficiency for a treatment process common to both
members of the industry became the prime consideration in
determining attainable treated effluent concentrations.
Final Bauxite-Ore Subcategorization
Based on the results of intensive study, facility inspections,
NPDES permit applications, and communication with the industry,
it was concluded that the bauxite mining and milling industry
should not be subcategorized beyond that presented below.
Bauxite mining and associated milling operations
(essentially grinding and crushing)
Ferroalloy Ores
In development of subcategories for the ferroalloy mining
and milling category, the following factors were considered
initially: type of process, and product, mineralogy, climate*
topography, land availability, size, age, and wastes or
treatability of wastes generated.
A tentative Subcategorization of the industry was developed after
collection and review of Initial data, based primarily on end
product (e.g., tungsten, molybdenum, manganese, etc.), with
further division on the basis of process, in some cases.
Further data, particularly chemical data on effluents and
more complete process data for past operations, indicated
that process was the dominant factor influencing waste-stream
character and treatment effectiveness. Examination of the
industry additionally showed that size of operation could
also be of great importance. Other factors, except as they
are reflected in or derived from the above, are not believed
to warrant industry Subcategorization.
Final Ferroalloy-Ore Subcategorization
It has been determined that the ferroalloy mining and milling
category should be divided into five subcategories for the
purpose of establishing effluent limitations and new source
performance standards:
I. Mines discharging water.
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II. Mills processing less than 5,000 metric tons
(5,512 short tons) per year of ore by methods
other than ore leaching.
III. Mills processing more than 5,000 metric tons per
year of ore by purely physical methods (e.g.,
crushing, ore washing, gravity separation,
and magnetic and electrostatic separation).
IV. Mills processing more than 5,000 metric tons per
year of ore and employing flotation.
V. Mills practicing ore leaching and associated
chemical beneficiation techniques.
The subcategory including mills processing less than 5,000
metric tons of ore per year is representative of operations
which are typically both intermittent in operation and eco-
nomically marginal. This subcategory is believed to contain,
at present, almost exclusively processors of tungsten ores.
Purely physical processing provides the minimum pollution
potential consistent with recovery of values from an ore
using water. All mills falling into this subcategory are
expected to share the same major pollution problem—namely,
suspended solids generated by the need for crushing and
grinding. The exposure of finely divided ore (and gangue)
to water may also lead to solution of some material, but,
in general, pretreatment levels will be low and treatment,
relatively simple. The dissolved material will clearly vary
with the ore being processed, but treatment is expected to
be essentially the same in all cases and to result in
similar maximum effluent levels. There are currently no
active major water-using physical processors in the ferro-
alloy Industry except in the case of nickel, where water use
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is not really in the process. Information has been drawn
heavily, therefore, from past data and related milling
operations—particularly, in the iron ore industry. The
close relationship between iron ores and manganiferous ores,
where such production is likely in the near future, as well
as the nature of the data itself, makes this transfer reason-
able. These milling processes are fully compatible with
recycle of all mill water.
The practice of flotation significantly changes the character
of mill effluent in several ways. Generally, mill water pH
is altered or controlled to increase flotation efficiency.
This, together with the fact that ore grind is generally
finer than for physical processing, may have the secondary
effect of substantially Increasing solubility of ore compon-
ents. Reagents added to effect the flotation may include
major pollutants. Cyanide, for example, is commonly used
and, though usage is low, may necessitate treatment. Oils
are also common flotation reagents which are undesirable in
effluent streams. The added reagents may have secondary
effects on the effluent as well; the formation of cyanide
complexes, for example, may increase solubility of some
metals and decrease treatment effectiveness. Some flotation
operations may also differ from physical processors in the
extent to which water may be recycled without process changes
or serious recovery losses.
Ore leaching operations differ substantially from physical
processors and flotation plants in effluent character and
treatment requirements. The use of large quantities (in
relation to ore handled) of reagents, and the deliberate
solubllization of ore components, characterizes these opera-
tions. The solubllization process is not, in general,
entirely specific, and the recovery of desired material is
less than 100 percent. Large amounts of dissolved ore may
be expected, therefore, to appear in the mill effluent,
necessitating extensive treatment prior to discharge. For
these operations, even commonly occurring ions (i.e., Na+,
SOA^, etc.) may be present in sufficient quantities to cause
major environmental effects, and total dissolved-solid levels
can become a real (although somewhat intractable) problem.
Wide variations in leaching processes might justify further
division of this subcategory (into acid and alkaline leach-
ing, as in the uranium industry, for example), but the limited
current activity and data available at this time do not sup-
port such a division.
IV-25
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Other Considerations. Climate, topography, and land avail-
ability are extremely important factors influencing effluent
volume, character, and treatment in the mining and milling
industry—particularly, the attainment of zero pollutant dis-
charge by means of discharge elimination. Zero discharge
may be attainable, for example, despite a net positive water
balance for a region because rainfall input to a tailing
impoundment balances part of the process water loss, includ-
ing evaporative losses in the mill and retention in tails.
On the other hand, it may be unattainable despite a negative
annual balance due to severe seasonal fluctuations, coupled
with soil porosity, which renders diversion ineffective, or
other topographic factors. Particular situations may also
exist where other environmental effects, such as massive
energy consumption, make zero discharge unfeasible. It is
anticipated that, under the Impetus of effluent limitations
established under PL 92-500, and the resultant pollution con-
trol costs, many mills in the defined subcategories will
choose the often less expensive option of discharge elimina-
tion.
Mercury Ores
The mercury industry in the United States currently is at a
reduced level of activity due to depressed market prices.
One facility was found to be operating at present,
although it is thought that activity will again increase
with increasing demand and rising market prices. The
decreased use of mercury due to stringent air and water
pollution regulations in the industrial sector may be offset
in the future by increased demand in dental and other uses.
Very little beneficiatiug of mercury ores is known in the
industry. Common practice for most producers (since rela-
tively low production characterizes most operators) is to
feed the cinnabar-rich ore directly to a kiln or furnace
without beneficiation. Water use in most of the operations
is at a minimum, although a rather large (20,000-flask-per-
year, or 695-metric-ton-per-year or 765-short-ton-per-year)
flotation operation with high water use is expected to be
operating in the near future. In the year 1985, the industry
could be producing 3,000 to 20,000 flasks (104 to 695 metric
tons, or 115 to 765 short tons) per year, depending on market
price, technology, and ore grade (U.S. Bureau of Mines pro-
jection) .
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Final Mercury Ore Subcategorization
Since most mercury operations are direct furnacing facilities,
the resulting Subcategorization represents that fact. Little
or no beneficiation is done in the industry, with few excep-
tions. There are a few operations from which mercury is
recovered as a byproduct at a smelter or refinery. A single
known flotation operation is expected in the near future and
is reflected in the Subcategorization scheme below based on
processing.
I. Mine(s) alone or mine(s) with crushing and/or
grinding prior to furnacing (no additional
beneficiation).
II. Mill(s) or mine/mill complex(es) using the process
of gravity separation for primary or byproduct
recovery of mercury.
III. Mill(s) or mine/mill complex(es) using flotation
for primary or byproduct recovery of mercury.
IV. Mill(s) where mercury is recovered as a byproduct
of base- or precious-metal concentrates.
(The recovery takes place at a refinery or
smelter.)
Uranium. Radium, and Vanadium Ores
The primary factors evaluated in consideration of Subcate-
gorization of the uranium, radium, and vanadium mining and
ore dressing Industry are: end product, type of processing,
ore mineralogy, waste characteristics, treatability of
wastewater, and climate, rainfall, and location. Based upon
an intensive literature search, plant inspections, NPDES permits,
and communications with the industry, this category is
categorized by milling process, mineralogy (and, thus,
product), and climatic factors. A discussion of each of the
primary factors as they affect the uranium/radium/vanadium
ore category follows.
The milling processes of this industry involve complex hydromecal-
lurgy. Such point discharges as might occur in milling processes
IV-2 7
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(i.e., the production of concentrate) are expected to con-
tain a variety of pollutants that need to be limited. Mining,
for the ores, is expected to lead to a smaller set of con-
taminants. While mining or milling of ores for uranium or
radium produces particularly noxious radioactive pollutants,
these are largely absent in an operation recovering vanadium
only. On the basis of these considerations, the SIC 1094
industry was tentatively subcategorized into: (1) The mining
of uranium/radium ores; (2) The processing of the ores of
the first subcategory to yield uranium concentrate and,
possibly, vanadium concentrate; (3) The mining of non-radio-
active vanadium ores; and (4) The processing of the ores of
the third subcategory to yield vanadium concentrate.
A careful distinction will be drawn between the radioactive
processes and the vanadium industry by including in the
former all operations within SIC 1094 that are licensed by
the U.S. Nuclear Regulatory Commission (NRC, formerly AEC,
Atomic Energy Commission) or by agreement states. The agree-
ment states, Including the uranium producing states of Colorado,
Texas, and Washington, have assumed all licensing, record-
keeping, and inspection responsibilities for radioactive
materials from the U.S. government upon establishing regula-
tions regarding radioactive materials that are at least as
stringent as those of the NRC(AEC). The licensing requirements,
as set forth in the code of Federal Regulations, Title 10,
Part 20 <10CFR20), already constitute restrictions on the
discharge of radioactive nuclides that form a minimum (larg-
est-discharge) standard superseding any less-stringent regu-
lations.
To further emphasize the distinction between the NRC-licensed
uranium subcategories and the pure vanadium subcategories,
the latter, whose products are used in the inorganic chemical
industry and, to a large extent, the ferroalloy smelting
Industry, are discussed further in connection with ferroalloy-
metal ore mining and dressing, in another portion of these
guidelines. The vanadium subcategories are summarized there
as members of the mining and hydrometallurgical process sub-
categories.
The variety of ores and milling processes discussed in Section
III might lead to the generation of as many subcategories
based on the major characteristics of the mill process as
there are ores and mills. It is possible, however, to group
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mills into fewer subcategories. This simplification is based
on the observations discussed below.
Raw wastewaters from mills using acid leaching remain acid
at the process discharge (not to be confused with a point
discharge), retain various heavy metals, and are generally
not suitable for recycling without additional and specialized
treatment. Those from the alkaline leach process are normally
recycled in part, since the leach process is somewhat selec-
tive for uranium and vanadium, and other metals remain in the
solid tailings. At one time, It was expected that mills
using solvent exchange would have a radically different raw-
waste character due to the discharge of organic compounds.
The fact that mills not using solvent exchange often process
ore that is rich In organics make this distinction less
important. As a result, a distinction must be made between
mills using acid leaching (or both acid and alkaline leaching)
of ore and mills using alkaline leaching of ore only.
While other differences between ores and processes, in addi-
tion to those mentioned above, can have an effect on wastewater
characteristics, they are not believed to Justify further
subcategorization. For example, there are some uranium/radium
ores that contain molybdenum and others that do not. Efflu-
ent limitations which may restrict molybdenum content must be
applied at all times and should not be restricted to those
operations which happen to run on ore containing molybdenum.
The two subcategories (acid and alkaline) retained reflect
not only differences in wastewater characteristics but also
(a) differences in the volume of wastewater that must be
stored and managed in a zero-effluent condition and (b) dif-
ferences in the ultimate disposition of wastes upon shutdown
of an operation.
Climatic conditions (such as rainfall versus evaporation
factors for a region), although subject to questions of
measurement, have an Important influence on the existence
of present-day point discharges and, thus, have been con-
sidered relative to present and future exploitation of uran-
ium reserves In the United States. Under the present defi-
nition of point discharge by the U.S. Environmental Protection
Agency (40 CFR125, para. K of the Federal Register), seepage
into the ground (and evaporation) is not counted as contri-
buting to a point discharge. With this factor taken into
consideration, and in view of the observation that economi-
cally exploitable uranium reserves are found in arid climates
IV-29
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for geological/chemical reasons, no point discharges are
needed to manage the raw wastewater from most current mining
and ore dressing operations in the uranium industry. In
addition, other operations that now discharge wastewater
plan to.terminate their discharges within a year or two. To
make possible a zero discharge from a sealed mine/mill pond,
the annual evaporation of water must exceed the annual pre-
cipitation of water. Since control of discharges to ground
water may become the goal of future state or federal legis-
lation, seepage should not in the future be counted on to
remove effluents.
There will be fluctuations in rainfall and evaporation from
year to year that may result in temporary accumulation of
excess water (or, conversely, in low levels) in wastewater
evaporating ponds. It is common engineering practice to
accommodate such fluctuations up to but not including the
abnormal "storm" that is observed only once every 25 years
or once every 100 years. In view of the useful life of mines
and mills, the integrated fluctuations short of those caused
by a 25-year storm should be stored in the holding ponds of
a zero-discharge operation without overflow. Beyond this
level, no fluctuations should result in progressive pond
failure, including at least those due to any documented
storm and also those due to the 100-year storm.
Ore characteristics were considered and, within a subcategory,
cause short-term effect on wastewater characteristics that
does not justify further subcategorization. Waste character-
istics were, as described above, considered extensively, and
it was found difficult to distinguish whether the acid/alkaline
leach distinction is based on process, mineralogy, waste
characteristics, or treatability of wastewater, since all
are interrelated. Vanadium operations which are not extracting
radioactive ore or covered under government licensing regula-
tions (NRC or agreement states) are subcategorized in the
ferroalloys section.
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Final Subcategorization of Uranium, Radium, and Vanadium
Category
The uranium, radium, and vanadium segment of the mining and
ore dressing industry considered here has been separated into
the following subcategories for the purpose of establishing
effluent guidelines and standards. These subcategories are
defined as:
I. Mines which extract (but not concentrate) ores
of uranium, radium, or vanadium under NRC
(formerly AEC) or Agreement State license.
II. Mills which process uranium, radium, or vanadium
ores to yield uranium concentrate and,
possibly, vanadium concentrate by either
acid or combined acid-and-alkaline leaching.
III. Mills which process uranium, radium, or vanadium
ores to yeild concentrates by alkaline
leaching only.
Problems Involved in Subcategorization. Those milling
operations surveyed which currently have a point discharge
either are changing the extraction process employed or
making plans to attain zero discharge. All of these opera-
tions are In arid areas. As uranium ore in the eastern U.S.
is exploited however, present methods of obtaining zero
discharge may have to be reexamined because of the humid
climate expected to be encountered.
IV-31
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Metal Ores. Not Elsewhere Classified
This group of metal ores was considered on a metal-by-metal
basis because of the wide diversity of mineralogies, processes
of extraction, etc. Most of the metal ores in this group
do not have high production figures and represent relatively
few operations. For this entire group, ore mineralogies and
type o* process formed the basis of subcategorization.
The metals ores examined under this category are ores of
antimony, beryllium, platinum, tin, titanium, rare earths
(including monazite), and zirconium.
Antimony Ores
Antimony mining and milling are practiced at two locations
in the United States. Although antimony is often found as
a byproduct of lead extraction, producers are often penalized
for antimony content at a smelter.
Final Antimony-Ore Subcategorization
The antimony ore mining and dressing Industry has been separated
into three subcategories for the purpose of establishing
effluent guidelines and standards. These subcategories are
defined as:
I. Mine(s) alone operating for the extraction of ores
to obtain primary or byproduct antimony ores.
II. Mill(s) or mine/mill complex(es) using a flotation
process for the primary or byproduct recovery
of antimony ore.
III. Mill(s) or mine/mill complex(es) obtaining antimony
as a byproduct of a base- or precious-metal
milling operation; antimony present in the
base- or precious-metal concentrate is recovered
at a smelter or refinery (antimony extraction
plant).
Beryllium Ores
Beryllium mining and milling in the United States are repre-
sented by one operating facility. Therefore, subcategorization
IV-32
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consists simply of division into mines and mills:
I. Mine(s) operated for the extraction of ores of
beryllium.
II. Mill(s) or mine/mill complex(es) using solvent
extraction (sulfuric-acid leach).
Platinum Ores
As discussed previously, most production of platinum in the
United States is as byproduct recovery of platinum at a
smelter or refinery from base- or other precious-metal con-
centrates. A single operating location mines and benefici-
ates ore by use of dredging, followed by gravity separation
methods. A single category, thus, is listed for platinum
ores:
I. Mine/mill complex(es) obtaining platinum concen-
trates by dredging, followed by gravity
separation and beneflclatlon.
Rare-Earth Ores
Rare-earth ores currently are obtained from two types of
mineralogies: bastnaesite and monazite. Monazlte is an
ore both of thorium and of rare-earth elements, such as
cerium. The subcategorization which follows is based pri-
marily upon division into mines and mills, as well as on
the type of processing employed for extraction of the rare-
earth elements.
I. Mine(s) operated for the extraction of primary
or byproduct ores of rare-earth elements.
II. Mill(s) or mine/mill complex(es) using flotation
process and/or leaching of the flotation
concentrate for the primary or byproduct
recovery of rare-earth minerals.
III. Mlll(s) or mine/mill complex(es) operated in con-
junction with dredging or hydraulic mining
methods; wet gravity methods are used in
conjunction with electrostatic and/or magnetic
methods for the recovery and concentration of
rare-earth minerals (usually, monazite).
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Tin Ores
A single operating location currently produces tin as a
byproduct of molybdenum mining and beneficlatlon. Placer
deposits of tin are found in the world and could be exploited
if discovered in the U.S. Therefore, a single subcategory
for miuing and one subcategory for milling are listed:
I. Mine(s) operating for the primary or byproduct
recovery of tin ores.
II. Mill(a) or mine/mill complex(es) using gravity
methods, flotation, magnetic separation
and/or electrostatic methods.
Titanium Ores
Titanium ores exploited in the United States occur in two
modes and mineralogical associations: as placer or heavy
sand deposits of rutile, llmenite, and leucoxene; and as
a titaniferous magnetite in a hard-rock deposit. The
titanium ore Industry, therefore, is subcategorized as:
I. Mine(s) obtaining titanium ore by lode mining
alone.
II. Mill(s) or mine/mill complex(es) using electro-
static and/or magnetic methods in conjunction
with gravity and/or flotation methods for
primary or byproduct recovery of titanium
minerals.
III. Mill(s) or mine/mill complex(es) In conjunction
with dredge mining operation; wet gravity
methods used In conjunction with electrostatic
and/or magnetic methods for the primary or
byproduct recovery of titanium minerals.
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Zirconium Ores
Zirconium ±s obtained from the mineral zircon in conjunction
with dredging operations. Mo additional subcategorizatlon
is required.
I. Mill(s) or mine/mill complex(es) operated in con-
junction with dredging operations. Wet gravity
methods are used in conjunction with electro-
static and/or magnetic methods for the primary
or byproduct recovery of zirconium minerals.
SUMMARY OF RECOMMENDED SUBCATEGORIZATION
Based upon the preceding discussion and choice of final
subcategorles, a summary of categories and subcategorles
recommended for the ore mining and dressing industry is
presented here In Table IV-1.
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TABLE IV-1. SUMMARY OF INDUSTRY SUBCATEGORIZATION RECOMMENDED
CATEGORY
IRON ORES
COPPER ORES
LEAD AND ZINC ORES
GOLD ORES
SILVER ORES
BAUXITE ORE
FERROALLOV ORES
MERCURY ORES
URANIUM. RADIUM
& VANADIUM ORES
Q ANTIMONY ORES
UJ
iL
5 BERYLLIUM ORES
U
u PLATINUM ORES
jjj RARE-EARTH ORES
IU
H
o
* TIN ORES
Ul
O
_j
< TITANIUM ORES
3
ZIRCONIUM ORES
SUBCATEGORIES
MINES
MILLS
MINES
Physical/Chemical Separation
Magnetic and Physical Separ anon (Mesabi Range)
Open-fit. Underground. Stripping
Hydrometellurgical (Laaehing)
Vat Leaching
Flotation
Process
Net Evaporation 2,76 2 cm I3O in I/year
Net Evaporation < 76 2 cm (30 in I/year
No Solubilmtion Potinlial
Solubiliiation Potantial
MILLS
MINES
MILLS
Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation
Byproduct of Base Maial Operation
MINES
MILLS
Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
Byproduct ol Base Metal Operation
MINES
MINES
MILLS
< 5.000 mi
itnc tons (5.512 short tonsl/year
> 5.000 metric tons/year by Physical Processes
> 5.000 m
Btric tons/year by Flotation
Leaching
MINES
MILLS
Gravity Separation
Flotation Process
Byproduct of Base/Precious-Metal Operation
MINES
MILLS
Acid or Acid/Alkaline Leaching
Alkaline Leaching
MINES
MILLS
Flotation Process
Byproduct ol Bese/Procious Metal Operation
MINES
MILLS
MINES OR MINE/MILLS
MINES
MILLS
Flotation 01 Leaching
Dredging or Hydraulic Methods
MINES
MILLS
MINES
MILLS
Electrostatic/Magnetic and Gravity/Flotation Procos&o
Physical Processes with Dredge Mining
MILLS OR MINE/MILLS
IV-36
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SECTION V
WASTE CHARACTERIZATION
INTRODUCTION
This section discusses the specific water uses in the ore
mining and dressing industry, as well as the amounts of
process waste materials contained in these waters. The
process wastes are characterized as raw waste loads emanat-
ing from specific processes used in the extraction of
materials involved in this study and are specified in terms
of kilograms per metric ton (and as pounds per short ton)
of product produced in ore processed. The specific water
uses and amounts are given" in terms of cubic meters (and
gallons) or liters per metric ton (and gallons per short
ton) of concentrate produced or ore mined. Many mining
operations are characterized by high water inflow and low
production, or by production rates that bear little resem-
blance to mine water effluent due to infiltration or precipita-
tion. Where this occurs, waste characteristics are expressed
in units of concentration (mg/1 = ppm). The discussion of the
necessity for reporting the data in this fashion in some
instances is discussed below under the heading "Mine Water."
The introductory portions of this section briefly discuss
the principal water uses found in all subcategories and
subcategories in the industry. A discussion of each mining
and milling subcategory, with the waste characteristics and
loads identified for each, concludes this section.
Because of widely varying wastewater characteristics, it was
necessary to accumulate data from the widest possible base.
Effluent data presented for each industry category were
derived from historical effluent data supplied by the indus-
try and various regulatory and research bodies, and from
current data for effluent samples collected and analyzed
during this study. The wastewater sampling program conducted
during this study had two purposes. First, it was designed
to confirm and supplement the existing data. In general,
only limited characterization of raw wastes has been pre-
viously undertaken by industry. Second, the scope of the
water-quality analysis was expanded to Include not only pre-
viously monitored parameters, but also waste parameters which
could be present in mine drainage or mill effluents.
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Mine Water
The effluent situation evident in the mining segment of the
ore mining and dressing industry is unlike that encountered
in most other industries. Usually, most industries (such as
the milling segment of this industry) utilize water in the
specific processes they employ. This water frequently becomes
contaminated during the process and must be treated prior to
discharge. In the mining industry, process water is present
only in placer operations operating by gravity methods, in
hydraulic mining, in washing operations, and in dust control.
Therefore, water is not normally utilized in the actual mining
of ores. It is an unwanted natural feature or problem that
presents problems to much of the mining industry. It enters
mines by ground-water infiltration and surface runoff and
becomes polluted by contact with materials in the host rock,
ore, and overburden. The polluted mine water then requires
treatment before it can be safely discharged into the surface
drainage network. These effluent quantities of ore mines
thus are unrelated, or only indirectly related, to production
quantities, except as noted above. Therefore, raw waste
loadings are expressed in terms of concentration rather than
units of production in many of the ore categories discussed
in Section IV.
In addition to handling and treating often massive volumes
of unwanted mine drainage during active mining operations,
metal ore mine operators are faced with the same problems
during startup, idle periods, and shutdown. Water handling
problems are generally minor during initial startup of a new
underground mining operation. These problems generally
Increase as the mine is expanded and developed and,
unless remedial action is taken, may continue after all min-
ing operations have ceased. The long-term drainage from
tailing disposal also presents long-term potential problems.
Surface mines, on the other hand, are somewhat more predict-
able and less permanent in their production of mine drainage
pollutants. Water handling within a surface mine is fairly
uniform throughout the life of the mine. It is highly depen-
dent upon precipitation patterns and precautionary methods
employed, such as the use of diversion ditches, burial of
toxic materials, and concurrent regrading and revegetation.
Because mine drainage with pollutants does not
necessarily cease with mine closure, a decision must be
V-2
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made as to the point at which a mine operator has fulfilled
his obligations and responsibilities for a particular mine
site. This point will be further discussed in Section VII,
"Control and Treatment Technology."
SPECIFIC WATER USES IN ALL CATEGORIES
Water is used in the ore mining and dressing industry for
ten principal uses falling under three major categories.
The principal water uses are:
(1) Noncontact cooling water
(2) Process water - wash water
transport water
scrubber water
process and product consumed
vater
(3) Miscellaneous water -
dust control
domestic/sanitary uses
washing and cleaning
drilling fluids
Noncontact cooling water is defined as that cooling water
which does not come into direct contact with any raw material,
intermediate product, byproduct, or product used In or
resulting from the process.
Process water is defined as that water which, during the bene-
ficiation process, comes into direct contact with any raw
material, intermediate product, byproduct, or product used
in or resulting from the process.
Noncontact Cooling Water
The largest use of noncontact cooling water in the ore mining
and dressing industry is for the cooling of equipment, such
as crusher bearings, pumps, and air compressors.
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Wash Water
Wash water comes into direct contact with either the raw
material, reactants, or products. An example of this type
of water usage is ore washing to remove fines. Waste efflu-
ents can arise from these washing sources because the resul-
tant solution or suspension may contain dissolved salts,
metals, or suspended solids.
Transport Water
Water is widely used in the ore mining and dressing industry
to transport ore to and between various process steps.
Water is often used to move crude ore from mine to mill,
to move ore from crushers to grinding mills, and to trans-
port tailings to final retention ponds.
Scrubber Water
Wet scrubbers are often used for air pollution control—
primarily, in association with grinding mills, crushers,
and screens.
Process and Product Consumed Water
Process water is primarily used in the ore mining and dress-
ing industry in wet screening, gravity separation processes
(tabling, jigging), heavy-media separation, flotation unit
processes (as carrier water), and leaching solutions; it is
also used as mining water for dredging and hydraulic mining.
Mine water is often pumped from a mine and discharged, but,
at many operations, mine water is used as part of processing
water at a nearby mill.
Miscellaneous Water
These water uses include dust control (primarily at crushers),
truck and vehicle washing, drilling fluids, floor washing and
cleanup, and domestic and sanitary uses. The resultant
streams are either not contaminated or only slightly contami-
nated with wastes. The general practice is to discharge such
streams without treatment or through leaching fields or septic
systems. Often, these streams are combined with process water
prior to treatment or discharged directly to tailing ponds.
Water used at crushers for dust control is usually of low
volume and is either evaporated or adsorbed on the ore.
V-4
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PROCESS WASTE CHARACTERISTICS BY ORE CATEGORY
Iron Ore
The nature and quantity of pollutants discharged in wastewater
from open-pit and underground iron mining operations and
beneficiation facilities vary from operation to operation.
In general, the quality of the water in mines is highly depen-
dent on the deposit mined and the substrata through which
the water flows prior to discharge.
Sources o|_ Waste. The main sources of waste in iron mining
and ore processing are:
(1) Wastewater from the mine itself. This may consist
of ground water which seeps into the mine, under-
ground aquifers intersected by the mine, or pre-
cipitation and runoff which enter from the surface.
(2) Process water, including spillage from thickeners,
lubricants, and flotation agents.
(3) Water used in the transport of tailings, slurries,
etc., which, because of the volume or impurities
involved, cannot be reused in processing or trans-
port without additional treatment.
In most casea, the last category constitutes the greatest
amount of waste.
Waste Loads and Variability. Waste loads from mines and
processing operations are often quite different, and there
is variability on a day-to-day and seasonal basis, both
within an operation and between operations. At times, mine
water is used as process feed water, and variability in its
quality Is reflected in the process water discharge.
Nature of_ Iron Mining Wastes* Mine wastewater can generally
be classified as a "clear water," even though it may contain
large amounts of suspended solids. The water may, however,
contain significant quantities of dissolved materials. If
the substrata are high in soluble material (such as iron,
manganese, chloride, sulfate, or carbonate), the water will
most likely be high in these components. Because rain water
V-5
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and ground water are usually slightly acidic, there will be
a tendency to dissolve metals unless carbonates or other
buffers are present.
Some turbidity may result from fine rock particles, generated
in blasting, crushing, loading, and hauling. This "rock
flour" will depend on the methods used in a particular mine
and on the nature of the ore.
Nitrogen-based blasting agents have been implicated as a
source of nitrogen in mine water. The occurrence of this
element (as ammonia, nitrite, or nitrate) would be expected
to be highly variable and its concentration a function of
both the residual blasting material and the volume of dilu-
tion water present.
These effluents in the iron mining operations are generally
unrelated to production quantities from the operation.
Therefore, waste loadings are expressed in concentration
rather than units of production. The principal contaminants
of mine water are:
(1) Suspended solids resulting from blasting, crushing,
and transporting ore; finely pulverized minerals
may be a constituent of these suspended solids.
(2) Oils and greases resulting from spills and leakages
from material handling equipment.
(3) Hardness and alkalinity associated with the host
rock or overburden.
(4) Natural levels of salts and nutrients in the intru-
sive water.
(5) Residual quantities of unburned or partially
burned explosives.
Processing Wastes. The processing of ore from the mine may
result in the presence of a number of waste materials in the
wastewater. Some of these are derived from the ore itself,
and others are added during processing. Still others are
not intentionally added but are inadvertent and inherent
contributions.
V-6
DRAFT
-------
DRAFT
Dissolved and suspended solids are contributed by the ore to
water used in transport and processing. Included in this are
metals. The nature and quantity of these are dependent on
the nature of the water, the ore, and the length of contact.
During processing, various flotation agents, acids, clays,
and other substances may be added and thereby become consti-
tue-'ts of wastewater. Oil and grease from machinery and
equipment may also contaminate the water.
Inadvertent additions include metals (such as zinc) from
buildings and machinery, runoff from the plant area and from
stockpiles which may contain dissolved and suspended solids,
and spills of various substances.
Sanitary sewage from employees and domestic sewage from wash-
rooms, lunchrooms, and other areas is usually disposed of
separately from process and transport wastes through munici-
pal or drainfield systems. Even when not, it would be
expected to constitute a minor part of the load.
The principal characteristics of the waste stream from the
mill operations are:
(1) Solid loadings of 10 to 50 percent (tailings).
(2) Unseparated minerals associated with the tailings.
(3) Fine particles of minerals (particularly, if the
thickener overflow is not recirculated).
(4) Excess flotation reagents which are not associated
with the iron concentrate.
(5) Any spills of reagents which occur in the mill.
One aspect of mill waste which has been poorly characterized
from an environmental-effect standpoint is the excess of flo-
ation reagents. Unfortunately, it is very difficult to detect
analytically the presence of these reagents—particularly, the
organics. COD, TOG, and surfactant tests may give some indi-
cation of the presence of organic reagents, but no definitive
information is related by these parameters.
V-7
DRAFT
-------
DRAFT
The substances present in mine-water discharges are given in
Table V-l; those present in process-water discharges are
given in Table V-2. These values are historically represen-
tative of what is present before and after discharge to the
receiving water. When mine water is used as processing water,
its characteristics often cannot be separated from those of
the processing water.
As part of this study, a number of mining and beneficiation
operations were visited and sampled. The results of the
sample analyses show certain potential problem areas with
respect to the discharge of pollutants. Summaries of the
major chemical parameters in raw wastes from mine and mill
water, measured as part of site visits, are given in Tables
V-3 and V-4. The basic waste characteristics, on the average,
are very similar for both mines and mills. Elevated concen-
trations of particular parameters tend to associate with a
particular mining area or ore body. For example, the dis-
solved iron and manganese tend to be much higher in Michigan
ores than in ores from the mining areas of the Mesabi Range
in Minnesota.
In the benef iciation of iron-containing minerals, as much
as 27.2 cubic meters of water per metric ton (7,300 gallons
per long ton) and as little as 3.4 cubic meters of water per
metric ton (900 gallons per long ton) of concentrate may be
used. The average amount of water per metric ton of ore
produced is approximately 11.8 cubic meters (3,200 gallons per
long ton) . Most processing water in beneficiation operations
is recycled to some extent. The amount of recycle is depen-
dent on the type of processing and the amount of water that
is included in the overall recycle system in the mill.
Mills that employ flotation techniques currently discharge
a percentage of their water to keep the concentration of
soluble salts from increasing to excessive levels. Soluble
salts—especially, those of the divalent ions—are deleterious
to the flotation process, causing excessive reagent use and
loss of recoverable iron. Even these operations currently
recycle at least 80 percent of their water.
Mills using physical methods of separation (magnetic, washing>
jigging, heavy media, spirals, and cyclones) can and do
recycle greater than 80 percent of their water. The amount
of water discharged from these operations is solely dependent
on how much water drains and accumulates into their impound-
ment systems.
V-8
DRAFT
-------
DRAFT
TABLE V-1. HISTORICAL CONSTITUENTS OF IRON-MINE DISCHARGES
PARAMETER
TSS
TDS
COO
PH
Oil and Grease
Al
Ca
Cr
Cu
Fe
Pb
Mg
Ho
Ni
Na
1UW
win
Zn
Chlonda
Cyanide
CONCENTRATION Img/W
BEFORE TREATMENT
MIN
1.000
140.0
0.200
5.00*
1.800
0.003
0.003
0.001
0.001
0.060
0.001
0.020
0.002
0.003
0.023
0.001
0.001
1.000
0.010
MAX
5.000.0
1380.0
36.0
8.40*
9.000
0.350
256.0
0.010
1.000
178.0
0.100
118.0
2.000
0.100
15.0
18.0
8.0
120.0
0.02
AVG
371.51
436.18
6.470
7.45"
4.511
0.066
85.39
0.007
0.167
13.3
0.018
39.35
1.001
0.024
7.511
2.462
1.869
27.143
0.013
NO.
19
17
10
IB
9
7
3
9
12
14
9
3
2
6
2
14
9
14
4
AFTER TREATMENT
MIN
1.000
100.0
0.026
6.800*
0.400
0.007
0.002
0.010
0.005
0.008
0.008
0.008
-
0.010
.
0.001
0.010
0.900
0.005
MAX
30.0
1,090.0
42.0
8.500*
20400
0.350
0.158
0.010
0.370
2.100
0.100
0.029
•
0.075
-
6.900
0.340
180.00
0.020
AVG
10.693
390.10
12.116
7.652*
4.313
0.131
0.045
0.010
0.120
0446
0.023
0.017
-
0.023
•
1.720
0.185
33.225
0.011
NO.
27
20
14
21
16
9
4
6
10
11
8
3
-
5
-
11
5
20
4
•Value in pH units
TABLE V-2. HISTORICAL CONSTITUENTS OF WASTEWATER FROM
IRON-ORE PROCESSING
PARAMETER
TSS
TDS
COO
PH
Oil and Grease
Al
Ca
Cu
Fe
Pb
Ni
Mn
Zn
Chloride
Cyanide
CONCENTRATION (mg/£)
BEFORE TREATMENT
MIN
1.20
0.500
0.200
5.000*
0.030
0.030
5S.O
-
0.200
0.100
0.010
0.007
0.006
1.000
MAX
9.999.0
356.0
36.0
8.300*
40400
5.000
250.0
-
10.0
5.0
0.050
20.0
10.0
110.0
AVG
1,894.8
207.1
16.986
7.187*
14.229
0.994
120.0
-
2.568
3.367
0.023
2.772
3.013
22.145
NO.
11
10
7
12
8
6
£
•
9
3
3
9
5
11
AFTER TREATMENT
MIN
0400
0.300
0.200
6.000*
0.100
0.009
82.0
0.010
0.050
0.045
0.010
0.016
0.010
0.350
0.008
MAX
200.0
1,090.0
90.0
8.300*
90.0
0.270
181.0
0.450
1.610
0.250
0.200
2.100
0.115
180.0
0.020
AVG
25.133
393.27
19.518
7.259*
12.0
0.107
131.5
0.230
0.453
0.111
0X187
0.529
0.056
42.875
0.013
NO.
15
16
12
16
13
8
2
2
10
4
3
10
4
16
4
'Value in pH units
V-9
DRAFT
-------
DRAFT
TABLE V 3. CHEMICAL COMPOSITIONS OF SAMPLED MINE WATERS
PAflAMETEH
pH
Alkalinity
COD
TSS
TOS
Conductinly
Total ft
Diuolrad F«
M»
Sulfate
CONCENTRATION (nq/ll IN EFFLUENT FROM MINE
o
£
73'
204
274
2
455
440'
004
<002
021
85
a
jj
i
72'
482
2
505
400*
<002
<002
<002
040
215
-
|
6
S
72'
176
45
30
246
310'
<002
<002
<002
45
_
|
6
£
8
:*•
10
<1
281
320<
<002
<002
COD2
28
„
ft
!
6
£
i
74'
183
21
169
215*
<002
0059
21
M
|
|
i
76'
211
228
20
271
140'
018
<002
<002
26
S
j
B
84*
218
18
6
1.302
1.950'
450
<002
320
152
r
6
1
I
S
71'
374
45
10
118
no'
2.80
<002
0025
11 2
1
0
3
72'
118
90
2
440
550'
130
004
0054
332
§
63-
181
27S
12
308
342'
030
002
065
387
|
!
2
o
7.9'
660
<10
48
1.290
1.125'
110
008
<002
780
1
!
I
i
i
754*
151
167
13.25
4995
5668'
086
0027
039
134
VtluB in pH unitt
f t/ilue m micromhas/cm
TABLE V-4. CHEMICAL COMPOSITIONS OF SAMPLED MILL WATERS
PARAMETER
PH
Alkalinity
COD
TSS
TDS
Conductance
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION <0.02
330.0
175
1103
Tailing-
Pond
Influent
7.5*
-
<1.0
7
198
350*
90.0
0.06
37.50
40
1105
Thickener
Tails
8.2*
-
9.2
2
287
375*
1180.0
0.10
320.0
55
1107
Underflow
From
Thickener
7.6*
-
225
10
712
0.70
<002
67.0
236
1108
Tails
Effluent
73*
-
13.5
-
230
130*
8.20
0.16
7.50
195
1109
Thickener
Underflow
9.5*
238
13.5
-
360
262*
004
004
16.0
20.7
1110
Thickener
Underflow
and Mine
Water
9.00*
13.4
11.9
16
2.360
1,900*
< 0.62
< 0.02
0.032
475
Average
(All
Mills)
8.2*
125.7
11.5
12.2
669
603* ;
212.8 ,
0.06
111.1
146
Value in pH units
Value in micromhos
V-10
DRAFT
-------
DRAFT
Typical mining operations take the water that accumulates
in the mine and pump it either to discharge or to a tailing
basin, where a portion is recycled" in the processing operat-
tion. Mine water is generally settled to remove suspended
matter prior to discharge or before use in plant processes.
A typical flow scheme for the treatment of mine water ±s
gixen in Figure V-l.
Process operations generally recycle high percentages of their
water. Water in the plant process is used to wash and trans-
port the ore through grinding processes. After separation
of the concentrate, the tailings are discharged to a tailing
pond, where the coarse and fine waste rock particles settle
(Figure V-2) . Clarified water is returned to be used in further
processing, and a portion is discharged to receiving waters.
Plants or mines that have zero discharge have not been dis-
cussed in this section because they discharge no waste
materials. It should be pointed out, however, that every
plant operation loses water to some degree and has to make
up this water loss to maintain a water balance. The main
sources of water loss are losses to within the concentrated
product, evaporation and percolation of water through
impoundment structures, loss of water to the tailings, and
evaporation or water loss during processing.
Process Descriptions
The following subsections discuss particular processing opera-
tions to demonstrate how water is utilized during different
ore processing, the water flow within each system, and the
waste loads generated.
Mine and Mill 1105. Mine and mill 1105 is a typical taconite
operation. Open-pit mines associated with the operation
produce an effluent, and the mill operates with a closed
water system.
Crude magnetic taconite is mined, mainly from the lower
cherty member of the Minnesota Biwabik formation, by con-
ventional open-pit methods and then milled to produce a
fine magnetite. The fine magnetite from the mill is
agglomerated in a grate-kiln system to produce approximately
V-ll
DRAFT
-------
DRAFT
Figure V-1. FLOW SCHEME FOR TREATMENT OF MINE WATER
MINE WATER
SEDIMENTATION BASIN
SETTLED
SOLIDS
TO
WASTE
CLARIFIED
EFFLUENT
TO RECEIVING
WATERS
TO PROCESS
WATER
Figure V-2. WATER FLOW SCHEME IN A TYPICAL MILLING OPERATION
WATER
TO
STOCKPILE
PROCESS
PRODUCT
PROCESS PLANT
COAGULANT
TO
WASTE
n
PROCESS
TAILING
RECYCLE (80-97%)
SETTLED CLARIFIED
SOLIDS EFFLUENT
TO RECEIVING
WATERS
V-12
DRAFT
-------
DRAFT
2.64 million metric Cons (2.6 million long tons) of oxide pellets
annually for blast-furnace feed.
The mine, mill, and pelletizing plant are located on a large site
controlled by the operating company, with 8094 hectares
(20,000 acres) utilized at present. An initial tailing pond
of 405 hectares (1000 acres) has been filled. A second 1,619-
hectare (4,000-acre) pond is now being used.
An open system is used in mine dewatering. A sketch of the
system with flow rates is shown in Figure V-3. Settling
basins are used to contain the water before it is discharged
to two lakes.
Chemical analysis of the mine discharge water after settling
shows the following chemical constituents:
Parameter Concentration (mg/1)
pH* 7.4
TSS 17
IDS 281
Iron (total) 0.10
Iron (dissolved) less than 0.02
Manganese less than 0.02
Sulfate 36
* value in pH units
The mill water system is a closed loop. Plant processes use
204 cubic meters per minute (54,000 gpm), with 189 cubic
meters per minute (50,000 gpm) returned from the 91.4-meter
(300-foot) diameter tailing thickener overflow and 15.1
cubic .meters per minute (4,000 gpm) returned from the tailing
pond or basin. The tailing thickener receives waste or tailings
in a slurry from the concentrate pellet plant. A nontoxic,
anionic polyacrylamide flocculant is added to the thickener
to assist in settling out solids. Tailing thickener under-
flow is pumped to the tailing basin.
Rotary machines are used in the mine to prepare blast holes
for the ammonium nitrate-fuel oil (ANFO) and metallized slurry
blasting agents. Electric shovels are used to load the
broken ore into 100-ton-capacity diesel/electric trucks for
haulage to the primary crusher.
V-13
DRAFT
-------
DRAFT
V 3. WAI tR BALANCE FOR MINE/MILL 1105 (SEPTEMBER 1974)
3 4 mj/min (900 gpntl
\
(INTERMITTENT)
>
3 SETTLING BASINS ^X
(IN SERIES) J
\
^ CREEK ^
i
Q LAK
2 PUMPS
C 11 4 m /mm
(3.000 gpm) EA
~~ s^llNTtRMI TTENT)
LL-^
p-
MINE OEWATERING
\
17 to 32 m3/mm (4.500 to 8.500 gpm)
MAX. 8.23 m3/mm (2.200 gpm)
(INTERMITTENT)
PLANT STORAGE TANK
i
15.1 m3/mm (4.000 gpm)
i
PLANT PROCESSES
1
204 m3/mm (54.000 gpm)
' 189 m3/min
TAILINGS THICKENER
\
^TAILIN
15 1 m3/mm
t
BPON^V^
(50.000 gpm)
4.000 gpm)
1 m3/mm (4.000 gpm)
\
L
{ 3 SETTLING BASINS^N
<^^ (IN SERIES) J
\
17 to 32 m3/mm
(4.600 to 8.500 gpm)
^" LAKE 2 ^
V-14
DRAFT
-------
DRAFT
The 1.52-meter (60-inch) primary crusher is housed in the
pit and reduces the ore to a size of less than 0.15 meter
(6 inches). From the crusher, coarse ore is conveyed to a
storage building.
Figure V-4 is a flowsheet showing the physical processing
used in the mill. Coarse ore assaying 22 percent magnetic
iron is reclaimed from the storage building and ground to
14-mesh size in the primary, air-swept dry grinding system.
Broken ore is removed from the mill by a heated air stream
and is air classified and screened. The coarse fraction goes
to a vertical classifier, and the fine fraction goes to two
cyclone classifiers. From the cyclone classifiers, the fine
product goes to a wet cobber to recover the magnetics for
the secondary grinding circuit. Coarse product of the
air classifiers is screened, and the oversize is returned
to the primary mill for further grinding. Undersize from
the classifiers is separated magnetically to produce a dry
cobber concentrate, a dry tailing, and a weakly magnetic
material which is recycled for further grinding and concen-
tration. About 37 percent of the crude weight is rejected
in the primary circuit.
Dust collected in sweeping the dry mill is pulped with water
and fed to a double-drum wet magnetic separator to produce a
final tailing and a wet concentrate for grinding in the
secondary mills.
Ball mills are used in the secondary wet grinding section to
reduce the size of the dry cobber and wet dust concentrates.
Slurry from the ball mills is sized in wet cyclones. Over-
size from the cyclones is returned to the ball mill. Under-
size ore from the cyclones is pumped to hydroseparators.
A rising current of water is used in the hydroseparator to
overflow a fine silica tailing. Hydroseparator underflow is
sent to finisher magnetic separators. The finisher separators
upgrade the hydroseparator underflow and produce a fine
tailing or discard. Finisher magnetic concentrate can be
further upgraded, if necessary, by fine screening and regrinding
and then reconcentrating the screen-oversize material.
The final concentrate is thickened and dewatered to about
10 percent moisture prior to the formation of "green balls'
from this material. A bentonite binder is blended with the
concentrate before balling in drums. The balling drums are
V-15
DRAFT
-------
DRAFT
Figure V-4. CONCENTRATOR FLOWSHEET FOR MILL 1105
FEED
DRY SEMIAUTOGENOUS GRINDING MILLS
I
VERTICAL DRY CLASSIFIER
OVERSIZE
UNDERSIZE
I
CYCLONE CLASSIFIER
OVERSIZE
UNDERSIZE
t
SCREEN
WET MAGNETIC COBBING
OVERSIZE UNDERSIZE
I t
I
CONCENTRATE
I
TAIL
DRY MAGNETIC ROUGHER
TAIL
i
CONCENTRATE
I
DRY MAGNETIC SCREENING
MIDDLING
I
WET SECONDARY
GRINDING MILLS
HYDROCYCLONES
UNDERSIZE OVERSIZE
TAIL
i
HYDROSEPARATION
CONCENTRATE
i
i
TAIL
I
WET FINISHER MAGNETIC SEPARATION
I
CONCENTRATE
TAIL
J
TO
PELLET
PLANT
TAILING THICKENER
UNDERFLOW
I
OVERFLOW
TO
WASTE
TO
TAILING POND
TO
REUSE WATER
V-16
DRAFT
-------
DRAFT
in closed circuit with screens to return undersize material
to the drum and to control the green ball size.
Fines are again removed from the green balls on a roller
feeder before they enter a traveling grate. These fines are
recirculated to a balling drum or to the pellet plant feed.
Green balls are dried in an updraft and downdraft section
of the grate. Dried balls then pass through a preheat section
on the grate. The magnetite begins to oxidize, and the
balls to strengthen, while passing through the preheat
section.
Balls go directly from the grate to a kiln, where they are
baked at 1315 degrees Celsius (2400 degrees Fahrenheit)
before they are discharged to a cooler, where oxidation of
the pellets is completed and pellet temperature is reduced.
The finished pellets contain 67 percent iron and 5 percent
silica and are transported for lake shipment to the steel
industry.
Mine and Mill 1104. This mine/mill complex is a typical
natural ore (an iron ore that contains moisture) operation,
with the mine and mill both producing effluents. Physical
processes are used in the mill to remove waste material from
the iron. The plant processes a hematite/limonite/goethite
ore and was placed in operation at the start of the 1962
shipping season. The operation is seasonal for 175 days
per year, from the last week in April to about the middle
of October.
Mine water from one of the two active pits is pumped to an
abandoned mine (settling basin) and overflows to a river at
an average rate of 7,086 cubic meters per day (1,872,000 gpd)
and at a maximum rate of 5,826 cubic meters per day (1,539,000
gpd) per day at Discharge No 1. Mill process water, mine
drainage from the other pit, and fine tailings from the mill
are pumped to a 105-hectare (260-acre) tailing basin. Process
water is recycled from the basin at a rate of 45 cubic meters
(]2,000 gallons) per minute. Excess water from the tailing
basin is siphoned to a lake intermittently at an average
rate of 3,717 cubic meters (981,900 gallons) per day at
Discharge No. 2. Table V-5 is a compilation of the chemical
characteristics and waste loads present in mine water
(Discharge No. 1—concentration only) and combined mine and
mill process effluent.
V-17
DRAFT
-------
DRAFT
TABLE V-5. CHEMICAL ANALYSIS OF DISCHARGE 1 (MINE WATER) AND
DISCHARGE 2 (MINE AND MILL WATER) AT MINE/MILL 1104,
INCLUDING WASTE LOADING FOR DISCHARGE 2
PARAMETER
PH
TSS
TDS
Total Fe
Dissolved Fe
Mn
CONCENTRATION (mg/l ) IN WASTEWATER
DISCHARGE 1
6.7 •
6
263
<0.02
<0.02
<0.02
DISCHARGE 2
7.3»
6
210
<0.02
<0.02
<0.02
RAW WASTE LOAD
g/metric ton
—
3.8
132
< 0.01 3
< 0.01 3
<0.013
Ib/short ton
—
0.0074
0.26
<0.00003
<0.00003
<0.00003
•Value in pH units
V-18
DRAFT
-------
DRAFT
Mining is carried out by conventional open-pLr methods.
Ammonium nitrate explosives are used in bLat.t-i.ng. Shovels
load the ore into trucks for transport to the plant.
At the mill, the ore, averaging 37 percent iron, is fed to a
preparation section for screening, crushing, and scrubbing.
A plant flowsheet is shown in Figure V-5.
Reversible conveyors permit rock coarser than 10.2 centi-
meters (4 inches) from the first stage of screening to be
removed as a reject and stockpiled or processed further
depending on the quality of the oversize material. Plant
feed is processed in a crusher/screen circuit to produce
fractions which are 3.2 cm by 0.64 cm (1.25 inches by 0.25 inch)
and less than 0.64 cm (0.25 inch). The material which is
3.2 cm by 0.64 cm (1.25 inches by 0.25 inch) goes to a heavy-
media surge pile. The fraction which is less than 0.64 cm
(0.25 inch) after classification to remove tailings which
are less than 48 mesh is sent to a jig surge pile.
Material from the heavy-media surge pile is split into
fractions which are 3.2 cm by 1.6 cm (1.25 inches x 0.63 inch)
and 1.6 cm x 0.64 cm (0.63 inch by 0.25 inch). Both fractions
go to identical sink/float treatment in a ferrosilicon
suspension. Float rejects or tailings from the heavy sus-
pension treatment are trucked to a stockpile. Concentrates
go directly to a railroad loading picket. The ferrosilicon
medium is recovered by magnetic separation. The magnetic
medium is recycled to the process. Nonmagnetic slimes go
to the tailing pond. The material which is less than 0.64
cm (0.25 inch) but greater than 48 mesh goes from the surge
pile to jigs, where pulsating water is used to separate the
concentrate and tailing. Concentrates are dewatered before
shipment, and water from this operation is recycled in the
plant. Jig tailings are sent to a dewatering classifier.
Sands from the classifier are trucked to a reject pile.
Overflow from the classifier is pumped to the tailing
basin.
Concentrates produced in the plant are shipped by rail and
boat to the lower lakes. The 58-percent-iron heavy-media
concentrate serves as blast-furnace feed. The 58-percent-
iron jig concentrate is later sintered at the steel plant
before entering the blast furnace.
V-19
DRAFT
-------
DRAFT
Figure V-5. FLOWSHEET FOR MILL 1104 (HEAVY-MEDIA PLANT)
I MINING I
I
CRUDE ORE
(J7» IRON)
*
DOUBLE DECK
SCREEN
--
OOUBLE DECK
SCREEN
> 16 2 cm
06 m)
I
10 2 la 16 2
(4lo6in)
*•* f* 42% Ft
CONE
CRUSHER
TO HOCK REJECTS
STOCKPILE
>064«n
OOJSin)
< 064 em
k026m)
1
HEAVY MEDIA
SURGE PILE
0 6« to 3 2 cm 10 25 u IX in )-
<064cm
KOJSml
FLOAT REJECTS
\
TO FLOAT REJECTS
STOCKPILE
TO
TAILING
POND
RECYCLE TO
HEAVY MEDIA
SEPARATORS
CONCENTRATES TO TRANSPORTATION
I- 32% OF CRUDE ORE)
V-20
DRAFT
-------
DRAFT
Mine and Mill 1108. This mine/mill complex is located in
Northern Michigan. The ore body consists of hematite (major
economic material), magnetite, martite, quartz, jasper, iron
silicates, and minor secondary carbonates. All of the consti-
tuents appear in the tailing deposit. The concentration plant
processes approximately 21,000 metric tons (20,700 long tons)
per day of low-grade hematite at 35.5 percent iron to produce
approximately 9,850 metric tons (9,700 long tons) per day of
concentrated ore at 65.5 percent iron. The remaining 11,200
metric tons (12,346 short tons), at approximately 10 percent
total iron, are discharged to the tailing basin.
Mine water is currently pumped from the actively mined pit
and discharged directly. The chemical constituents of the
discharged water are given in Table V-6.
Water in the concentration process is utilized at a rate of
114 cubic meters (30,000 gallons) per minute. Ore is first
ground to a fine state (80 percent less than 325 mesh) and
the argillaceous slime materials removed by wet cycloning.
A simplified flow scheme is included in Figure V-6. Subse-
quently, the concentrated ore is floated using tall oil -
fatty acid. The flotation underflows are discharged to a
tailing stream, which is discharged directly to a 385-hectare
(950-acre) tailing basin. Approximately 80 percent of the
water from the tailing pond is returned to the concentrating
plant as reuse water (untreated) . The remaining 20 percent
is discharged, after treatment, to a local creek. This dis-
charged wastewater is first treated with alum, then with a
long-chain polymer to promote flocculation. It then passes
to a 8.5-hectare (21-acre) pond, where the flocculated particles
settle. The concentration of chemical parameters and the
waste loading in this discharge are given in Table V-7.
Copper Ore
Frequently, discharged wastes encountered in the copper ore
mining and dressing industry include waste streams from
mining, leaching, and milling processes. These waste
streams are often combined for use as process water or
treated together for discharge. Other wastes encountered
in this segment are discharge wastes from copper smelting
and refining facilities, treated sewage effluent, storm
drains, and filter backwash. The uses of water in copper
mining and milling are summarized below.
V-21
DRAFT
-------
DRAFT
TABLE V-6. CHEMICAL CHARACTERISTICS OF DISCHARGE WATER
FROM MINE 1108
PARAMETER
PH
Alkalinity
COD
TSS
TDS
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION
(mg/£ )
7.2*
118
9.0
2
440
1.3
0.04
0.054
33.2
•Value in pH units
V-22
DRAFT
-------
DRAFT
Figure V-6. SIMPLIFIED CONCENTRATION FLOWSHEET FOR MINE/MILL 1108
MINING
T
CRUDE ORE
24.500 metric tont
(20.700 long tons)
per day
REGRIIMOING.
FLOTATION. THICKENING.
AND FILTRATION
I
SECONDARY
CONCENTRATE
i
PEL..ETIZING
OPERATION
PELLETS
t
9 3 m3/min (5 5 cfs)
17 m3/min (10 cfi)
41 6 m3/min (24 5 cfi)
14 4 m3/rnin (3 5 cfs)
17 m-Vrnin (10 cfs)
0 85 m3/min (0 5 cfs)
TO STOCKPILE
8 1% SOLIDS
100 m3/mm (59 cfs)
Y
TO TAILINGS
V-23
DRAFT
-------
DRAFT
TABLE V-7. CHARACTERISTICS OF MILL 1108 DISCHARGE WATER
PARAMETER
pH
AllMlmlly
COD
TSS
TDS
TotilF*
DaaohwIF*
Mi
SulUM
PROG RAM SAMPLE
CONCENTRATION
Ima/D IN
WASTE WATER
7.1*
82.0
22.6
10
16O
2.06
O93
005
5
WASTE LOAD
in g/nwlric ton
(Ib/lhort ton)
PRODUCT
-
213 (O42I
77 « (01SI
3.4 100071
660 (108)
7 OS (0 013)
32 (0008)
0 17 (0 0003)
172 (0.034)
10MONTH AVERAGES
AVERAGE
CONCENTRATION
Iml/tl
70-
-
-
86
-
-
078
066
—
WASTE LOAD
in o/iMtric ton
lib/Ikon ton)
PRODUCT
_
-
-
20.7 10040)
-
-
183(00036)
1 66(00031)
-
HIGH
CONCENTRATION
lino/I)
79-
-
-
S3
-
-
360
680
-
LOW
CONCENTRATION
Img/ll
66*
-
-
1
-
-
001
001
-
•WlminpHunili
V-24
DRAFT
-------
DRAFT
I. Mining:
a. Cooling
b. Dust control
c. Truck washing
d. Sanitary facilities
e. Drilling
II. Hydrometallurgical processes associated with
mining: Dump, heap, and in situ leaching
solutions.
III. Milling Processes:
a. Vat leach
1. crusher dust control
2. Vat leach solution
3. Wash solutions
b. Flotation
1. Crusher dust control
2. Carrier water for flotation
Copper Ore Mining. Most of the domestic copper is mined in
low-grade ore bodies in the western United States. All mining
and milling activities adjust to the type of copper minerali-
zation which is encountered. The principal minerals exploited
may be grouped as oxides or sulfides and are listed in Table
V-8. Porphyry copper deposits account for 90 percent of the
domestic copper ore production and are mined by either block-
caving or open-pit methods. The choice of method is deter-
mined by the size, configuration, and depth of the ore body.
Open-pit (undercut) mining accounted for 83 percent of the copper
produced in the United States in 1968. The mining sequence
includes drilling, blasting, loading, and transportation.
Primary drilling involves sinking vertical or near-vertical
blast holes behind the face of an unbroken bank. Secondary
drilling is required to break boulders too large for shovels
to handle, or to blast unbroken points of rock that project
above the digging grade in the shovel pit. Ore and over-
burden are loaded by revolving power shovels and hauled by
large trucks (75 to 175 ton capacity) or by train. Ore and
waste may be moved by tractor-drawn scrapers or belt conveyors.
Some mines have primary crushers installed in the pit which
send crushed and semi-sorted material by conveyor to the
mill.
V-25
DRAFT
-------
DRAFT
TABLE V-8. PRINCIPAL COPPER MINERALS USED
IN THE UNITED STATES
MINERAL
Chalcocite
Chalcopyrite
Bornito
Covell ite
Enargite
COMPOSITION
OCCURRENCE*
SULFIDES
Cu2S
CuFeS2
Cu5FeS4
CuS
Cu3A$S4
SW.NW. ••
SW. NW. ••
NW,SW
NW,SW
NW
OXIDES
Chryiocolla
Malachite
Azurite
Cuprite
Tenorite
CuSiO3-H2O
Cu2(OH)2-CO3
Cu3(OH)2-(C03)2
Cu2O
CuO
SW"
SW, NW
SW, NW"
SW
SW
NATIVE ELEMENTS
Copper 11 Cu
NC.SW"
*SW = Southwest U.S.
NW = Northwest U.S.
NC = Northcentral U.S.
"Major minerals
V-26
DRAFT
-------
DRAFT
In 1968, 445 million metric tons (490 million short tons)
of waste material were discarded (mostly from open-pit
operations) after production of 154 million metric tons
(170 million short tons) of copper ore. The cutoff grade
of ore, which designates it as waste, is usually less than
0.4 percent copper. However, oxide mineralization of 0.1
to 0.4 percent copper in waste is separated and placed in
special dump areas for leaching of copper by means of sulfuric
Underground mining methods provided 17 percent of the U.S.
copper in 1968. Deep deposits have been mined by either
caving or supported stopes. Caving methods include block
caving and sublevel caving. For supported stope mining,
installation of systematic ground supports is a necessary
part of the mining cycle. In underground mining, solid
waste may be left behind. More than 60 percent of the
material produced is discarded as too low in copper content
or as oxide ore, which does not concentrate economically
by flotation.
Water Sources and Usage. In the mining of copper ores,
water collected from the mines may originate from subsurface
drainage or seepage from surface runoff, or from water pumped
to the mine when its own resources are insufficient. A
minimal amount of water in mining is needed for cooling,
drilling, dust control, truck washing, and/or sanitary
facilities (Figure V-7). For safety, excess mine water not
consumed by evaporation and seepage must be pumped from the
mines. Table V-9 lists the amount of mine water pumped from
selected mines and the ultimate fate of this wastewater at
surveyed mines. Open-pit mines pumped 0 to 0.27 cubic meter
per metric ton (0 to 64.7 gallons per short ton) of ore produced,
while underground mines pumped 0.008 to 3.636 cubic meters
per metric ton (1.91 to 871 gallons per short ton) of ore
produced.
Solid wastes produced are summarized in Table V-10 as metric
tons (or short tons) of waste (actually, overburden and
wastes) per metric ton (short ton) of ore produced. Under-
ground operations rarely have waste. Those mines which do
produce wastes yield relatively small amounts in comparison
to open-pit mining operations.
V-27
DRAFT
-------
DRAFT
Figure V-7. WASTEWATER FLOWSHEET FOR PLANT 2120-B PIT
NATURAL DRAINAGE.
SEEPAGE. AND
RUNOFF
i
MINING
(DRILLING,
BLASTING,
AND
LOADING)
ORE
16,560,000 metric tons/year
(18.250,000 short tons/year)
TO
MILL
EXCESS
MINEWATER
o
0.06 m /metric ton
(14.4 gal/short ton)
c
POND
LIME PRECIPITATION
0.06 m3/metric ton
(14.4 gal/short ton)
DISCHARGED
V-28
DRAFT
-------
DRAFT
TABLE V-9. MINE-WATER PRODUCTION FROM SELECTED MAJOR COPPER-PRODUCING
MINES AND FATE(S) OF EFFLUENT
MINE
2101
2102
2103
2104
2107
2108
2109
2110
211.1
2113
2114
2115
2116
2117
2118
2119
2120
2121
2122
2123
2124
TYPE*
OP
UG
OP
OP
UG
OP
OP
OP
OP
OP
OP
UG
OP
UG
OP
UG
UG.OP
UG
OP
OP
OP
MINE-WATER PRODUCTION
m3/metric ton
ore produced
0270
0.008
NE.
0.086
N/A
N.E.
N.E
N.E
N.E
0.015
40 5 (avg|T
1769
0.030
0886
0014
0.654
0.486
0170
0034
0075
N.E.
gal/short ton
ore produced
64.7
1 85
N.E
20.6
N/A
N.E.
NE
N.E.
NE
35
9.715 0
-------
DRAFT
TABLE V-10. SUMMARY OF SOLID WASTES PRODUCED BY PLANTS SURVEYED
MILL
MILL
2101
2102
2103
2104
2107
2108
2109
2110
2111
2112
2113
2114
2115
2116
2117
2118
2119
2120
2121
2122
2123
2124
HAULED WASTE (1973)
metric tons
38.32 1.250*
21.532.400>
57.213,000
22,129,279'
0 (UG)
12,566.400*
26.700.000
115.000
9.420.000
50.000 (UG)
19.773.594
12,000,000*
20.000 (UG)
35,557.000*
0(UG)
37.063.000*
91.200IUG)
36.500.000*
0(UG)
97.500.000*
12.000.000*
16.909.000
short tons
42,241,897
23.735.379
63.066.462
24.393,325
0
13352,068
29.431.677
126,766
10.383.760
56.116
21.796.630
13.227.720
22^)46
39,194,836
0
40354.915
100,531
40.234,315
0
107,475,220
13.227.720
18,638,959
MILL ORE
metric tons
7.934.320
8.782.600
15,407,000
8,10.1,784
N/A
3,927,000
1,727,800
4.092.000
1.632,000
700,000
10,343,337
143.723
519,593
12.638,000
1.335.626
18,360.000
21.974.500
25,730,000
8384,136
38,300,000
2.172,000
8.722,000
(1973)
short tons
8,746.080
9.681.148
16,983,290
8330.678
N/A
4,328,771
1304371
4310.663
1.798370
771,617
11,401363
158,427
572.753
13330393
1.472.274
20,238.411
24.222.111
28.362.436
9.793.072
42.218.473
2.394.217
9.614.348
RATIO
(WASTE/ORE)
4.83
2.45
3.71
2.73
—
3.20
15.45
0.03
5.77
0.07
1.91
83.5T
0.04
2.81
—
2.02
0.004
1.42
—
2.55
5.53
1.94
• All or a portion leached
* Stripping operation
N/A = Not available
UG - Underground
V-30
DRAFT
-------
DRAFT
Air quality control within open-pit nines consists of spraying
water on roads for dust control. Underground mines may employ
scrubbers, which produce a sludge of particulates. The
sludge is commonly evaporated or settled in holding ponds.
Wastewater Characterization. The volume of mine water pumped
from mines was previously summarized in Table V-9. The
chemical characteristics of these waters are summarized in
Table V-ll, which includes the flow per day, concentration
of constituents, and raw-waste load per day.
A portion of the copper industry (less than 5 percent) must
contend with acid mine water produced by the percolation of
natural water through copper sulfide mineralization associated
with deposits of pyrite (FeS2_). This results in acid water
containing high concentrations of iron sulfate. Acid iron
sulfate oxidizes metal sulfides to release unusually high
concentrations of trace elements in the mine water. The pH
of mine water most often is in the range of 4.0 to 8.5. In
the southwestern U.S., mine water is obtained from under-
ground shafts, either in use or abandoned on the property.
This source of water is valuable and is used for other
copper-producing processes. In contrast, mine water in Utah,
Montana, Colorado, Idaho, Oklahoma, Michigan, Maine, and
Tennessee—especially, in underground mines—is often unwanted
excess, which must be disposed of if reuse in other processes
(such as leaching and flotation) is not possible.
The primary chemical characteristics of mine waters are:
(1) occasional presence of pH of 2.0 to 9.5; (2) high dissolved
solids; (3) oils and greases; and (4) dissolved
metals. Often, mine water is characterized by high sulfate
content, which may be the result of sulfide-ore oxidation
or of gypsum deposits. Mine water—particularly, acid mine
water—may cause the dissolution of metals such as aluminum,
cadmium, copper, iron, nickel, zinc, and cobalt. Selenium,
lead, strontium, titanium, and manganese appear to be
indicators of local mineralogy and are not solubilized
additionally by acid mine water.
Handling of Mine Water. As shown in Table V-9, mine
waters are pumped to leach and mill operations as a water
source for those processes whenever possible. However, four
V-31
DRAFT
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 1 of 4)
O
30 <
> I
•n W
PARAMETER
Flow
pH
IDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
S«
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
MINE 2119
CONCENTRATION
(mg/ ;i
42.0135m3/day
964*
544
8
1
5
<10
02
OS
< 005
< 0003
< 050
< 007
< 005
< 02
380
< 005
< 005
< 010
< 02
013
00008
< 005
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric Ions
752 m3/1000 metric tons
964'
4185
62
077
385
<769
0 154
0385
<0038
<0002
<0385
<0054
<0038
< 0 154
2923
< 00385
< 00385
<0077
<0 154
0 10
000062
<0038
lb/1000 short tons
180.332 gal/1000 short tons
964'
8370
124
1 54
770
<1538
0308
0770
< 0076
< 0004
< 0770
< 0108
< 0076
< 0308
5846
< 00770
< 00770
< 0154
< 0308
020
0 00124
< 0076
MINE 2120-K
CONCENTRATION
(mg/5.1
27,524 Sm3/day
349*
4.590
4
<1.0
31
20
010
920
032
N/A
< 002
<007
1720
< 05
2.0000
100
033
024
< 05
1 35
00784
< 01
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
15.173 m3/1000 metric tons
349'
69.630.3
607
<1517
713
3034
1 52
1.3956
485
N/A
< 303
< 1 06
2.6092
< 759
30.340
1.517
501
364
< 759
2048
1 19
< 152
lb/1000 short tons
3.635.997 gal/1000 short tons
3.49*
139.2606
1214
< 3034
1426
606.8
304
2.791 2
97
N/A
< 606
< 212
5.2184
< 1517
60.680
3.034
100?
728
< 15 17
4096
2.38
< 304
O
33
>
•n
•Value in pH units
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 2 of 4)
O
3) <-
> f
•Tl CO
PARAMETER
Flow
PH
TOS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
Ai
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
MINE 2120-B
CONCENTRATION
< 1 0
32
<10
004
530
0 1
0007
<02
< 007
31 25
< 05
600
265
1 3
013
<05
1 55
00005
<01
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric Ions
60 08 m3/1000 metric tons
61'
1293
24
<0060
0 192
<0601
0002
0318
0006
00004
<0012
<0004
1 88
<003
0361
1 592
0781
0008
<003
0093
000003
<0006
lb/1000 short tons
14.400 gal/1000 short tons
61'
2586
48
<012
0384
< 1 202
0004
0636
0012
00008
< 0024
<0008
376
< 006
0722
3184
1 562
0016
<006
0186
000006
<0012
MINE 2120 -CE
CONCENTRATION
Img/i)
272S2m3/day
47-
454
34
170
23
<10
001
62
006
0042
< 02
< 007
617
< 05
86
1 42
0034
< 005
< 05
009
00005
< 01
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
17 685 m3/1000 metric tons
47'
803
060
030
0041
<0 177
00002
Oil
00011
000074
< 00035
< 00012
0109
< 0009
0 152
0025
00006
< 00009
< 0009
0002
0000009
< 0002
lb/1000 short tons
4.239 gal/1000 short tons
47'
1606
1 2
06
0082
< 0354
00004
022
00022
000148
< 0007
< 00024
0218
< 0018
0304
005
00012
< 00018
< 0018
0004
0000018
< 0004
O
3)
'Value in pH units
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 3 of 4)
D
3D
f
to
PARAMETER
Flow
pH
TOS
TSS
Oil and Greaie
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Si
HE
Pb
MINE 2121
CONCENTRATION
Img/Zl
3.8153m3/day
737'
29.250
69
<10
<45
819
219
087
<004
<0077
060
<007
28
<05
<01
222
<002
<005
<05
119
< 0 0001
<01
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric Ions
17 28 m /1 000 metric tons
737-
5.053 9
11 9
<0173
<0778
141 5
0378
0150
<0007
<0013
0104
<0012
0484
<0086
<0017
0384
<0003
<0009
<0086
206
< 0 00002
<0017
Ib/ 1000 short tons
4.141 gal/1000 short tons
737-
10,107 8
238
<0346
< 1 556
283
0756
03
<0014
<0026
0208
<0024
0968
<0 172
<0034
0768
<0006
<0018
<0172
41.2
< 000004
<0034
MINE 2122
CONCENTRATION
(mg/2)
3.274m3/day
761'
2.288
2
3
21
389
Oil
1.90
190
<0003
0.2
<007
133
<02
95
083
<005
013
<02
0.83
< 00001
<05
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
34 m3/1000 metric tons
7.6V
7869"
0069
0103
0722
134
0004
0065
0065
< 00001
0007
< 0002
0046
< 0007
0327
0029
< 0002
0004
< 0007
0.029
< 0 000003
< 0017
lb/1000 short tons
8.053 gal/1000 short tons
76V
15738
0 138
0206
1 444
268
0.008
0130
0130
< 00002
0014
<0004
0092
<0014
0654
0058
<0004
0.008
<0014
0.058
< 0000006
<0034
o
a
•Value m pH unit*
-------
TABLE V-11. RAVV WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 4 of 4)
1
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
a
Cu
Co
Se
Te
fa
Zn
So
Fe
Mn
Cd
Mi
Mo
Sr
Hg
Pb
MINE 2123
CONCENTRATION
< 006
0096
< 02
< 001
01
< 05
< 0 1
09
< 003
< 005
< 02
08
< 00001
< 05
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric Ions
75 m^/1000 metric Ions
696*
101
02
05
075
03
0005
008
< 0005
0007
< 002
< 00008
0008
< 004
< 0008
007
< 0002
< 0004
< 002
006
< 0000008
< 004
lb/1000 short Ions
18,000 gal/1000 short Ions
696-
202
04
1 0
1 5
06
001
0 16
< 001
0014
< 004
<: 00016
0016
< 008
< 0016
0 14
< 0004
<: 0008
< 004
012
<: 0000016
< 008
o
30
J>
• Value in pH units
-------
DRAFT
of the plants surveyed discharge all of their mine water
to surface waters. Half of these treat the water first by
lime precipitation and settling.
Process Description-Hydrometallurgical Extraction Processes
(Mining)
The use of acid leaching processes on low-grade oxide ores
and wastes produces a significant amount of cement copper each
year. All leaching is performed west of the Rocky Mountains.
Figure V-8 is a flow diagram of the process of acid leaching.
Leaching of oxide mineralization with dilute sulfuric acid
or acid ferric sulfate may be applied to four situations
of ore. Dump leaching extracts copper from low-grade (0.1
to 0.4 percent Cu) waste material derived from open-pit mining.
The cycle of dissolution of oxide mineralization covers many
years.
Most leach dumps are deposited upon existing topography. The
location of the dumps is selected to assure impermeable sur-
faces and to utilize the natural slope of ridges and valleys
for the recovery and collection of pregnant liquors. In
some cases, dumps have been placed on specially prepared
surfaces. The leach material is generally less than 0.61
meter (2 feet) in diameter, with many finer particles.
However, it may include large boulders. Billions of tons
of material are placed in dumps that are shaped as truncated
cones.
The leach solution is recycled from the precipitation or other
recovery operation, along with makeup water and sulfuric
acid additions (to pH 1.5 to 3.0). It is pumped to the top
of dumps and delivered by sprays, flooding, or vertical pipes.
Factors such as climate, surface area, dump height, mineralogy,
scale of operation, and size of leach material affect the
choice of delivery method. Figure V-9 summarizes the reactions
by which copper minerals are dissolved in leaching.
V-36
DRAFT
-------
DRAFT
Figure V-8. FLOWSHEET OF HYDROMETALLURGICAL PROCESSES USED IN
ACID LEACHING AT MINE 2122
TO ATMOSPHERE
TO ATMOSPHERE
43 m3/RMlnc ton
110.282 oriMiort ton)
241 m3/nwtnc ton
(57.834 pl/tfiort ton)
EVAPORATION
EVAPORATION
27 m3/mitnc ton
(6.426 gri/lhort ton)
-SEEPAGE —
296 m3/nutnc ton
I70.B79 gal/dion ton)
2.294 m3/nwtnc ion
(549.873 pl/ihort ton)
360 m3/nwtnc Ion
(86,303 gri/inort ton)
107 n>3/m«nc Ion
<25.704 gd/ihort ton)
2.386 m3/Rwtnc Ion
(571.914 gri/riiort ton)
16 m3/m«tnc ton
(3.727 ojlhhon ton)
OJ m3/metnc ton
(64 gil/ihort ton)
V-37
DRAFT
-------
DRAFT
Figure V-9. REACTIONS BY WHICH COPPER MINERALS ARE DISSOLVED IN
DUMP. HEAP, OR IN-SITU LEACHING
AZURITE
Cu3(OH)2-(CO3)2 + 3H2SO4 ~^ 3CuSO4 + 2CO2
MALACHITE
Cu2(OH)2-CO3 + 2H2SO4 ~—^ 2CuSO4 + C02 + 3H2O
CHRYSOCOLLA
CuSiO3-2H20 + H2SO4 ~—^ CuSO4 + Si02 + 3H2O
CUPRITE
Cu2O + H2SO4 ^ CuSO4 + Cu + H2O
Cu2O + H2SO4 + Fe2(SO4)3 ^ 2CuSO4 + H2O + 2FeSO4
NATIVE COPPER
Cu + Fe2(SO4)3 ^ CuSO4 + 2FeSO4
TENORITE
CuO + H2SO4 "* CuSO4 + H2O
3CuO + Fe2(SO4)3 + 3H2O "^"^ 3 CuSO4 + 2Fe(OH)3
4CuO + 4FeSO4 + 6H2O + O2 ^~~^ 4CuSO4 + 4Fe(OH)3
CHALCOCITE
Cu2S + Fe2(S04)3 ~ ** CuS + CuSO4 + 2FeSO4
Cu2S + 2Fe2(SO4»3 " *" 2CuSO4 + 4FeSO4 + S
COVE L LITE
CuS + Fe2(SO4)3 ^ CuSO4 + 2FeSO4 + S
Chalcopyrite will slowly dissolve in acid ferric sulfate solutions and also
will oxidize according to:
CuFeS2 + 2O2 ^ CuS + FeSO4;
CuS + 2O2 ^ CuSO4.
Pyrite oxidizes according to:
2FeS2 + 2H2O + 7O2 ~ ^ 2FeSO4
V-38
DRAFT
-------
DRAFT
Heap leaching of wastes approaching a better grade ore is
usually done on specially prepared surfaces. The time cycle
is measured in months. Copper is dissolved from porous oxide
ore. Very little differentiates heap from dump leaching.
In the strictest sense, the pad is better prepared, the volume
of material is less, the concentration of acid is greater,
acid is not regenerated due to the absence of pyrite in the
ore, and the ore is of better copper grade in heap leaching,
compared to dump leaching.
In-situ leaching techniques are used to recover copper from
shattered or broken ore bodies in place on the surface or in
old underground workings. Oxide and sulfide ores of copper
may be recovered over a period of years. The principle is
the same as in dump or heap leaching. Usually, abandoned
underground ore bodies previously mined by block-caving
methods are leached although, in at least one case, an ore
body on the surface of a mountain was leached after shattering
the rock by blasting. In underground workings, leach solution
is delivered by sprays, or other means, to the upper areas
of the mine and allowed to seep slowly to the lower levels,
from which the solution is pumped to the precipitation plant
at the surface. The surface leaching operation is similar
to a heap or dump leach.
Recovery o_f_ Copper From Leach Solutions. Copper dissolved
in leach solutions may be recovered by iron precipitation,
elettrowinning, or solvent extraction (liquid ion exchange).
Hydrogen reduction has been employed experimentally.
Copper is often recovered by iron precipitation as cement
copper. Burned and shredded scrap cans are most often used
as the source of iron, although other iron scrap and sponge
iron may also be used. In 1968, 12 percent of the domestic
mine copper production was in the form of cement copper re-
covered by iron precipitation. Examples of iron launders
and cone precipitators are shown in Figures V-10 and V-ll.
The pregnant copper solution (0.5 to 2.2 g/1) is passed over
shredded or burned iron scrap and precipitates copper by
replacement according to the reaction:
CuS04 + Fe«=^Cu + FeS04
V-39
DRAFT
-------
DRAi-T
Figure V-10. TYPICAL DESIGN OF GRAVITY LAUNDER/PRECIPITATION PLANT
DRAINS
DRAINS
SIDE VIEW
, - C
' - \
' ' \
...
'\
. -j!
i i-
u.
f CELL
SOLUTION
j- CELL FLOW -
5*
i
4*- _
i
r-
J- > *
C
T 0
1- 3
r
j*
j
T
-1- *-
J
J
\
x
N
J
TOP VIEW
CANS
FLOW
DRYING
PAD
PERFORATED
SCREEN
DECANT
BASIN
'•v.':;.r-n\W^v ^\-^^-w* vwl-
1 / -:-N W •*' ' •
r
END VIEW
SOURCE: REFERENCE 23
V-40
DRAFT
-------
DRAFT
Figure V-11. CUTAWAY DIAGRAM OF CONE PRECIPITATOR
BARREN
SOLUTION
COOPER
SETTLING AND
COLLECTION
ZONE
COPPER DISCHARGE
SCRAP IRON
DYNAMIC
ACTION
ZONE
COPPER-BEARING
SOLUTION
SOURCE: REFERENCE 23
V-41
DRAFT
-------
DRAFT
Scrap iron of other forms and sponge iron may be employed.
Gravity iron launders employ gravity to allow solutions to
trickle over and through iron scrap. Spray water washes
remove copper frequently from the can surfaces. Occasionally,
solution is introduced from below and flows upward through
the iron to produce a coarser, but highly pure, cement copper.
(See Figure V-10.)
Cone precipitators may be employed for copper recovery. Solu-
tion is injected, through nozzles at the bottom of the cone,
into the shredded iron scrap. This injection, under pressure,
both precipitates copper rapidly and removes it from the iron
surface by the turbulent action. (See Figure V-ll.)
Precipitated copper is recovered by draining and scooping out
the solids. Recovery from pregnant solution may be 60 per-
cent. The resulting cement copper is 85 to 99 percent pure
and is sent to the smelter for further purification.
The barren solution from a precipitation plant is recycled from
a holding pond to the top of the ore body, after sulfuric
acid and makeup water are added, if necessary.
Leach solutions containing greater than 25 to 30 grams per
liter of copper are usually sent to electrowinning facili-
ties. The cathode copper produced is highly pure and
does not require smelting.
Solvent extraction of copper from acid leach solutions by
organic reagents is rapidly becoming an important method of
recovery. When pregnant liquors contain less than 30 grams
of copper per liter, the process is most applicable. (See
Figure V-12.)
In solvent extraction, a reagent with high affinity for
copper and iron in weak acid solutions, and with low affinity
for other ions, is carried in an organic medium. It is
placed in intimate contact with copper leach solutions,
where H+ ions are exchanged for Cu(-H-) ions. This regener-
ates the acid, which is recycled to the dump. The organic
medium, together with copper, is sent to a stripping cell,
where acidic copper sulfate solutions exchange H+ ion for
Cu(-H-) . This regenerates the organic/H+ media and passes
copper to the electrolytic cells, where impurity-free copper
V-42
DRAFT
-------
DRAFT
Figure V-12. DIAGRAM OF SOLVENT EXTRACTION PROCESS FOR RECOVERY OF
COPPER BY LEACHING OF ORE AND WASTE
MINE
DUMP
I
WEAK CuSO4
LEACH SOLUTION
RAFFINATE
'RECYCLED "
i
SOLVENT
EXTRACTION
PLANT
Cu"1"1" ON
ORGANIC CARRIER
RECYCLED
ORGANIC
CARRIER (H+)
i
STRIPPING
AREA
RECYCLED
ACIDIC ELECTROLYTE
(H2S04)
T
CuS04
ELECTROLYTE
i
ELECTROLYTIC
RECOVERY
PLANT
CATHODE
COPPER
I
TO
STOCKPILE
V-43
DRAFT
-------
DRAFT
(99.98 to 99.99 percent Cu) is electrolytically deposited on
cathodes (electrowinning). Typically, 3.18 kg (7 Ib) of
acid is used per 0.454 kg (1 Ib) of copper produced.
Acid Leach Solution Characterization. Water sources for
heap, dump, and in situ leaching are often mine water, wells,
springs, or reservoirs. All acid water is recycled. Makeup
water needs result only from evaporation and seepage;
therefore, the water consumption depends largely on climate.
Table V-12 lists the amount of water utilized for various
operations.
The buildup of iron salts in leach solutions is the worst
problem encountered in leaching operations. The pH must be
maintained below 2.4 to prevent the formation of iron salts,
which can precipitate in pipelines, on the dump surface, or
within the dump, causing uneven distribution of solution.
This may also be controlled by the use of settling or hold-
ing ponds, where the iron salts may precipitate before recycl-
ing.
Table V-13 lists the chemical characteristics of barren leach
solutions at selected plants. This solution is always recycled
and is almost always totally contained.
Other metals, such as iron, cadmium, nickel, manganese, zinc,
and cobalt, are often found in high concentrations in leach
solutions. Total and dissolved solids often build up so that
a bleed is necessary. A small amount of solution may be sent
to a holding or evaporation pond to accomplish the control
of dissolved solids.
Handling and Treatment of Water. No discharge of pollutants
usually occurs from leaching operations, except for a bleed,
which may be evaporated in a small, nearby lagoon.
Process Description - Mill Processing
Vat Leaching. Vat leaching techniques require crushing and
grinding of high-grade oxide ore (greater than 0.4 percent Cu).
(See Figure V-13.) The crushed ore, either dry or as a
slurry, is placed in lead-lined tanks, where it is leached
V-44
DRAFT
-------
DRAFT
TABLE V-12. 1973 WATER USAGE IN DUMP, HEAP. AND IN-SITU
LEACHING OPERATIONS
MILL
2101
2103
2104
2107
2108
2110
2116
2118
2120
2122
2123
2124
2125
WATER USAGE (1973)
m3/metric ton
precipitate produced
4,848.6
1,600.0*
1. 335.1 1
967.8*
1,096.5
1,308.7
N/A
1,185.3
4,264.0
1,973.6
922.2
746.3
626.0
gallons/short ton
precipitate produced
1,162.131
383,490*
320.000t
231.967*
262.800
313,683
N/A
284,108
1.022,000
473,040
221,026
178,876
150,048
•Estimated from 1972 copper-in-precipitate production and
assuming precipitates are 85% copper (Source: Copper - A
Position Survey, 1973, Reference 24)
t Production taken from NPDES permit application
N/A = Production not available; only flow available
V-45
DRAFT
-------
DRAFT
TABLE V-13. CHEMICAL CHARACTERISTICS OF BARREN HEAP,
DUMP, OR 1N-SITU ACID LEACH SOLUTIONS
(RECYCLED: NO WASTE LOAD)
PARAMETER
pH
TS
TSS
COD
TOC
Oil and Qreasa
S
At
B
Cd
Cu
Fa
Pb
Mn
Hg
Ni
Tl
Se
Ag
Ta
Zn
Sb
Au
Co
Mo
Sn
Cyanide
CONCENTRATION (mg/£l IN LEACH SOLUTION FROM MINE
2120
3.66*
28.148
14
61 5 X
1.3
<1.0
<0.6
<0.07
0.11
7.74
38.0
2.880.0
0.1
260.0
0.0009
2.40
<1.0
< 0.003
<0.1
1.0
940.0
<0.6
<0.06
3.30
.
.
<0.01
2124
2.82 •
47.764
188
1.172
28.0
8.0
<0£
023
031
0.092
146.0
6.300.0
<0.1
94.0
O.O012
720
< 0.1
< 0.040
<0.1
1.0
2BS
<0£
<0.05
3.80
0.76
-
<0.01
2123
3.56*
44,368
162
80
27.6
2.0
<0.6
0.07
<0.01
6.66
97.0
660.0
0.1
123.6
0.0010
5.68
<0.1
0.030
<0.1
1.8
33.0
<0.5
<0.05
7.3
1.33
•
<0.01
2122
249*
83.226
34
386.1
46.0
6.0
<0.6
<0.01
0.08
4.60
72.0
3.600.0
1.14
190.0
0.0003
31.1
<0.1
<0.003
0.038
2.6
74.6
<2JO
<0.06
72.0
0.36
2.40
<0.01
2126
4.24*
29/494
218
440.0
11.0
0.0
<0*
<0.07
0.03
0.20
7.00
3.688.0
<0.1
1494
0.0007
6.90
<0.1
< 0.020
<0.1
1.1
21.0
<0.6
<0.06
13.70
0.6
•
<0.01
2104
3J9-
•
•
•
•
•
"
0.04 to 0.60
0.66
6226
•
0.68
•
0.0003
•
"
0.13
~
•
•
"
'
•
•
•
•Value In pH units
V-46
DRAFT
-------
DRAFT
Figure V-13. VAT LEACH FLOW DIAGRAM (MILL 2124)
TO ATMOSPHERE
34 m /metric ton
<8.166gal/ihort ton)
1 1 m"/metnc ton
(264 gaJ/ihort ton)
290 m'/ralric ton
(69.506 oil/ihort ton)
TO ATMOSPHERE
EVAPORATION
SLIMES
166 m /metric ton
(39.453 gil/ihart ton)
28 m'/metric tan
(5.620 pl/ihort ton)
10 m /metric ton
(2.411 gal/ihort ton)
TO MILL
TO MILL
TO WASTE
BARREN SOLUTION •
176 m3/metnc ton
(42.181 B»l/ihort ton)
_ COPPER-RICH
ELECTROLYTE
207 m'/metric ton
(49.578 gal/ihort ton)
189 mj/metnc ton
(45.176 gal/short ton)
31 m /metric ton
(7.397 gal/ihon ton)
ELECTROWINNING
TO
STOCKPILE
TO LEACH DUMPS
AND PRECIPITATION
PLANT
V-47
DRAFT
-------
DRAFT
with sulfuiic acid for approximately four days. This method
is applicable to nonporous oxide ores and is employed for
better recovery of copper in shorter time periods.
The pregnant copper solution, as drawn off the tanks, contains
very high concentrations of copper, as well as some other
metals. The copper may be recovered by iron precipitation
or by electrowinning.
Water is utilized in the crusher for dust control, as leach
solution, and as wash water. The wash water is low in copper
content and must go to iron precipitation for copper recovery.
Table V-14 summarizes water usage at vat leach plants. The
vat ores are washed and discarded in a dump. If the sulfide
concentration is significant, these ores may be floated in
the concentrator to recover CuS.
Vat Leach Water Characterization. Table V-15 summarizes the
chemical characteristics of vat leach solutions. These solu-
tions are recycled directly. Makeup water is usually required
when there are evaporative losses from the tanks and recovery
plants.
Of the three vat leach facilities surveyed, one recycles
directly. Another employs holding (evaporative) ponds for
dissolved-iron control. Still another reuses all the leach
solution in a smelter process and requires new process water.
Therefore, no discharge results.
Variation Within the Vat Leach Process. Ores which are
crushed prior to the vat leach process may be washed in a
spiral classifier for control of particulates (slimes) unde-
sirable for vat leaching. These slimes may be floated in
a section of the concentrator to recover copper sulfide and
then leached in a thickener for recovery of oxide copper.
The waste tails (slimes) are deposited in special evapora-
ting ponds. The leach solution undergoes iron precipitation
to recover cement copper, and the barren solution is sent
to the evaporation pond as well. These wastes are character-
ized in Table V-16. No effluent results, as the wastes are
evaporated to dryness in the special impoundment.
V-48
DRAFT
-------
DRAFT
TABLE V-14. WATER USAGE IN VAT LEACHING PROCESS AS A FUNCTION OF
AMOUNT OF PRODUCT (PRECIPITATE OR CATHODE COPPER)
PRODUCED
MILL
2102
2116
2124
WATER USAGE (1973)
m^/metric ton
product
133.7
52.4
206.85
gallons/short ton
product
32,040
12.568 1
49,578
METHOD OF RECOVERY
Solvent Extraction/Iron
Precipitation*
Electrowinni ng* *
Electrowinning**
* Product is cement copper or copper precipitate
t No 1973 data were received through surveys. 1972 data from Reference 24
were used to calculate a value which may be a low estimate of water use.
"Product is cathode copper
V-49
DRAFT
-------
DRAFT
TABLE V-15. CHEMICAL CHARACTERISTICS OF VAT-LEACH BARREN ACID
SOLUTION (RECYCLED: 1MO WASTE LOAD)
PARAMETER
pH
IDS
TSS
COD
TOC
Oil and Grease
Al
Cd
Pb
Cr
Cu
Fe
Mn
Ni
V
Tl
Se
Ag
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/t )
1.1*
169,000
515
331
96
1.0
1,540.0
0.42
2.0
17.0
27,800
4,800.0
47.3
1.70
2.50
<0.03
< 0.003
0.17
11.5
51.0
2.0
< 0.01
"Value in pH units
V-50
DRAFT
-------
DRAFT
TABLE V-16. MISCELLANEOUS WASTES FROM SPECIAL HANDLING OF
ORE WASH SLIMES IN MINE 2124 (NO EFFLUENT)
PARAMETER
PH
TDS
TSS
COD
TOC
Oil and Grease
Al
Cd
Cu
Fe
Pb
Mn
H8
Ni
Se
Ag
Ti
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/J.)
SLIME LEACH-THICKENER
UNDERFLOW
2.4*
19.600
292.000
515
21
4.0
320.0
0.27
4.800
5,500
0.22
2.7
0.0026
1.5
< 0.003
0.057
3.8
8.9
1.0
0.5
<0.01
SLIME PRECIPITATION-
PLANT BARREN SOLUTION
1.8*
23.000
277
226
8
1.0
305.0
0.40
4,800
4.500
0.59
3.0
0.0560
1.75
< 0.003
0.054
4.2
35.0
1.0
3.75
<0.01
* Value in pH units
V-51
DRAFT
-------
DRAFT
The process has application when mined ores contain signifi-
cant amounts of both oxide and sulfide copper.
Process Description - Froth Flotation
Approximately 98% of ore received at the mill is beneficiated
by froth floation at the concentrator. The process includes
crushing, grinding, classification, flotation, thickening,
and filtration. (See Figure V-14.)
Typically, coarse ore is delivered to the mill for two- or
three-stage reduction by truck, rail or conveyor and is
then fed to a vibrating grizzly feeder, which passes its over-
size material to a jaw crusher. The ore then travels by con-
veyor to a screen for further removal of fines ahead of the
next reduction stage. Screen oversize material is crushed
by a cone crusher. When ore mineralogy is chalcopyrite, or
contains pyrite, an electromagnet is inserted before secondary
crushing to remove tramp iron. Crushing to about 65 mesh is
required for flotation of porphyry copper.
The crushed material is fed to the mill for further reduction
in a ball mill and/or rod mill. A spiral classifier or
screen passes properly sized pulp to the flotation cells.
Ahead of the flotation cells, conditioners are employed to
properly mix flotation reagents into the pulp. (See Figure
V-15.)
Reagents employed for this process might include, for instance:
Reagent Example of lb/short ton kg/metric ton
type Reagent mill feed mill feed
pH control lime 10.0 5.0
collector Xanthate 0.01 0.005
collector Minerec 0.03 0.015
compounds
frother MIBC 0.02 0.04
The specific types of reagents employed and amounts needed
vary considerably from plant to plant, although one may
classify them, as in Table V-17, as precipitating agents,
pH regulators, dispersants, depressants, activators, collec-
tors, and frothers.
V-52
DRAFT
-------
DRAFT
Figure V-14. FLOW DIAGRAM FOR FLOTATION OF COPPER (MILL 2120)
MINING
ORE
CRUSHERS
IBS m3/metrie ton
(46.736 g>l/diart ton)
RECYCLE
REAGENTS
196 m3/matrw Ion
(46.736 ffl/ihort ton)
PROCESS
BALL MILL
CONCENTRATION
CuS
FROTH
i
IATEH
196m3/
(46.735
WATER
SOFTENING
TO ATMOSPHERE
TAILINGS-
THICKENERS
FRESH WATER
27 m3/nwtrw ton
(6.491 gal/ihort ton)
0.06 m3/metne ton
(11.1 gdMrart ton)
RECYCLES (*)+ (1»V 118 m3/n»trh ton
\*-S V-X 128.189 mri/ihort «
(28.189 ga)/ihort ton)
RECYCLE-
THICKENER
THICKENED
TAILS
108 m3/nwtric ton
(25.964 QriMiort ton)
77 m3/mitiic ton .
(18^46 orf/dnit ton) J
IS m3/mtrle ton
(3.709 gri/diort ton)
EVAPORATION
TAILING
POND
Il6m3/matne ton
(3.709 gal/ihort ton)
OVERFLOW
(IF ANY)
DISCHARGE
V-53
DRAFT
-------
DRAFT
Figure V-15. ADDITION OF FLOTATION AGENTS TO MODIFY MINERAL SURFACE
PULP FROM GRINDING CIRCUIT
(25-45% SOLIDS)
REAGENTS TO
ADJUST pH
'
CONDITIONER
i
WETTING AGENT
DISPERSANT
CONDITIONER
I
COLLECTOR
CONDITIONER
•FROTH-
i
TO
ADDITIONAL
PROCESSING
FLOTATION
CELLS
ACTIVATOR
(OR DEPRESSANT)
-TAILINGS-
TO
WASTE
V-54
DRAFT
-------
DRAFT
TABLE V-17. EXAMPLES OF CHEMICAL AGENTS WHICH MAY BE
EMPLOYED IN COPPER FLOTATION
MINERAL
Bornite
Chalcocite
Chalcopyrite
Native Copper
Azunte
Cuprite
Malachite
PRECIPITATION AGENT
—
—
—
—
Sodium momnulfide
Sodium monosulfidc
Sodium monosulfide
PH
REGULATION
Lime
Lime
Lime
Lime
Sodium urbonate
Sodium carbonate
Sodium carbonate
DISPERSANT
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
DEPRESSANT
Sodium cyanide
Sodium cyanide
Sodium cyanide
Sodium cyanide
Quebracho
Quebracho
Tannic acid
ACTIVATOR
—
—
—
—
Polysuliide
Polysulfidc
Polysulfide
COLLECTOR
Xanthdte
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats.
Fatty acids
and salts
Fatty acids
and sdlts.
Xanthates
Fatty acids
and salts.
Xanthates
FROTHER
Pine oil
Pine oil
Pine oil
Pine oil
Pine oil.
Vapor oil,
Cresylic
acid
Pine oil.
Vapor oil.
Cresylic
acid
Pine oil.
Vapor oil.
Cresylic
acid
Source Reference 25
V-55
DRAFT
-------
DRAFT
Rougher-cell concentrate is cleaned in cleaner flotation
cells. The overflow is thickened, filtered, and sent to
the smelter. Tailings (sands) from the cleaner cells are
returned to the mill for regrinding. Tailings from the
rougher cells are sent to the tailing pond for settling of
solids. Scavenger cells, in the last cells of the rougher
unit, return their concentrate (overflow) to one of the
first rougher cells.
In flotation, copper sulfide minerals are recovered in the
froth overflow. The underflow retains the sands and slimes
(tailings). The final, thickened and filtered, concentrate
contains 15 to 35 percent copper (typically, 25 to 30 percent)
as copper sulfide. Copper recoveries average 83 percent,
so a significant portion of the copper is discarded to tailing
ponds. Tailings contains 15 to 50 percent solids (typically,
30 percent) and 0.05 to 0.3 percent copper.
Selective or differential flotation is practiced in copper
concentrators, which (for example) are used for separation
of molybdenum from copper concentrate, for separation of
copper sulfide from pyrite, and for separation of copper
sulfide from copper/lead/zinc ore. Silver may be floated
from copper flotation feed; gold and silver may be leached
by cyanide from the copper concentrate, with precipitation
by zinc dust.
Water Usage iri Flotation. The major usage of water in the
flotation process is as carrier water for the pulp. The carrier
water added in the crushing circuit also serves as contact
cooling water. Sometimes, water sprays are used to control
dust in the crusher. Process water for flotation comes from
mine-water excess, surface and well water, recycled tailing
thickener, and lagoon water. The majority of the copper
industry recycles and reuses as much water as is available
because the industries are located in an arid climate
(i.e., Arizona, New Mexico, and Nevada). There are plants
in areas of higher rainfall and less evaporation which have
reached 70, 95, and 100 percent recycle (or zero discharge)
and are researching process changes and treatment technology
in order to attain zero discharge of all mill water. Three
major copper mills discharge all process water from the
tailings at this time.
V-56
DRAFT
-------
DRAFT
Table V-18 outlines the amount of water used in flotation per
ton of concentrate produced.
Noncontact cooling water in the crushers, if not entirely
in a closed circuit, may be reused in the flotation circuit
and either settled in holding ponds prior to recycle or
evaporated. The use of noncontact cooling water in crushing
appears to be rare, since pulp carrier water serves as contact
cooling water.
Waste Characterization. The chemical characteristics of
tailing-pond (settled) decant water are summarized in Table
V-19. Residual flotation agents or their degradation pro-
ducts may be harmful to aquatic biota, although their consti-
tuents and toxicity have not been fully determined. Their
presence (if any), however, does not appear to hamper the
recycling of tailing decant water to the mill process. Water
is characterized by 1 to 4 grams per liter of dissolved
solids and by the presence of alkalinity, sulfate, surfactant,
and fluoride. Dissolved metals in decant water are usually
low, except for calcium (from lime employed in flotation
process), magnesium, potassium, selenium, sodium, and
strontium—which do not respond to precipitation with lime.
On occasion, cyanide, phenol, iron, lead, mercury, titanium,
and cobalt are detectable in the decant. However, in these
cases, the water is either recycled fully or partially dis-
charged.
Handling or Treatment of Decanted Water From Mill Tailing
Ponds. The majority of the industry recycles all mill pro-
cess water from the thickeners and the tailing pond due to
the need for water in the areas of major copper-ore production.
Of the balance of the industry, which includes approximately
six major copper producing facilities and an undetermined
number of operations producing copper as a byproduct, at
least half (50 percent) are currently working toward attaining
recycle of mill process water. Also, of the six, three
have sophisticated lime and settling treatment, or are
installing it, to protect the quality of the discharge.
Three of the copper mills surveyed , aJl ->t wh^ch discharge
water from the railing pond, are compared Ln Table V-20 as
to the quality of, and the amount of loading in, the discharged
V-57
DRAFT
-------
DRAFT
TABLE V-18. WATER USAGE IN FROTH FLOTATION OF COPPER
MILL
2101
2102
2103
2104
2106
2108
2109
2111
2112
2113
2114
2115
2116
2117
2118
2119
2120
2121
2122
2123
2124
WATER USAGE (1973)
m^/ metric ton
concentrate
produced
95.8
188.7
77.6
474.3
36.0
141.9
N.P.
*
280.4
78.6
68.3
85.5
366.7
51.8
145.0
112.0
161.6
234.7
149.4
160.9
370.9
110.3
gal/short ton
concentrate produced
22.967
45.233
18.610
113.674
8.625
34,009
N.P.
67.201 •
18,847
16,377
20,503
87,888
12.417
34,763
26,846
38,738
56,257
35,801
38.570
88,905
26,440
•Concentrate production estimated from known copper content and
assuming concentrate contains 20.43% copper, as in 1972
N.P. = No (1973) production
SOURCE: Reference 24
V-58
DRAFT
-------
DRAFT
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAILING PONDS (Sheet 1 of 4)
PARAMETER
Flow
pH
SES'
TOS
TSS
Oil >nd Giuw
s.o2
Al
A>
Cd
Cu
ft
Pb
Mn
Hg
Mi
S.
Ag
Sf
Zn
Sb
Co
Au
Mo
PfMMpfWM
Cyan*.
OpiMtina
din/rMr
Annual
Production
oi Conctnttata
MILL 2119
CONCENTRATION
\rntU)
21 3,966 m3/d.¥
156.630.000 B»l/dav)
912-
3OX
1.422
4
<10
-
< 10
<007
< 05
<006
OOOM
<01
<0003
< 01
0*3
< 005
<02
< 005
<005
< 02
< 001
RAW WASTE LOAD PER UNIT PRODUCT
koVIOOOmairic torn
168 m^/matric Ion
91Z-
-
239.032
872
<1681
-
<840
4707
1S969
<84
<84
00672
<1«8
<0504
<168
13952
<840
<336
<840
<8«0
<336
<168
Ib/IOOO ihon toni
40.290 pi/ikon ion
912-
-
478.063
1.345
<3362
-
<3362
< 2353
< 1681
94 13
31938
1681
<1661
01345
<336
< 1 OO9
<336
27904
<1681
<672
<1681
<1681
<672
<33S
357
454 420 matne ton* (5O0.9O4 ihon lorn)
MILL 2120
CONCENTRATION
Imj/t >
105.980 m3/d.v
I28.OOO.OOO gal/d>*l
11 08-
35%
2.652
< 2
30
-
-
-
<002
077
520
<01
007
00008
<01
-
01
<004
<005
-
<001
RAW WASTE LOAD PER UNIT PRODUCT
kg/lOOO niauic tons
109 1 m3/m«ric Ion
11 O9"
-
289.692
< 218
3277
-
-
-
<218
8411
56802
<109
765
00874
<546
-
<109
-
1092
<437
<109
tb/10OO ihon tons
261.829«il/ihonion
1109*
-
579.384
<437
6554
-
-
-
< 437
16822
1.13605
<218
1529
01748
<1O92
-
<21 8
-
2185
0092
< 874
O092
-
< 218
360
349.272 nmiic loni (385.000 ihon lonl
•Valu* in pH uml*
V-59
DRAFT
-------
DRAFT
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAILING PONDS (Sheet 2 of 4)
PARAMETER
Flo*
pN
SES<
TDS
TSS
Oil and Gru»
s.o2
Al
Ai
Cd
Cu
Ft
Pb
Mn
Hg
Mi
S.
Ag
Si
Zn
Sb
Co
Au
Mo
Phosphate
Cvinid.
Operating
dar»/y«
Annual
Product ion
of Conc«ntrM«
CONCENTRAIION
<002
350
10 OS
022
025
00098
<0 1
OO7
O9
<05
<004
<05
024
<001
MILL 2121 (Slim. Tub)
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 malnc ton*
1047m3/nnlriclon
93'
-
45.063
21,151
1047
49.737
6178
<733
<209
36648
1 O5223
2304
2618
10261
< 105
713
9424
<524
<4 19
< 524
<524
25 13
<10S
Ita/IOOO ihon tom
25,097 gal/shon Ion
93*
-
91.725
42.302
2094
99.474
1.236
<1466
<419
73296
2.10465
4607
5235
20523
<1047
4607
<209
1466
18848
< 1047
< 838
< 1047
< 1047
5026
<209
360
223.318 matric tom 1246 162 short ions)
MILL 2121 hand!)
CONCENTRATION
lmo/41
32 825 m3/d.y
(8.C72.4OO gil/dayl
928*
10*
310
6
<1
-
900
<007
003
46
1.216
040
48
00001
172
<0003
<01
006
850
<05
1 1
COOS
<001
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 m«lne lam
52 9 m'/metric Ion
928*
-
16.4O4
317
<370
1 59
2.4341
643453
21 17
2.5399
00053
9101
< 0159
< 53
317
44978
<265
5821
< 265
<265
-
<053
lb/1000 ihon lam
1 2.683 g>l/ihort ion
928-
_
32.808
63S
<10S8
_
95248
< 741
317
4.8682
128.6907
4233
50799
00106
18203
<0317
< 106
635
89956
< 529
11641
<529
<529
_
-------
DRAFT
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAILING PONDS (Sheet 3 of 4)
PARAMETER
Flo-
PH
SESr
IDS
TSS
Oil ind GIMM
S*>?
Al
A.
Cd
Cu
Ft
Pb
Mn
Hg
Ni
$•
A«
S>
Zn
Sb
Co
A»
Mo
PhoiptMt.
Cyanida
OpH«.no
dayi/yur
Anniul
Product 100
ol Concnural*
MILL 2122
CONCENTRATION
(mg/ei
278 O84 n>3/d«y
<73.470.000oil/d«vl
854'
16*
4,778
24
3
122S
<1
<007
<005
008
<01
278
OM7
OOO02
<01
0022
<01
181
<10
008
<005
<02
015
<001
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 mttiic tons
134 m3/m*lric ton
854*
-
573.188
3.Z17
4021
1,67580
<134
<93S
<67
1072
<134
37399
63
00268
<134
2949
< 134
24263
< 67
<134
1072
< 67
< 268
2011
<1 34
lb/1000 ihan toni
32.079 oilMon Un
854-
-
1.146.378
6.434
8043
3.361 19
<2681
< 1877
< 134
2145
< 268
74799
126
00536
<26B
5898
< 268
48525
<134
<26B1
2145
< 134
< 536
4021
<268
3S7
740 EO2 mtnc torn (817.836 short ton)
MILL 2123
CONCENTRATION
Img/ll
11 446 m3/d«y
13.024.000 atl/diy)
1300-
3O%
2.494
20
10
27
1
<007
<003
077
015
<01
<006
00019
<05
<006
<005
<2S97
< 11 13
28563
5564
< 3709
< 2226
07048
<185S
25966
< 3709
83833
<18S5
O8S5
< 2226
O8SS
<1B55
7.493 1
371
Ib/lOOO short lant
88.905 gal/short ton
1300*
-
1.85O.272
1483B
7.4189
20.031
7419
< 51 93
<2226
57125
11128
<7<19
<«S)
1 4O96
<3709
51932
< 7419
1.67667
< 3709
<3709
<4451
<3709
<3709
149862
742
362
11 ITOiMtnclom 112.313 Ihoil lam)
*V>lua in pH unili
V-61
DRAFT
-------
DRAFT
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAI LING PONDS (Sheet 4 of 4)
PARAMETER
FlOT.
pH
SES'
IDS
TSS
Oil and Graan
ao2
Al
A.
Cd
Cu
F.
Pb
Mn
HO
Ni
Sa
Al
Sr
Zn
Sb
Co
Au
Mo
Phosphata
Cvanida
Oparalmg
dayWyaar
Annual
Production
ol Concarwau
MILL 212*
CONCENTRATION
Img/EI
19.322 n.3/day
(5.1O4800galfdavl
1005*
50%
2.848
6
1
4675
<05
<007
005
912 6
1.982
035
31
00006
28
<01
1 2
56
<005
2929
208
< 001
RAW WASTE LOAD PEN UNIT PRODUCT
kfHOOO moliic ion*
100 8 m3/nwlric Ion
1005*
-
286995
60S
1008
4,714 34
<5O4
< 706
504
920178
1998677
3529
3.1261
00605
28236
COM3
11008
121 01
56471
< 5O4
2.953 65
2.0975
<10I
lb/1OOO>hort lam
24 ITOgil/irionton
10 OS1
-
573.989
1J2IO
2017
9.428 67
<1O08
< 14 12
1008
184.035 G
399 735 5
7059
6.263 2
0 1210
56471
< 06OS
< 20 17
24202
1 12942
<1008
33883
<1008
5.90729
4.195 0
< 202
362
69.362 metric lent {76.457 then iwul
• Valuo in pH uniti
V-62
DRAFT
-------
TABLE V-20. WASTEWATER CONSTITUENTS AND WASTE LOADS RESULTING
FROM DISCHARGE OF MILL PROCESS WATERS
O
30
o>
10
PARAMETER
FLOW
pH
TDS
TSS
Oil end Graase
At
B
Cd
Cu
Fa
Pb
Mn
Hg
Ni
Se
Si
Zn
Co
Cyanide
CONTROL
TREATMENT
MILL 2120-
CONCENTRATION
15.140"
96'"
3,328
8
150
(007
<001
<0005
006
< 0 10
< 01
003
00011
< 005
0043
240
<004
<001
WASTE
18.628"
96'"
61.994
149
279
< 1 304
<0186
< 0003
1 12
< 1 863
<1863
056
002
<093
0801
4471
<0931
<0745
CO 186
LOAD
42 .229" •
1 5Tt
96™
123988
298
5588
< 1 608
<0372
<0186
224
<3726
<3726
1 12
004
< 186
1 602
8942
< 1862
< I 490
<0372
70% RECVCLE
LIMING AND SETTLING
MILL 2121*
CONCENTRATION
101.937"
84'"
1 142
6
40
< 007
095
< 002
026
< 01
< 01
004
< 00001
<005
<004
<001
WASTE
2366"
84'"
190.269
1 000
666
<1166
1583
< 333
4332
0666
<1666
666
< 0017
< 833
250
5665
< B33
< 666
< 167
LOAD
56.7 16'"
84™
380538
20OO
1 332
<2382
3166
< 666
8664
< 3332
3332
1332
< 0034
< 1666
< 1332
< 334
NONE
LIMING AND SETTLING
MINE 2122
CONCENTRATION
<001
0 16
<0005
0 12
093
20
006
< 00001
<010
0030
098
<005
012
<001
WASTE
30.897"
839'"
88001
484
120 9
< 0302
434
<0 15
363
28 105
6044
1 81
<0003
< 302
0.91
2962
<151
363
<0302
LOAD
•>/10OO them loni
7.243.078' "
839'"
176.002
968
2418
<0604
968
<030
726
56.21
12088
362
<0006
<604
1.82
5924
<302
7.26
<0604
35* RECYCLE
NONE
O
a
•Influenced by acid mma water and laach solution
'influancad by mine water and smeller wastes
"Inm3/day
"in n>3/1000 metric tons
•"In gal/1000 ihort tons
111 Value in pH unit!
-------
DRAFT
decant water. In the calculations made to present these
data, no allowance was made for incoming process water.
As discussed previously, noncontact cooling water, if present,
remains either in a closed system or joins the carrier water
to the flotation cells.
Sewage from the mill is either handled in a treatment plant
or, in one case, is sent to an acid leach holding reservoir.
Overflow from the treatment plants is either discharged or
sent to the tailing pond.
Variations in Flotation Process. Flotation tailings may be
separated at the concentrator into slimes and sands. The sands
usually are transferred directly to the tailing pond.
However, in one case, the slimes (fines) are leached in a
thickener prior to rejoining the thickener underflow with
the sand tails. Sand and slimes are then sent to the tailing
pond. Thickener overflow is sent to a precipitation plant
for recovery of oxide copper (Figure V-16). This variation
is employed when mined ores contain a mixture of sulfide and
oxide copper.
Variations in Mill Processes
Dual Process. Ores which contain mixed sulfide and oxide
mineralization in equal ratios (greater than 0.4 percent copper
sulfides or oxides) may be treated with vat leaching, as well
as with froth flotation, in a dual process (Figure V-17).
Ore is crushed and placed in vats for leaching with sulfuric
acid, as described under "vat leaching." The leachate is
sent to iron precipitation or electrowinning plants for
recovery of copper. The residue, or tails, remaining in the
vats contains nonleachable copper sulfides and is treated
by froth flotation to recover the copper, as described under
"Froth Flotation."
Water usage and tailing-water quality are similar to the
processes of vat leaching and froth flotation. No discrete
discharge differences result from this variation compared to
vat leaching and froth flotation.
V-64
DRAFT
-------
DRAFT
Figure V-16. FLOWSHEET FOR MISCELLANEOUS HANDLING OF FLOTATION
TAILS (MILL 2124)
ORE
SULFIDE
FLOTATION CELLS
I
TAILINGS
CONCENTRATE
TO
STOCKPILE
i
HYDROSEPARATOR
•SANDS-
SLIMES
i
ACID
LEACH
(THICKENER)
UNDERFLOW
(SLIMES)
TO
TAILING
POND
PREGNANT
SOLUTION
BARREN
SOLUTION
PRECIPITATION
PLANT
CEMENT
COPPER
TO
STOCKPILE
V-65
DRAFT
-------
DRAFT
Figure V-17. DUAL PROCESSING OF ORE (MILL 2124)
MINING •
ORE
RECYCLED
WATER ORE
LEACH
TAILS
ACID
SOLUTION'
RECYCLED ACID-
ELECTROWINNING
TAILING
THICKENERS
RECYCLED
DECANT
CATHODE
COPPER
TO
STOCKPILE
V-66
DRAFT
-------
DRAFT
Leach/Precipitation/Flotation (LPF) Process. Mixed sulfide
and oxide mineralization may also be handled by the leach/pre-
cipitation/flotation process. Crushing may be in two or three
stages (Figure V-18). Both rod and ball mills may be employed
to produce a pulp of less than 65 mesh and 25 percent solids.
The pulp flows to acid-proof leach agitators. Sulfuric acid
(to a pH of 1.5 or 2.0) is added to the feed. The leaching
cycle continues for approximately 45 minutes. The acid pulp
then is fed to precipitation cells, where burned and shredded
cans or finely divided sponge iron (less than 35 mesh) may be
used to precipitate copper by means of an oxidation/reduction
reaction, which increases the pH of the pulp to 3.5 to 4.0:
CuS04_ + Fe —»• Cu + FeS04^
(excess)
Copper precipitates as a sponge, and the entire copper sponge,
together with pulp-sponge iron feed, is carried to flotation
cells. Flotation recovers both sponge copper and copper
sulfide in the froth by means of the proper conditioning re-
agents, such as Minerec A as a collector and pine oil as a
frother. Flotation is accomplished at a pH of 4.0 to 6.0
(+0.5). The concentrate is thickened and filtered before
it is shipped to the smelter. Copper recovery may be as high
as 91 percent. An example of reagent consumption for this
process is:
Reagent kg/metric ton Ib/short ton
type of mill feed of mill feed
Sulfuric acid 12.5 25
Sponge iron 18 36
Minerec A 0.09 0.18
Pine oil 0.04 0.08
V-67
DRAFT
-------
DRAFT
Figure V-18. LEACH/PRECIPITATION/FLOTATION PROCESS
COPPER SULFIDE CONCENTRATE
AND
SPONGE COPPER
TO STOCKPILE
V-68
DRAFT
-------
DRAFT
Lead and Zinc Oret.
The chemical characteristics of raw mine drainage are deter-
mined by the ore mineralization and by the local and regional
geology encountered. Pumping rates for required mine dewatering
in the lead and zinc ore mining industry are known to range
from hundreds of cubic meters per day to as much as 200,000
cubic meters per day (52.84 million gallons per day).
The chemical characteristic of raw wastewater from the milling
operation appear to be considerable less variable from
facility to facility than mine wastewater. The volume of
mill discharge varies from as little as 1000 cubic meters
per day (264,200 gallons per day) to as much as 16,000 cubic
meters per day (4.23 million gallons per day). When expressed
as the amount of water utilized per unit of ore processed,
quantities varying from 330 cubic meters per metric ton per
day (79,070 gal/short ton/day) to 1,100 cubic meters per
metric ton per day (263,566 gal/short ton/day) are encountered.
The sources and characteristics of wastes in each recommended
subcategory are discussed below.
Sources of_ Wastes - Mine Water (No Solubilization Potential) .
The main sources of mine water are:
(1) Ground-water seepage.
(2) Water pumped into the mine for machines and drinking.
(3) Water resulting from hydraulic backfill operations.
(4) Surface-water infiltration.
The geologic conditions which prevail in the mines in this
subcategory consist of limestone or dolomitic limestone with
little or no fracturing present. Pyrite may be present, but
the limestone is so prevalent that, even if acid is formed,
it is almost certainly neutralized j.n situ before any metals
are solubilized. Therefore, the extent of heavy metals in
solution is minimal. The principal contaminants of such mine
waters are:
V-69
DRAFT
-------
DRAFT
(1) Suspended solids resulting from the blasting,
crushing, and transporting of the ore. (Finely
pulverized minerals may be constituents of these
suspended solids.)
(2) Oils and greases resulting from spills and leakages
from material-handling equipment utilized (and,
often, maintained) underground.
(3) Hardness and alkalinity associated with the host
rock and ore.
(4) Natural nutrient level of the subterranean water.
(5) Dissolved salts not present in surface water.
(6) Small quantities of unburned or partially burned
explosive substances.
A simplified diagram illustrating mining operations and mine
wastewater flow for a mining operation exhibiting no solubil-
ization potential is shown in Figure V-19. Typically, mine
water may be treated and discharged or used in a nearby mill
as flotation-process water.
The range of chemical constituents measured for seven mines
sampled as part of this program is given in Table V-21.
The data, although limited to 4-hour composite samples obtained
during three site visits, generally confirm other data with
a narrower range of parameters. Generally, raw mine water
from this class of mine is of good quality, and any problem
parameters appear to be readily remedied by the current
treatment practice of sedimentation-pond systems.
Sources of Wastes - Mine Water (Solubilization Potential)
The sources of water in this subcategory are the same as those
for mines with no solubilization potential. The key differ-
ence in this category is the local geologic conditions that
prevail at the mine. These conditions lead to either gross
or localized solubilization caused by acid generation or
solubilization of oxidized minerals. The resultant wastewater
pumped from the mine contains the same waste parameters as
that from the preceding subcategory but also contains sub-
stantial soluble metals.
V-70
DRAFT
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DRAFT
Figure V-19. WATER FLOW DIAGRAM FOR MINE 3105
SEEPAGE . . •*-
DRILL
WATER 270mJ/day
(72,000 gpd)
MINE
1
PUMPING
7.600 m3/day
(2.000,000 gpd}
MILL FEED-WATER
RESERVOIR
FUEL AND LUBRICANT
SPILLAGE AND
LEAKAGE
EXPLOSIVE
WASTE PRODUCTS
V-71
DRAFT
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DRAFT
TABLE V-21. RANGE OF CHEMICAL CHARACTERISTICS OF SAMPLED
RAW MINE WATER FROM LEAD/ZINC MINES 3102, 3103,
AND 3104
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
NH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/S, )
7.4 to 8.1 •
180 to 196
200 to 330
2 to 138
326 to 510
< 10 to 631
<1to4
3 to 29
0.03 to 0.15
< 0.05 to 1.0
< 0.0001 to 0.0001
< 0.2 to 4.9
0.03 to 0.69
<0.02
<0.002 to 0.015
<0.02
< 0.02 to 0.06
<0.02to0.90
37 to 63
3 to 57
0.3 to 1.2
•Value in pH units
V-72
DRAFT
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DRAFT
The following reactions are the basic chemical reactions that
describe an acid mine-drainage situation:
Reaction 1^ — Oxidation of Sulf ide to Sulfate
When natural sulfuritic material in the form of a sulfide
(and, usually, in combination with iron) is exposed to the
atmosphere (oxygen) , it may theoretically oxidize in two
ways with water (or water vapor) as the limiting condition:
(A) Assuming that the process takes place in a dry
environment , an equal amount of sulfur dioxide
will be generated with the formation of (water-
soluble) ferrous sulfate:
FeS£ + 302^ - > FeSCM + S02^
(B) If, however, the oxidation proceeds in the
presence of a sufficient quantity of water
(or water vapor) , the direct formation of
sulfuric acid and ferrous sulfate, in equal
parts, results:
2FeS2^ + 2H20 + 70,2 - =» 2FeS04. + 2H2S04^
In most mining environments in this subcategory (underground,
as well as in the tailing area), reaction (B) is favored.
Reaction 2_ — Oxidation of Iron (Ferrous to Ferric)
Ferrous sulfate, in the presence of quantities of sulfuric
acid and oxygen, oxidizes to the ferric state to form
(water-soluble) ferric sulfate:
2Fe.2( 504)3^ + 2H20
Here, water is not limiting since it is not a requirement
for the reaction but, rather, is a product of the reaction.
Most evidence seems to indicate that bacteria (Thiobacillus
f errobacillus , Thiobacillus sulfooxidans) are involved in
the above reaction and, at least, are responsible for
accelerating the oxidation of ferrous iron to the ferric
state.
V-73
DRAFT
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DRAFT
Reaction 3_ — Precipitation of Iron
The ferric iron associated with the sulfate ion commonly
combines with the hydroxyl ion of water to form ferric hydrox-
ide. In an acid environment, ferric hydroxide is largely
insoluble and precipitates:
Fe2_(S04).3_ + 6H20 - * 2Fe(OH)_3 + 3H2S04
Note that the ferric ion can, and does, enter into an oxidation/
reduction reaction with iron sulfide whereby the ferric ion
"backtriggers" the oxidation of further amounts of sulfuritic
materials (iron sulfides, etc.) to the sulfate form, thereby
accelerating the acid-forming process:
Fe2_(S04)_3 + FeS£ + H20 - > 3FeS04 + 2S
S + 30 + H20 - > H2S04_
The fact that very little "free" sulfuric acid is found in
mine waste drainage is probably due to the reactions between
other soluble mineral species and sulfuric acid.
In some ore bodies, such reactions — and subsequent solubili-
zation of metals — may occur in local regions in which little
or no limestone or dolomite is available for neutralization
before the harmful solubilization occurs. Once a metal such
as copper, lead, or zinc is in solution, the subsequent
mixing and neutralization of that water may not precipitate
the appropriate hydroxide unless a rather high pH is obtained.
Even if some of the metal is precipitated, the particles may
be less than 0.45 micrometer (0.000018 inch) in size and,
thus, appear as soluble metals under current analytical
practice.
Conditions compatible with solubilization of certain metals —
particularly, zinc — are associated with heavily fissured ore
bodies. Although the minerals being recovered are sulfides,
fissuring of the ore body allows the slight oxidation of the
ore to oxides, which are more soluble then the parent minerals.
V-74
DRAFT
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DRAFT
When conditions exist which provide a potential for solubili-
zation, the mine water resulting is of a quality which requires
treatment beyond conventional sedimentation. The best current
practice suggests that the treated mine water is likely to
be of a quality inferior to raw discharge from mines where
the potential for such solubilization does not exist.
A flow diagram illustrating flows encountered in a mine of
the type described in this subcategory is shown as Figure
V-20. The characteristics of mine waters from this subcategory
are illustrated by Table V-22, which amplifies the above
observations.
These data suggest that particular problems are encountered
in achieving zinc and cadmium levels approaching the levels
of raw mine water from the class of mines with no solubili-
zation potential.
Process Description - Mill Flows and Waste Loading
The raw wastewater from a lead/zinc flotation mill consists
principally of the water utilized in the flotation circuit
itself, along with any housecleaning water used. The waste
streams consist of the tailing streams (usually, the under-
flow of the zinc rougher flotation cell), the overflow from
the concentrate thickeners, and the filtrate from concentrate
dewatering. The water separated from the concentrates is
often recycled in the mill but may be pumped with the tails
to the tailing pond, where primary separation of solids
occurs. Usually, surface drainage from the area of the
mill is also collected and sent to the tailing-pond system
for treatment.
The principal characteristics of the waste stream from mill
operations are:
(1) Solid loadings of 25 to 50 percent (tailings).
(2) Unseparated minerals associated with the tails.
(3) Fine particles of minerals—particularly, if the
thickener overflow is not recirculated.
(4) Excess flotation reagents which are not associated
with the mineral concentrates.
(5) Any spills of reagents which occur in the mill.
V-75
DRAFT
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DRAFT
Figure V-20. WATER FLOW DIAGRAM FOR MINE 3104
SEEPAGE-
MINE
(ALL WATER REQUIRED
FOR DRILLING FROM
SEEPAGE)
T
PUMPING
3.460 m3/day
(915,000 gpd)
XSEDIMENTATIONN
\^^ BASIN ^^/
DISCHARGE
FUEL AND LUBRICANT
SPILLS AND LEAKAGE
EXPLOSIVE WASTE
PRODUCTS
V-76
DRAFT
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DRAFT
TABLE V-22. RANGE OF CHEMICAL CHARACTERISTICS OF RAW MINE WATERS
FROM FOUR OPERATIONS IN SOLUBILIZATION-POTENTIAL
SUBCATEGORY
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
NH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£ )
IN RAW MINE WATER
3.0 to 8.0*
14.6 to 167
178 to 967
< 2 to 58
260 to 1.722
15.9 to 95.3
1to11
Oto3
0.020 to 0.075
< 0.05 to 4.0
0.0001 to 0.001 3
< 0.0001 to 0.0001
0.1 to 0.3
1.38 to 38.0
< 0.02 to 0.04
0.01 6 to 0.055
0.17 to 0.42
< 0.02 to 57 .2
0.12 to 2.5
48 to 775
< 0.01 to 220
0.06 to 0.80
•Value in pH units
V-77
DRAFT
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DRAFT
Figure V-21 illustrates the sources, flow rates, and fates
of water used for the flotation process in beneficiation of
lead and zinc ores.
One aspect of mill waste which has been relatively poorly
characterized from an environmental-effect standpoint is the
excess flotation reagents. Unfortunately, it is very diffi-
cult to analytically detect the presence of these reagents—
particularly, those which are organic. The TOC and MBAS
surfactant parameters may give some indication of the presence
of the organic reagents, but no definitive information is
implied by these parameters.
The raw and treated waste characteristics of four mills
visited during this program are presented in Table V-23.
Information for a mill using total recycle and one at which
mill wastes are mixed with metal refining wastes in the
tailing pond are not included in this summary. Feed water
for the mills is usually drawn from available mine waters;
however, one mill uses water from a nearby lake. These data
illustrate the wide variations caused by the ore mineralogy,
grinding practices, and reagents utilized in the industry.
Gold Ores
Water flow and the sources, nature, and quantity of the wastes
dissolved in the water during the processes of gold-ore mining
and beneficiation are described in this section.
Water Uses
The major use of water in this industry is in beneficiation
processes, where it is required for the operating conditions
of the individual process. Water is normally introduced at
the grinding stage of lode ores (shown in the process diagrams
of Section III) to produce a slurry which is amenable to
pumping, sluicing, or classification into sand and slime
fractions for further processing. In slurry form, the ground
ore is most amenable to beneficiation by the technology
currently used to process the predominantly low-grade and
sulfide gold ores—i.e., cyanidation and flotation. The
gravity separation process commonly used to beneficiate placer
gravels also requires water as a medium for separation of the
fine and heavy particles.
V-78
DRAFT
-------
DRAFT
Figure V-21. FLOW DIAGRAM FOR MILL 3103
WATER
FROM MILL
FEED RESERVOIR
9 500 m3/d»y
(2.500.000 gpd)
Zn SCAVENGERS Zn ROUCHERS
TO TAILING-
POND SYSTEM
TO STOCKPILES
(b) MILL PROCESS
V-79
DRAFT
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DRAFT
TABLE V-23. RANGES OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE
LOADS FOR MILLS 3102, 3103, 3104, 3105, AND 3106
PARAMETER
pH
Alkalinity
Hwdnn
TSS
TDS
COD
TOC
OilwdGf«n»
MBAS SurfKUnti
P
Ammonia
Hg
Pb
Zn
Cu
Cd
C>
Mn
ft
Cv«n«J.
Sulhu
Chloral*
Fluorri*
RANGE OF
CONCENTRATION
Img/CI
IN WASTEWATER
lam limit
79-
26
310
<2
670
714
11
a
018
OO42
< 00001
<01
012
<002
OOOS
<002
<002
DOS
<001
295
21
013
uppar him
88-
609
1760
108
2.834
1636
36
8
37
01SO
14
0101
19
046
036
0011
067
008
053
003
1825
395
026
RANGE OF RAW WASTE LOAD
pr* unit or* imltad
kg/1000 nutncioni
kXMr linul
_
410
460
7
940
6
636
5
0236
0108
OO64
<0127
0089
CO 026
OOOS
<0026
<0026
0064
<02S
019
<0062
0016
<0052
<0026
260
40
074
upper hrnii
_
3.200
94OO
S70
17.000
9.600
260
42
26
175
528
00052
138
344
0316
0036
344
0580
232
0218
9600
1740
168
p>r unit concMitraH produad
kg/1000 mi
tomf limit
_
1450
2.290
30
4300
30
30
30
205
054
032
< 000168
<0900
062
<018
<018
<018
< 045
0012
0091
1.260
210
203
UK lam
upon limit
-
10700
32500
2.000
50300
50000
580
130
607
254
IBS
0130
32.2
860
1.96
885
1 36
100
0198
0509
33.700
4070
545
Ib/IOOO dion tani
tomrhmn
_
2.900
4380
60
8.600
60
60
60
570
108
064
< 000336
< 18
124
< 036
< 036
<036
<090
<0024
0182
2.620
420
406
upper limit
-
20.400
66.OOO
4OOO
101JBOO
lOOjOOO
1 160
260
1214
508
370
0260
644
172
392
177
272
20
0396
1 18
67.4OO
8140
109
•VlIlM in pH umli
V-80
DRAFT
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DRAFT
Other uses of water in gold mills include washing of floors
and machinery and domestic applications. Wash water is nor-
mally combined with the process waste effluent but constitutes
only a small fraction of the total effluent. Some fresh water
is also required for pump sealing. A large quantity of water
is required in the vat leach process to wash the leached sands
and residual cyanide from the vats. Because the sands must
be slurried for pumping twice, the vat leach process requires
approximately twice the quantity of water necessary for the
milling of gold ore by any of the other leaching processes.
With the exception of hydraulic mining (dredging), water
is not normally directly used in mining operations but,
rather, is discharged as an indirect result of a
mining operation. Cooling is required in some underground
mines, and water is used to this end in air conditioning
systems. This water does not come into direct contact with
the materials or the mine and is normally discharged separately
from the mine effluent.
Water flows of four gold mining and milling operations visited
during this study are presented in Figure V-22.
Sources of Wastes
There are two basic sources of effluents containing pollutants:
(1) mines and (2) beneficiation processes. Mines may be either
open-pit or underground operations. In the case of an open
pit, the source of the pit discharge, if any, is precipitation,
runoff, and ground-water seepage into the pit. Ground-water
seepage is the primary source of water in underground mines.
However, in some cases, sands removed from mill tailings are
used to backfill stopes. These sands may initially contain
30 to 60 percent moisture, and this water may constitute a
major portion of the mine effluent. The particular waste
constituents present in a mine or mill discharge are a function
of the mineralogy and geology of the ore body and the parti-
cular milling process employed. The rate and extent to which
the minerals in an ore body become solubilized are normally
increased by a mining operation, due to the exposure of sulfide
minerals and their subsequent oxidization to sulfuric acid.
At acid pH, the potential for solubilization of most heavy
metals is greatly increased. Not all mine discharges are
acid, however; in those cases where they are alkaline, soluble
arsenic, selenium, and/or molybdenum may present problems.
V-81
DRAFT
-------
DRAFT
Figure V-22. WATER FLOW IN FOUR SELECTED GOLD
MINING AND MILLING OPERATIONS
(a) MINE/MILL 4101
3.817 n,3/
-------
DRAFT
Wastewater from a placer operation is primarily water that
was used in a gravity separation process. Where a
placer does not occur in a stream, water is used to fill a
pond on which the barge is floated. The process water is
generally discharged into either this pond or an on-shore
settling pond. Effluents of the settling pond usually are
combined with the dredge-pond discharge, and this constitutes
the final discharge. The principal wastewater constituents
from placer operations are high suspended solids.
Wastewater emanating from mills consists almost entirely
of process water. High suspended-solid loadings are the most
characteristic waste constituent of a mill waste stream.
This is primarily due to the necessity for fine grinding of
the ore to make it amenable to a particular beneficiation
process. In addition, the increased surface area of the ground
ore enhances the possibility for solubilization of the ore
minerals and gangue. Although the total dissolved-solid
loading may not be extremely high, the dissolved heavy-metal
concentration may be relatively high as a result of the
highly mineralized ore being processed. These heavy metals,
the suspended solids, and process reagents present are the
principal waste constituents of a mill waste stream.
Depending on the process conditions, the waste
stream may also have a high or low pH. The pH is of concern,
not only because of its potential toxicity, but also because
of the resulting effect on the solubility of the waste
constituents.
Process Description - Mining
Gold is mined from two types of deposits: placers and lode
(vein) deposits. Placer mining consists of excavating gold-
bearing gravel and sands. This is currently done primarily
by dredging but, in the past, has included hydraulic and
drift mining of buried placers too deep to strip. Lode
deposits are mined either by either underground (mines 4102,
4104, and 4105) or open-pit (mine 4101) methods, the parti-
cular method chosen depending on such factors as size and shape
of the deposit, ore grade, physical and mineralogical character
of the ore and surrounding rock, and depth of the deposit.
The chemical composition of raw mine effluent measured at
two of the mines visited is listed in Table V-24. Although
incomplete chemical data for mine 4102 are listed, considerable
variability was observed with respect to several key components
(TS, TDS, S04—, Fe, Mn, and Zn) .
V-83
DRAFT
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DRAFT
TABLE V-24. CHEMICAL COMPOSITION OF RAW MINE WATER FROM
MINES 4105 AND 4102
PARAMETER
PH
Alkalinity
Color
Turbidity (JTU)
TOS
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Al
As
Be
Be
B
Cd
Ca
Cr
Cu
Total Fa
Pta
CONCENTRATION (mg/£)
MINE 4105
.
275
34*
2.40
1,190
1,176
14
733
35.01
12.0
1
0.095
<0.2
0.03
< 0.002
<0.02
87.0
<0.02
<0.02
1.2
<0.1
MINE 4102"
6.15'
.
.
-
535
530
5
-
27
.
<0.1
.
0.143
0.084
-
-
-
0.025
-
-
0.056
25.11
0.82
PARAMETER
Mg
Mn
Hg
Ni
Tl
V
K
Ag
Na
Sr
Te
Ti
Zn
So
Mo
Sulfate
Nitrate
Phosphate
Cyanide
Phenol
Chloride
Fluoride
CONCENTRATION (mg/£)
MINE 4105
80.0
0.14
< 0.0001
0.10
<0.05
<0.2
44.0
<0.02
80.0
0.78
0.10
<0.5
<0.01
<1
<0.2
618.75
5.05
<0.05
<0.01
<0.01
1.8
2.05
MINE 4102"
.
11.93
.
0.06
.
.
-
-
.
.
.
.
7.31
-
.
1,192
.
.
<0.01
.
3
•
•Value in pH units
tValue in cobalt units
*• Industry data
TOS ° Total Solids
V-84
DRAFT
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DRAFT
Process Descriptions - Milling
The gold milling processes requiring water usage with subsequent
waste loading of this water, as discussed previously, are:
(1) cyanidation,
(2) amalgamation, and
(3) flotation.
There are four variations of the cyanidation process currently
being practiced in the U.S. :
(1) agitation-leaching,
(2) vat leaching,
(3) carbon-in-pulp, and
(4) heap leaching.
In general, the cyanidation process involves solubilization
of gold with cyanide solution, followed by precipitation of
gold from solution with zinc dust. (See Figure III-9.)
The agitation-leach process employed by mill 4401 requires
water to slurry the ground ore. Cyanide solution is added
to this pulp in tanks, and this mixture is agitated to main-
tain maximum contact of the cyanide with the ore. Pregnant
solution is separated from the leached pulp in thickeners,
and gold is precipitated from this solution with zinc dust.
(See Figure 111-10.)
The vat leaching process is employed by mill 4105. In this
process, vats are filled with ground ore slurry, and the
water is allowed to drain off. Cyanide solution is then
sprayed into the vats, and gold is solubilized by cyanide
percolating through the sands. Pregnant solution is collected
at the bottom of the vats, and gold is precipitated with zinc
dust.
V-85
DRAFT
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DRAFT
The carbon-in-pulp process is also used by mill 4105. This
process was designed to recover gold from slimes generated
in the ore grinding circuit. Water is added to the ore to
produce a slurry in the grinding circuit which is subsequently
cycloned. Cyclone underflows (sands) are treated by vat
leaching, while cyclone overflow is treated by the carbon-
in-pulp process. In this process, the slimes are mixed with
cyanide solution in large tanks, and contact is maintained
by agitation of the mixture (much the same as for agitation
leach). This mixture is then caused to batch flow through
a series of vats, where the solubilized gold is collected
by adsorption onto activated charcoal, which is held in
screens and moved through the series of vats countercurrent
to the flow of the slime mixtures. Gold is stripped from
this charcoal using a small volume of hot caustic. An
electrowinning process is used to recover the gold from this
solution. (See Figure III-9.)
Heap leaching has had only limited application in recent years.
This inexpensive process has been used primarily to recover
gold from low-grade ores. As the price of gold has risen
dramatically since 1970, the principal use of heap leaching
during this time has been in the recovery of gold from old
mine waste dumps. This process essentially consists of
percolating cyanide solution down through piled-up waste rock.
The leachate is usually collected by gravity in a sump; in
some cases, use is made of a specially constructed pad to
support the rock and collect the leachate.
Amalgamation can be done in a number of ways. The process
employed by mill 4102 is termed "barrel amalgamation."
This essentially consists of adding mercury to gold-containing
sands in a barrel. The barrel is then rotated to facilitate
maximum contact of mercury with the ore. The amalgam is
collected by gravity, and the gold and mercury are separated
by pressing in a hand-operated press.
Water is used by mill 4101 to slurry ground ore, making
it amenable to a flotation process. The slurried ore is
transported to conditioner tanks, where specific reagents are
added; essentially, this causes gold-containing minerals to
float and be collected in a froth, while other minerals sink
and are discarded. This separation is achieved in flotation
cells in which the mixture is agitated to achieve the frothing.
The froth is collected off the top of the slurry and is further
V-86
DRAFT
-------
DRAFT
upgraded by filtering and thickening. Tailings from the
flotation process of mill 4101 are further processed by the
cyanidation/agitation-leach process to recover residual gold
values.
In addition to suspended solids and dissolved metals, reagents
used in the mill beneficiation process also add to the pollu-
tant loading of the waste stream. The particular reagents
used are a function of the process employed to concentrate
the ore. In the gold milling industry, cyanide and mercury,
clearly, are the most prominent reagents of the cyanidation
and amalgamation processes. These reagents are also of primary
concern due to their potential toxicities. Table V-25 indicates
the quantity of each of these reagents consumed per ton of
ore milled. The bulk of these reagents which are used in
the process are present in the waste stream.
Because there is a potential solubilization of the ore minerals
present, heavy metals from these minerals may exist in the
mill waste stream. Table V-26 lists the minerals most commonly
associated with gold ore. Since settleable solids and most
of the suspended solids are collected and retained in tailing
ponds, the dissolved and dispersed heavy metals present in
the final discharge are of ultimate concern. Depending upon
the extent to which they occur in the ore body, particular
heavy metals may be present in a mill waste stream in the
range of from below detectable limits to 3 to 4 mg/1.
Calcium, sodium, potassium, and magnesium are found at con-
centrations of less than 100 mg/1 to over 1000 mg/1.
High levels of soluble metals usually result from the leaching
processes, and this is well-illustrated by the cyanide leach
process in the gold industry. Table V-27 summarizes the
chemical composition and raw waste loads resulting from four
gold milling operations. The processes represented include
amalgamation, cyanidation/agitation-leach, cyanidation/vat
leach, and the cyanidation/"carbon-in-pulp" process.
Silver Ores
Water flow and the sources, nature, and quantity of the wastes
dissolved in the water during the processes of silver-ore
mining and beneficiation are described in this section.
Coproduct recovery of silver with gold is common, and similar
methods of extraction are employed.
V-87
DRAFT
-------
DRAFT
TABLE V-25. PROCESS REAGENT USE AT VARIOUS MILLS BENEFICIATING
GOLD ORE
MILL
4105
4105
4101
4102
MILL PROCESS
Cyanidation/Leach
Cyanidation/Char-in-pulp
Cyanidation/Agitation Leach
Amalgamation
REAGENT CONSUMPTION
CYANIDATION
kg/metric ton
ore milled
0.13
0.58
0.18
-
Ib/short ton
ore milled
0.26
1.16
0.35
—
AMALGAMATION
kg/metric ton
ore milled
-
-
-
0.001
Ib/short ton
ore milled
-
-
-
0.002
TABLE V-26. MINERALS COMMONLY ASSOCIATED
WITH GOLD ORE
MINERAL
Arsenopynte
Pyrite
Chalcopynte
Galena
Sphalerite
Greenockite
Cinnabar
Pentlandite
Calverite
Sylvanite
Native Gold
Selenium
COMPOSITION
FeAsS
FeS
Cu FeS
PbS
ZnS
CdS
HgS
(Fe. Ni)g SB
Au Te2
(Au. Ag) Te2
Au
Se«
•Accompanies sulfur in sulfide minerals
V-88
DRAFT
-------
DRAFT
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
MILLING OPERATIONS (Sheet 1 of 2}
MINE/MILL
41O2
(Amaloinutnnl
4101
(Agnation Lose hi
4105
(Vat Luch)
4105 (Carbon
MI Pulp)
MINE/MILI
4102
(AmalgamBtioiO
41O1
(Agitation Luch)
4105
(W.I Laach)
4105 (Carbon
in Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laach)
41OS
(Vit Loch)
4105 (Carbon
in Pulp)
MINEflMLL
4102
(Amalgamation)
4101
(Agnation LMdil
4105
(Vil Luch)
4105 (Carbon
m-Pulpl
TSS
CONCEN
TRATION
(mat I)
495.000
545 OOO
48!>6ob ~
WASTE LOAD
in ha/1000 matnc tons
llb/IOOO shorl tons)
ol ooneentuta produced
61 695 31 5 OOO
1123.3906301
11 541.485000
123 082 93O.OOO)
-
47, I0]|
94 x 1011
rn ki/1000 matnc toni
(Ib/tOOO short torn)
of on millad
2.871 000
(5 742 000)
436.0OO
1872.000)
_
4.I71.0OO
(8.342.000)
IOC
CONCEN
TRATION
fm»/ I t
343
50 O
~~
970
CONCEN
TRATION
Ima/fcl
003
017
~
20
CONCEN
TRATION
lrng/41
1 5
< torn)
of ora millad
199
1398)
40
(80)
-
830
11 660)
Cu
WASTE LOAD
in hg/IOOO metric font
llh/IOaO short lon,l
ol concentrate produced
3740
I748O)
36OO
I72OO)
_
1941 OOO
(18820001
MI kg/IOOO nHlnc Ions
(lb/1 000 short tons)
of ora rrnllad
02
(041
01
(02)
^
17
(34)
Ft
WASTE LOAD
in kg/lOOO nwtric ions
llb/IOOO riiort ton,)
of concentrate producad
187,000
(374 OOO)
00600
K21.200I
—
74.70O.OOO
(149.400000)
ai kg/1000 marric Ions
(lb/10OOlhonton>l
of ors millad
87
11741
-04
< 081
_
66O
(9.320)
TDS
CONCEN
TRATION
(mg/£)
462
4.536
-
886
WASTE LOAD
in kft/IOOO matnc ions
(Ib/IOOO short ions)
of concantrata producad
I9^42j000
(39884.000)
96 060 OOO
1192120000)
_
8b9,900 OOO
(1719800000)
in kg/1000 malnc tons
(lb/1000 short tons)
of ora millad
930
(1.860)
3600
(7.2001
-
7.600
(15,2001
COD
CONCEN
TRATION
Img/ei
1142
43
-
17894
WASTE LOAD
in ko/IOOO matric tons
llb/IOOO titan ions)
of concanlr ata producod
1423000
12,847 000)
9 11. OOO
(1822000)
-
173.700.000
1347400000)
in kg/ 1000 matric ions
llb/tOOO short tons)
of or« millod
66
(1321
34
1681
_
1 540
130801
Ai
CONCEN
TRATION
(mg/ei
< 8.700
K 17. 400)
IO6
12121
1.510000
(30200001
-
in kg/1000 malnc ions
llb/IOOO short tonsl
of oi« millad
<04
K08I
O04
(O08I
14
(281
-
Zn
CONCEN
TRATION
(mo/ei
1 3
31
-
022
WASTE LOAD
in kg/IOOO matric tons
llb/IOOO shot tons)
of concanlrata producad
162000
(324000)
65600
(131 200)
-
213,000
(426,0001
in kg/IOOO matnc tons
Hb/lOOOihort Ions)
of ora rnillad
75
115 1)
25
151
-
2
(4)
V-89
DRAFT
-------
DRAFT
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
MILLING OPERATIONS (Sheet 2 of 2)
MINE/MILL
4102
(Amalgamation)
4IO1
(Agitation Loch)
41OS
(V.I Laach)
4105 (Carbon-
in-Pulp)
MINE/MILL
41O2
(Amalgamation)
4101
(Agitation Loch)
4105
(V.t Laachl
41 OS (Carbon.
in Pulp)
MINE/MILL
4102
(Amalgamation)
41O1
(Agnation L«achl
4105
(V.I Laachl
4105 (Carbon-
m-Pulp>
Pb
CONCEN
TRATION
(ma/ HI
<01
<01
WASTE LOAD
m kd/IOOO matrictoni
(Ib/IOOO short torn)
of eoncantrata producad
<12.5OO
K 25.000)
<2.100
K 4.200)
~
07.000
K194.OOO)
m kg/1000 matric torn
(ll/IOOO short torn)
ol or. nalrad
<06
kUt
<09
«18)
Hg
CONCEN-
TRATION
Imo/i)
00011
—
0004
00042
WASTE LOAD
in kg/IOOO matric lorn
(IbSIOOO short ions)
ofl eoncantrata producad
137
1274)
_
1.70O
13.400)
4.07O
18.140)
in kg/1000 matric torn
(Ib/IOOO short ton)
ol ora rmllad
00064
(00128)
-
0016
(0032)
0036
(0 072)
SULFIOE
CONCEN-
TRATION
(mg/ll
<05
02
1 7
WASTE LOAD
in kg/IOOO matric tons
(Ib/IOOOrfioil font)
ol concantrata produoad
< 62.000
K124.OOOI
<1O.600
K 2 1.2001
86.000
I172.OOOI
1.650.000
(3.300.0001
m kg/1000 matric toni
(fc/1000 short tons)
ol ora millad
<29
K58I
^04
K08)
08
116)
15
(30)
Cd
CONCEN-
TRATION
Img/e)
<002
010
<001
<002
WASTE LOAD
m kg/1000 matric tons
(B>/10OOdionlora)
ol concantr.ta produoad
< 2.500
K5.000I
2.100
(4^00)
< 4.300
K 8.6001
< 19.400
K 38.800)
in kg/ 1000 matrn tdni
(Ib/IOOO short tons)
ol ora millad
<0 1
CO 2)
008
(016)
<004
«0 08)
<017
K0.34)
CVANIOE
CONCEN
TRATION
(mg/ei
<001
506
-
006
WASTE LOAD
in kg/IOOO matric tons
(Ib/IOOO shon tons)
ol coneantrala producad
<1J50
K2300)
107000
(214.000)
__
58000
1116.000)
in kg/1000 matric tons
(Ib/IOOO short ion si
ol ora millad
<006
K012I
4
18)
-
052
(104)
V-90
DRAFT
-------
DRAFT
Water Uses
The major use of water in the silver-ore milling industry is
in the beneficiation process, where it is required for the
operating conditions of the process. It is normally intro-
duced at the ore grinding stage of lode ores (see process
diagrams, Section III) to produce a slurry which is amenable
to pumping, sluicing, or classification for sizing and feed
into the concentration process. In slurry form, the ground
ore is most amenable to beneficiation by the technology
currently used to process the predominantly low-grade sulfide
silver ores—i.e., froth flotation. A small amount of silver
is recovered from placer gravels by gravity methods, which
also require water as a medium for separation of the fine and
heavy particles.
Other miscellaneous uses of water in silver mills are for
washing floors and machinery and for domestic purposes.
Wash water is normally combined with the process waste effluent
but constitutes only a small fraction of the total effluent.
Some fresh water is also required for pump seals.
With the exception of hydralic mining and dredging, water
is not normally directly used in mining operations; rather,
it is usually discharged where it collects as an indirect
result of a mining operation. Cooling is required in some
underground mines for the air conditioning systems. This
water does not come into direct contact with the mine and is
normally discharged separately from the mine effluent.
Water flows of some silver mining and milling operations
visited during this program are presented in Figure V-23.
Sources of Wastes
There are two basic sources of effluents containing pollutants:
mines and the beneficiation process. Mines may be either
open-pit or underground operations. In the case of an open
pit, the source of the pit discharge, if any, is precipita-
tion, runoff and ground-water seepage into the pit. Ground-
water seepage is the primary source of water in underground
mines. However, in some cases, sands removed from mill
tailings are vised to backfill stopes. These sands may initially
V-91
DRAFT
-------
DRAFT
Figure V-23. WATER FLOW IN SILVER MINES AND MILLS
S49m3/day
(145.000 gpd)
FLOTATION
M 264 m3/day
(nt 67.000 gpd)
./ 1.109 m3/d«¥
i
\
MILL
t
3.161m3/d»Y ^V rmm J ' V POI
(835^00 gpd) ^ ^ ^^-__
1.635 m3 (432.000 gri)Atay
1.132 m3 (299.000 gall/dcy
(a) MINE/MILL4401
UNDERGROUND
MINE
DISCHARGE
2.933 m3/d«v
I (776,000 gpd)
V J 646 m3/
^ ^ <144,0a
... ^ FLOTATION
^ ^ MILL
igpd» ' 1 '
RAIN
I 12m3/d>v
4(3,340 gpd)
1.600 m3/dav V *OH° / 7 7 n, J B m3/rtlv
(396.000 gpd) ^^. — __— " (715 to 914 gpd)
' 964m3/div
(252.000 gpd)
(b) MINE/MILL 4402
V-92
DRAFT
-------
DRAFT
contain 30 to 60 percent moisture, and this water may
constitute a major portion of the mine effluent.
The particular waste constituents present in a mine or mill
discharge are a function of the mineralogy and geology of the
ore body and the particular milling process employed. The
rate and extent to which the minerals in an ore body become
solubilized are normally increased by a mining operation,
due to the exposure of sulfide minerals and their subsequent
oxidization to sulfuric acid. At acid pH, the potential for
solubilization of most heavy metals is greatly increased.
Not all mine discharges are acid, however; in those cases
where they are alkaline, soluble arsenic, selenium, and/or
molybdenum may present problems in the silver-ore mining
and dressing industry.
Very minor production is obtained for silver from placer
deposits as a byproduct of gold recovery. Wastewater placer
operations utilize primarily the water that was used in the
gravity separation process. The process water is generally
discharged into either a barge pond or an onshore settling
pond. The effluent of the settling pond usually is combined
with the dredge pond discharge, and this comprises the final
discharge. The principal wastewater constituent from any
placer operations, whether silver, gold, or other materials, is
high suspended solids.
Wastewater emanating from silver mills consists almost entirely
of process water. High suspended-solid loadings are the most
characteristic waste constituent of silver-mill waste streams.
This is caused by fine grinding of the ore, making it amenable
to a particular beneficiation process. In addition, the in-
creased surface area of the ground ore enhances the possibility
for solubilization of the ore minerals and gangue. Although
the total dissolved-solid loading may not be extremely high,
the dissolved heavy-metal concentration may be relatively
high as a result of the mineralization of the ore being pro-
cessed. These heavy metals, the suspended solids, and process
reagents present are the principal waste constituents of a
mill waste stream. In addition, depending on the process
conditions, the waste stream may also have a high or low pH.
The primary method of ore beneficiation in the silver-ore
milling industry is flotation. As a result, mill waste
streams can be expected to contain process reagents.
V-93
DRAFT
-------
DRAFT
Process Description - Mining
As discussed previously, very little water use is encountered
ins silver-ore mining, with the exception of dredging for
recovery of silver from gold mining operations. As a result
of sampling and site visits to mining operations in the silver
mining industry, the waste constituents of raw silver-mine
water were determined and are presented here in Table V-28.
Suspended-solid concentrations are low, while dissolved-solid
concentrations constitute the measured total-solid load.
Chlorides and sulfates are the principal dissolved-solid
constituents observed. Heavy-metal concentrations observed
are not notable, with the exception of total iron, total
manganese, and antimony.
Process Description - Milling
Milling processes of silver ore which require water and result
in the waste loads present in mill water are:
(1) flotation,
(2) cyanidation, and
(3) amalgamation.
The selective froth flotation process can effectively and
efficiently beneficiate almost any type and grade of sulfide
ore. This process is employed by mills 4401 and 4403 to
concentrate the silver-containing sulfide mineral tetrahedrite
and by mill 4402 to concentrate free silver and the silver
sulfide mineral argentite. In this flotation process, water
is added in the ore grinding circuit to produce a slurry for
transporting the ore through the flotation circuit. This
slurry first flows through tanks (conditioners), where various
reagents are added to essentially cause the desired mineral
to be more amenable to flotation and the undesired minerals
and gangue to be less amenable. These reagents are generally
classified as collectors, depressants, and activators, according
to their effect on the ore minerals and gangue. Also, pH
modifers are added as needed to control the conditions of the
reaction. Following conditioning, frothing agents are added,
and the slurry is transported into the flotation cells,
where it is mixed and agitated by aerators at the bottom
of the cells. The collector and activating agents cause the
V-94
DRAFT
-------
DRAFT
TABLE V-28. RAW WASTE CHARACTERISTICS OF SILVER
MINING OPERATIONS
PARAMETER
PH
Acidity
Alkalinity
Color
Turbidity (JTU)
TOS
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
Pb
Mg
CONCENTRATION (mg/£)
MINE 4401
8.0«
10.2
85.0
47t
2.0
504
504
<2
2403
11.9
17
4
0.085
<0.2
<0.07
< 0.002
<0.6
0.11
<0.02
46.0
C0.1
<0.02
0.33
<0.1
27.5
MINE 4403
.
4.2
76.2
<5*
23
622
622
<2
424.8
19.8
16
2
0.030
<0.2
< 0.002
<0.5
0.09
<0.02
44.5
<0.1
<0.02
2.05
0.18
32.0
PARAMETER
Mn
Hg
Ni
Tl
V
K
Se
Ag
Na
Sr
Ta
Ti
Zn
Sb
Mo
Chloride
Sulfata
Nitrate
Phosphate
Cyanide
Phenol
Fluoride
Kjeldahl N
Sulfide
Si02
CONCENTRATION (mg/£)
MINE 4401
0.43
0.0020
0.09
<0.1
<0.2
8.0
0.126
<0.02
7.0
0.15
<0.3
<0.5
<0.02
<0.2
<0.2
4.2
175
2.45
0.3
<0.01
<0.01
0.26
<0.2
<0.5
9.75
MINE 4403
6.3
0.0004
0.06
<0.1
<02
145
0.068
<0.02
12.0
0.21
<0.3
<0.5
0.03
<0.2
<0.2
1.15
338
0.10
0.25
<0.01
<0.01
0.21
-
<0.5
13.0
•Value in pH units
Value in cobalt units
TOS <= Total Solids
V-95
DRAFT
-------
DRAFT
desired mineral to adhere to the rising air bubbles and
collect in the froth, while the undesired minerals or gangue
are either not collected or are caused to sink by depressing
agents. The froth containing the silver mineral(s) is
collected by skimming from the top of the flotation cells and is
further upgraded by filtering and thickening (Flow sheets-Section III)
Recovery of silver is also accomplished by cyanidation at
mill 4105. This process has been discussed in the part of
Section V covering gold ores.
Currently, amalgamation is rarely used for the recovery of
silver because most of the ores containing easily liberated
silver have been depleted. The amalgamation process is
discussed in Sections III and V under gold-ore beneficiation
methods.
Quantity of Wastes
Discharge of water seldom exists from open-pit mines. However,
most underground mines must discharge water, and the average
volume of this water from the crossection of mines visited
ranges from less than 199 cubic meters per day (50,000 gallons
per day) to more than 13,248 cubic meters per day (3.5 million
gallons per day). Where mine discharges occur, the particular
metals present and the extent of their dissolution depend
on the particular geology and mineralogy of the ore body and
on the oxidation potential and pH prevailing within the mine.
Concentrations of metals in mine effluents are, therefore,
quite variable, and a particular metal may range from below the
limit of detectability upwards to 2 ppm. Calcium, sodium,
potassium, and magnesium may be present in quantities of less
than 5 ppm to about 50 ppm for each metal. However, the heavy
metals are of primary concern, due to their toxic effects.
Minerals known to be found in association with silver in
nature are listed in Table V-29.
For the facilities visited, the volumes of the waste streams
discharging from mills processing silver ore range from 1,499
to 3,161 cubic meters per day (396,000 to 835,200 gallons per day).
These waste streams carry solids loads of 272 to 1,542 metric
tons per day (300 to 1,700 short tons per day) from a mill,
depending on the mill. Where underground mines are present,
the coarser solids may be removed and used for backfilling
stopes in the mine. While the coarser material is easily
V-96
DRAFT
-------
DRAFT
TABLE V-29. MAJOR MINERALS FOUND ASSOCIATED
WITH SILVER ORES
MINERAL
Tetrahedrhe
Tennantite
Galena
Sphalerite
Chalcopyrite
Pyrite
Naumannite
Greenockite/
Xanthochroite
Garnierite
Pentlandite
Native Bismuth
Argenite
Stephanite
Stibnite
COMPOSITION
(Cu, Fe. Ag)i2 AS4$13
(Cu, Fe. Ag)i2Sb4Si3
PbS
ZnS
CuFeS2
FeS
Ag2S
CdS
(Mg, Ni) 0- Si 02 • x H2O
(Fe, Ni)g S8
Bi
Ag2S
Ag5 Sb $4
Sb2S3
V-97
DRAFT
-------
DRAFT
settled, the very fine particles of ground ore (slimes) are
normally suspended to some extent in the wastewater and often
present removal problems. The quantity of suspended solids
present in a particular waste stream is a function of the ore
type and mill process because these factors determine how
finely ground the ore is.
Heavy metals present in the minerals listed in Table V-29
may also be present in dissolved or dispersed colloidal form
in the mill waste stream. Since the settlable solids, and
most suspended solids, are collected and retained in tailing
ponds, the dissolved and dispersed heavy metals present in
the final discharge are of concern. Depending on the extent
to which they occur in the ore body, particular heavy metals
may be present in a mill waste stream in the range of from
below detectable limits to 2 to 3 ppm. Calcium, sodium,
potassium, and magnesium normally are found at concentrations
of 10 to 250 ppm each. In addition to the suspended solids
and dissolved metals, reagents used in the mill beneficiation
process also add to the pollutant loading of the waste stream.
The particular reagents used are a function of the process
employed to concentrate the ore. In the silver milling
industry, the various flotation reagents (frothers, collectors,
pH modifiers, activating agents, and depressants) are the most
prominent reagents of the flotation process. Table V-30
indicates the quantity of these reagents consumed per ton of
ore milled. A portion of these reagents which are consumed
in the process is present in the waste stream. Note that a
large number of compounds fall under the more general categories
of frothers, collectors, etc. At any one mill, the particular
combination of reagents used is normally chosen on the basis
of research conducted to determine the conditions under which
recovery is optimized. While flotation processes are
generally similar, they differ specifically with
regard to the particular reagent combinations. This is
attributable, in part, to the highly variable mineralization
of the ore bodies exploited. Waste characterizations and raw
waste loadings for mill effluents employing flotation and
cyanidation in four mills are presented in Table V-31. These
characterizations and loadings are based upon analysis of raw
waste samples collected during site visits.
V-98
DRAFT
-------
DRAFT
TABLE V-30. FLOTATION REAGLNTS USED BY THREE MILLS TO BENEFICIATE
SILVER-CONTAINING MINERAL TETRAHEDRITE (MILLS 4401 AND
4403) AND NATIVE SILVER AND ARGENTITE (MILL 4402)
REAGENT
M.I. B.C. (Methylisobutylcarbinol)
0-52
Z-200 (Isopropl ethylthiocarbamate)
Lime (Calcium oxide)
Sodium cyanide
PURPOSE
MILL 4401
Frother
Frother
Collector
pH Modifier
and Depressant
Depressant
CONSUMPTION
g/metric ton
ore milled
0.00498
0.00746
0.00187
0.109
0.00498
Ib/short ton
ore milled
0.00000995
0.0000149
0.00000373
0.000219
0.00000995
MILL 4402
Cresylic acid
Mineral oil
Dowfroth 250 (Polypropylene glycol
methyl ethers)
Aerofroth 71 (Mixture of 6/9-carbon
alcohols)
Aerofloat 242 (Essentially Aryl
dithiophosphoric acid)
Aero Promoter 404 (Mixture of
Sulfhydryl type compounds)
Z-6 (Potassium amyl xanthate)
Sulfuric acid
Soda ash (Sodium carbonate)
Caustic soda (Sodium hydroxide)
Hydrated lime (Calcium hydroxide)
Frother
Frother
Frother
Frother
Collector
Collector
Collector
pH Modifier
pH Modifier
pH Modifier
pH Modifier
2.83
6.9
0.545
10
90
1.82
70
250
1.260
3.03
320
0.00566
0.0138
0.00109
0.02
0.18
0.00363
0.13
0.49
2.51
0.00605
0.64
Ml LL 4403
Cresylic acid
Hardwood tar oils
M.I. B.C.
Aerofloat 242
Aerofloat 31 (Essentially Aryl
dithiophosphoric acid)
Xanthate Z-11 (Sodium ethyl xanthate)
Aero S-3477
Zinc sulfate
Sodium sulfite
Frother
Frother
Frother
Collector
Collector
Collector
Collector
Depressant
Depressant
1.25
1.25
3.75
7.51
5.00
250
25
150
200
0.0025
0.0025
0.0075
0.015
0.01
0.005
0.05
0.3
0.4
V-99
DRAFT
-------
DRAFT
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
AT MILLS 4401, 4402. 4403, AND 4105 (Sheet 1 of 2)
MILL
4401
4106
(Company Data
only)
MILL
4401
4403
4402
410S
(Company Data
only!
MILL
440.1
4402
it OS
(Company Gala
only!
MILL
44O1
4403
44O2
41OS
(Company OiU
only)
CONCEN
TRATION
Img/lt
650.000
7O3.000
9000O
CONCEN-
TRATION
Img/ei
220
240
290
CONCEN
TRATION
img/l)
026
OO3
022
CONCEN
TRATION
<01
056
TSS
WASTE LOAD
in kg/lOOO matric tom
lib/1000 atiart tan)
of concontral. produced
99.000.000
(198.0000001
33.901 OOO
162.802.0001
9720.000
119 440 OOO)
-
m MB/1 OOO matrie loni
(lb/1 OOO riiort torn)
ol or. millad
2.475.000
(4.9SO.OOO)
1S43000
1X088.0001
990,000
(1.980.000)
=
TOC
WASTE LOAD
in kg/lOOO matiie font
llb/1 000 dia.1 torn)
of oDimnlrali producad
4000
(8.0OO)
4.000
(B.OOO)
3.130
(6.2601
-
n kg/1000 imtric loni
(tb/1000 ihorl tom)
ol or. mllad
100
(2OOI
180
(360)
320
(640)
-
Cu
WASTE LOAD
in ka/lOOO mnric font
llb/lOOOriiorltom)
of concantrat. producad
46
1901
5
(10)
24
(481
-
in kg/IOOO ma»ic tont
Ib/IOOO riiort torn)
ol ora milUd
1
(2)
023
(046)
24
(481
-
Pb
WASTE LOAD
in fca/lOOO matrie tone
llb/100Ofhar< ron.l
of tonmnim pioducad
<18
K36I
< 17
K34I
60
1120)
-
in kg/1000 matnc rant
<045
K090)
<08
Kl 6)
62
11241
-
TDS
CONCEN
TRATION
lino/21
470
684
960
~
WASTE LOAD
in kg/ 1000 mafric Mni
lib/1000 inortlonil
of conoantnla producad
84.600
(169.2001
97,500
1195.000)
104 .OOO
1208^100)
-
in kg/1000 malnc ton»
lib/1000 Ihort lout)
ol ora millad
2.110
14.2201
4.440
(8,8801
10600
(21,2001
COD
CONCEN
TRATION
lmg/£)
595
159
227
WASTE LOAD
m kg/IOOO manic Mm
lib/1000 ihort torn!
ol CMcantrita producad
10.700
121.4001
2.700
15.400)
2.450
I4JOO)
-
in kg/1000 manic lont
(Ib/IOOO than tontl
ol ora millad
270
(540)
120
12401
250
1500)
-
Zn
CONCEN
TRATION
Irao/t)
002
017
037
WASTE LOAD
in kg/1000 imtric lorn
lib/1000 ibon lonil
ol concantraw producad
36
(72)
28
(S6I
40
ISO)
1
in kg/tOOO matfic tont
(Ib/IOOD ihort tonil
of oia millad
009
(018)
13
126)
41
182)
-
At
CONCEN
TRATION
(mg/CI
<007
< 007
007
350
WASTE LOAD
in kg/1000 matiic tani
lib/1000 ihon loni)
ol eonc.nlr.ti producad
<13
K26)
<12
K24I
8
1161
2265.000
14^30.000)
in kg/ 1000 metric lont
(Ib/lOOOihortlont1
ol or. millad
<03
K06)
-------
DRAFT
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 2 of 2)
MILL
44O1
4403
4402
4105
(Company Dill
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
44O1
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
41O5
(Company Date
only)
CONCEN
TRATION
Img/tl
00024
00008
01490
OOO4
CONCEN
TRATION
Img/l)
<03
<03
•CO 3
CONCEN
TRATION
(mg/ei
<002
<007
CONCEN
TRATION
(mg/t)
<02
<02
054
<54
K108I
<50
kiooi
<32
ft 641
-
in kg/1000 matric tom
(lb/1000 riiort ton)
of ora millad
<1 3
K26I
<23
K46I
<33
«66)
~
Ag
WASTE LOAD
in kg/1000 metric font
(lb/1000 riiort tonil
ol concantrata producad
< 36
K72I
<3
K6i
<2
K4)
-
in kg/1000 matric tont
(Il/IOOO thorium)
of ora mtllad
<009
K018)
<015
K030)
<02
K04I
-
Mo
WASTE LOAD
in kg/IOOO manic lon>
(lb/1000 short tont)
of concantrata producad
<36
«72)
<30
K60)
58
(116)
< 32.400
K 64.800)
in kg/1000 matric toni
(lb/1000 diorl torn)
of ora millad
<09
«1 8)
<4
«8)
6
1121
<03
«06)
Cd
CONCEN
TRATION
Img/t)
<002
<002
<002
<001
WASTE LOAD
m kg/1000 matric tons
(lb/1000 short tons)
of concanf rat* producad
<36
K72I
<33
K66)
<22
K44)
< 6,500
K 13.000)
in kg/1000 matric tons
(lb/1000 short tons)
of ora millad
<009
K018)
<015
K030)
<02
K04I
<006
K012)
Sa
CONCEN-
TRATION
lmg/£)
0154
0144
-
—
WASTE LOAD
m kg/1000 matric tons
(lb/1000 short tons)
of concantrata producad
28
(56)
24
(48)
-
in kg/IOOO matric tons
(lb/1000 short tons)
of ora milled
07
(1 4|
1 1
12.21
-
Ni
CONCEN
TRATION
I me/ 11
0 14
005
010
010
WASTE LOAD
in kg/1000 matric tons
lib. 1000 short tons)
of concantrata producad
250
(5001
8
1161
11
(22)
64.700
1129400)
m kg/1000 malnc Ions
(lb/1000 short tons)
of ora milled
063
1126)
04
108)
1
(21
06
112)
So
CONCEN
TRATION
Img/ll
185
230
••02
-
WASTE LOAD
m kg/1000 matric Ions
lib/1000 short tons)
of concantrata producad
333
(6661
384
1768)
<22
K44)
-
in kg/IOOO metric tons
(lb/1000 short tons)
of ore milled
83
(166)
17
134)
<2
K4)
1
V-101
DRAFT
-------
DRAFT
Bauxite Ores
Water handling and quantity of wastewater flow within surface
bauxite mines are largely dependent upon precipitation patterns
and local topography. Topographic conditions are often modi-
fied by precautionary measures, such as diversion ditching,
disposal of undesirable materials, regrading, and revegeta-
tion. In contrast, underground mine seepage occurs as a result
of controlled drainage of the unconsolidated sands in the over-
burden. These sands are under considerable water pressure,
and catastrophic collapses of sand and water may occur if
effective drainage is not undertaken. Gradual drainage accu-
mulates in the mines and is pumped out periodically for
treatment and discharge. As in other mining categories,
dewatering is an economic, practical, and safe-practice
necessity.
Beneficiation of bauxite ores is not currently practiced beyond
size reduction, crushing and grinding. No water use, other
than dust suppression, results.
Mining Technique and Sources ojE Wastewater
Open-Pit Mining. The sequence of operations that occurs
in a typical open-pit mining operation is that the mine site
is cleared to trees, brush, and overburden and then stripped
to expose the ore. Timber values are often obtained from
areas undergoing site preparation.
Depending upon the consolidation of the overburden, the
material may be vertically drilled from the surface, and explo-
sive charges—generally, ammonium nitrate—are placed for
blasting. This sufficiently fractures the overburden material
to permit its removal by earthmoving equipment, such as draglines,
shovels, and scrapers. Removal of this overburden takes
the greatest amount of time and frequently requires the
largest equipment.
Following removal of the overburden material, the bauxite is
drilled, blasted, and loaded into haulage trucks for transport
to the vicinity of the refinery. Extracted overburden or
spoils are often placed in abandoned pits or other convenient
locations, where some attempts have been made at revegetation.
V-102
DRAFT
-------
DRAFT
Regardless of the method of mining, water use at the two
existing operations is generally limited to dust suppression,
and water removal is required because it results in a hindrance
to mining. As such, mine dewatering and handling are a required
part of the mining plan at all bauxite mines.
The bauxite mining industry presently discharges about 57,000
cubic meters (15 million gallons) of mine drainage daily
at two locations. The open-pit mining technique is largely
responsible for accumulation of this water. Underground mining
accounts for only a fraction of a percent of the total.
In association with the open-pit approach to bauxite mining,
water drainage and accumulation occur during the processes
of mine site preparation and during active mining.
For the open-pit mine represented in Figure V-24, rainfall
and ground water intercepted by the terrain undergoing site
preparation are diverted to outlying sumps for transfer to
a main collection sump. Diversion ditching and drainage
ditches segregate most surface water, depending upon whether
it has contacted lignite-containing material. Contaminated
water is directed to the treatment plant, while fresh water
is diverted to other areas. At other mines, drainage occurring
during site preparation and mining is not treated, and segre-
gation of polluted and unpolluted waters may or may not be
practiced.
Water from the main collection sump is pumped to a series
of settling ponds, where removal of coarse suspended material
occurs. These ponds also aid in regulation of flow to the
treatment plant. A small sludge pond receives treated waste-
water for final settling before discharge.
Bauxite mining operations characteristically utilize several
pits simultaneously and may practice site preparation con-
current with mining. Since both bauxite producers have large
land holdings (approximately 4,050 hectares or 10,000 acres),
mines and site-preparation activities may be located in remote
areas, where the economics of piping raw mine drainage to a
central treatment plant are unfeasible. For larger quantities
of mine drainage in remote areas, separate treatment plants
appear necessary. Portable and semi-portable treatment plants
appear feasible for treating smaller accumulations of waste-
water at times when pumping of mine water for discharge is
required.
V-103
DRAFT
-------
DRAFT
Figure V-24. PROCESS AND WASTEWATER FLOW DIAGRAM FOR OPEN-PIT BAUXITE
MINE 5101
EXPLORATION AND ORE-BODY EVALUATION:
GEOLOGICAL SURVEY
TEST DRILLING
i
SITE PREPARATION:
CLEARING
STRIPPING
RUNOFF
AND
GROUND WATER
MINING:
BLASTING
LOADING
HAULING
RUNOFF
AND
GROUND WATER
MILLING:
CRUSHING AND GRINDING
STORAGE
BLENDING
2.76 m'/metric ton
(664 gal/short ton)
BAUXITE
REFINING:
COMBINATION PROCESS
2.76 ra3/metric ton
' ' (664 gal/short ton) BAUXITE
WATER TREATMENT
PLANT
o
2.76 m /metric ton
' (664 gal/short ton) BAUXITE
PRODUCTION = 2,594 metric tons (2,860 short tons) per day
WATER TREATED DAILY = 7.165 m3 (1,900,000gal)
2.76 m3/metric ton
(664 gal/short ton) bAUXiTfc'
DISCHARGE
V-104
DRAFT
-------
DRAFT
Underground Mining. Underground mining occurs where low-
silica bauxite is located deep enough under the land surface
so that economical removal of overburden is not feasible.
The underground operations have been historically notable
for relatively high recovery of bauxite under adverse con-
ditions of unconsolidated water-bearing overburden and unstable
clay floors. Controlled caving, timbered stope walls, and
efficient drainage systems—both on the surface and under-
ground—have minimized the problems and have resulted in
efficient ore recovery.
Initially, shafts are sunk to provide access to the bauxite
deposits, and drifts are driven into the sections to be mined.
A room-and-pillar technique is then used to support the mine
roof and prevent surface subsidence above the workings.
Configurations of rooms and pillars are designed to consider
roof conditions, equipment utilized, haulage gradients, and
other physical factors.
Ore is removed from the deposits by means of a "continuous
miner," a ripping-type machine which cuts bauxite directly
from the ore face and loads it into shuttle cars behind the
machine. Initial development of the room leaves much bauxite
in pillars, and it has been the practice to remove the pillars
and induce caving along a retreating caveline. However,
resultant roof collapse and fracturing can greatly increase
overburden permeability, facilitating mine-water infiltration
and subsequently increasing mine drainage problems. Recent
charges in mining technique have resulted in a cessation of
induced caving, but drainage still occurs in the mines.
Raw mine drainage accumulates slowly in the underground mines
and is a result of controlled drainage. The seepage is pumped
to the surface at regular intervals for treatment, with
subsequent settling and discharge. Excessive water in the
underground mine can lead to wetting of clays located in
drift floors and in resultant upheaval of the floor.
The most influential factor which determines mine-water
drainage characteristics is mineralization of the substrata
through which the drainage percolates. Underground mines
receive drainage which has migrated through strata of unconso-
lidated sands and clays, whereas open-pit drainage is exposed
V-105
DRAFT
-------
DRAFT
to sulfide-bearing minerals in the soil. As shown in this
section, open-pit and underground mine drainages differ
qualitatively and quantitatively; but, as a factor affecting
raw mine-drainage characteristics, mineralization does not
constitute a sufficient basis for subcategorization.
Study of NPDES permit applications and analysis of samples
secured during mine visitations revealed that the bauxite
mining industry generates two distinct classes of raw mine
drainage: (1) Acid or ferruginous, and (2) alkaline—
determined principally by the substrata through which the
drainage flows. Acid or ferruginous raw mine drainage is
defined as untreated drainage exhibiting a pH of less than
6 or a total iron content of more than 10 rag/liter. Raw
mine drainage is defined as alkaline when the untreated
drainage has a pH of more than 6 or a total iron content of
less than 10 mg/liter.
The class of raw mine drainage corresponds closely with
mining technique, and open-pit drainage is characteristically
acid. Acid mine water is produced by oxidation of pyrite
contained in lignite present in the soil overburden of the
area.
Acid mine drainage with pH generally in the range of 2 to 4
is produced in the presence of abundant water. The sulfuric
acid and ferric sulfate formed dissolve other minerals,
including those containing aluminum, calcium, manganese,
and zinc.
In areas undisturbed by mining operations, these reactions
occur because the circulating ground water contains some
dissolved oxygen, but the reaction rate is rather slow.
Mining activity which disturbs the surface of the ground
creates conditions for a greatly accelerated rate of sulfide-
mineral dissolution.
Alkaline mine water, characteristic of underground mines, may
migrate through the lignitic clays located in strata overlying
the mines before collecting in the mines, but pH is generally
around 7.5. Data evaluation reveals that underground mine
drainage differs significantly from open-pit mine drainage
(acid), as shown in Tables V-32, V-33, and V-34.
V-106
DRAFT
-------
DRAFT
TABLE V-32. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID RAW
MINE DRAINAGE FROM OPEN-PIT MINE 5101
PARAMETER
PH
Specific Conductance
Acidity
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION (mg/£)
THIS STUDY
2.8*
1.000f
397
0
560
<2
7.2
3.5
23.8
0.82
0.3
500
0.29
INDUSTRY DATA
3.0*
250
0
617
2
21.8
3.23
18.6
1.19
0.31
490
0.048
NPDES PERMIT
APPLICATION
3.5*
1.903f
—
40
1,290
10
7.0
4.2
38
1.0
0.37
700
1.4
'Value in pH units
Value in micromhos
TABLE V-33. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID
RAW MINE DRAINAGE FROM OPEN-PIT MINE 5102
PARAMETER
PH
Specific Conductance
Acidity
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sr
Sulfate
Fluoride
CONCENTRATION (mg/£)
THIS STUDY
3.2*
1.580f
782.0
0
1.154
< 2
64.0
7.7
88.0
0.36
0.063
0.1
887.5
0.59
INDUSTRY DATA**
2.8*
2.652t
533
-
-
416
62.2
-
44.6
-
-
-
726
-
NPDES PERMIT
APPLICATION
3.0*
2.000f
—
0
96
1,280
20.6
9.0
51.0
0.8
0.01
—
226
0.26
'Value in pH units '''Value in micromhos "Averages of eight or more grab samples taken in 1974
V-107
DRAFT
-------
DRAFT
TABLE V-34. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ALKALINE RAW
MINE DRAINAGE FROM UNDERGROUND MINE 5101
PARAMETER
PH
Specific Conductance
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sr
Sulfate
Fluoride
CONCENTRATION img/i)
THIS STUDY
7.2»
1.260t
280
780
<2
1.4
0.88
0.8
<0.02
<0.02
1.82
228.8
1.25
INDUSTRY DATA
7.6*
222
862
26
2.3
0.87
<0.05
<0.01
<0.01
246
0.07
NPDES PERMIT
APPLICATION
7.8*
3.281*
150
550
300
5.0
5.0
2.0
1.6
0.01
50
2.5
•Value in pH units
Value in micromhos
V-108
DRAFT
-------
DRAFT
Though these mine drainages differ with respect to mining
technique, all mine drainages sampled proved to be amenable
to efficient removal of selected pollutant parameters by
liming and settling, as exhibited in Section VII. Attainable
treated-effluent concentrations are directly related to
treatment efficiency, and these two interrelated factors
do not justify establishment of subcategories.
Due to acid conditions and general disruption of soils caused
by stripping of overburden for open-pit mines, natural
revegetation proceeds extremely slowly. The lack of vegetative
cover aids in accelerating the weathering of the unconsolidated
overburden and compounds the acid mine-water situation.
Extensive furrowed faces of exposed silt and sandy clays are
evidence of the erosion which infuses the mine water with
particulate matter. Fortunately, this material settles
rapidly, either in outlying pits or in pretreatment settling
basins, and presents no nuisance to properly treated discharges.
Raw Waste Loading
As discussed earlier in this Section, effluents from bauxite
mining operations are unrelated, or only indirectly related,
to production quantities and exhibit broad variation from
mine to mine. Loadings have been calculated for open-pit
mine 5101 and underground mine 5101, as shown in Tables
V-35 and V-36.
Potential Uses of_ Mine Water. Since both domestic bauxite
mines are intimately associated with refineries, the plausi-
bility of utilizing a percentage of mine water in the refinery
arises. Though the bauxite refining process intrinsically
has a substantial negative water balance, water is supplied
from rainfall on the brown-mud lake or from fresh-water
impoundments. More importantly, the brown-mud-like water
posseses a high pH (approximately 10) and remains amenable
to recycling in the caustic leach process.
To minimize the effects of dissolved salts in the refining
circuit, evaporators are sometimes used to remove impurities
from spent liquor. However, mine water contains many dissolved
constituents (particularly, sulfate) in large quantities, the
V-109
DRAFT
-------
DRAFT
TABLE V 35. WASTEWATER AND RAW WASTE LOAD FOR OPEN-PIT MINE 5101
PARAMETER
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION
(rng/H)
IN WASTEWATER
560 to 1290
< 2 to 10
7.0 to 21 .8
3.23 to 4.2
18.6 to 38
0.82 to 1.19
0.3 to 0.37
490 to 700
0.048 to 1.4
RAW WASTE LOAD
kg/metric ton
1.55 to 3.56
< 0.006 to 0.028
0.02 to 0.06
0.01 to 0.01
0.05 to 0.10
0.002 to 0.003
0.0008 to 0.001
1.35 to 1.93
0.0001 to 0.004
Ib/short ton
3.10 to 7.12
< 0.012 to 0.056
0.04 to 0.1 2
0.02 to 0.02
0.10 to 0.20
0.004 to 0.006
0.001 6 to 0.002
2.70 to 3.86
0.0002 to 0.008
Daily flow of wastewater = 7,165 m3 (1.900,000 gal)
Daily mine production = 2,594 metric tons (2,860 short tons)
TABLE V-36. WASTEWATER AND RAW WASTE LOAD FOR UNDERGROUND MINE 5101
PARAMETER
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION
(mg/&)
IN WASTEWATER
550 to 862
< 2 to 300
1.4 to 5.0
0.87 to 5.0
< 0.05 to 2.0
< 0.01 to 1.6
< 0.01 to 0.01
50 to 246
0.07 to 2.5
RAW WASTE LOAD
kg/metric ton
0.12 to 0.18
< 0.0004 to 0.06
0.0003 to 0.001
0.0002 to 0.001
< 0.00001 to 0.0004
< 0.000002 to 0.0003
< 0.000002 to 0.000002
0,01 to 0.05
0.00001 to 0.0005
Ib/short ton
0.24 to 0.36
<0.0008 to 0.12
0.0006 to 0.002
0.0004 to 0.002
<0.00002 to 0.0008
< 0.000004 to 0.0006
< 0.000004 to 0.000004
0.02 to 0.10
0.00002 to 0.0010
Daily flow of wastewater = 83 m (22,000 gal)
Daily mine production = 390 metric tons (430 short tons)
V-110
DRAFT
-------
DRAFT
effects of which are detrimental or undetermined at this time.
The exacting requirements of purified alumina, and the specific
process nature of the refinery, largely preclude the intro-
duction of new intake constituents via alternative water
sources (treated or untreated mine water) at this time.
Ferroalloy Ores
Waste characterization for the ferroalloy-ore mining and
milling industry has, of necessity, been based primarily on
presently active operations. Since these comprise a somewhat
limited set, many types of operations which may or will be
active in the future were not available for detailed waste
characterization. Sites visited in the ferroalloy segment
are organized by category and product in Table V-37. Since
some sites produce multiple products, and/or employ multiple
beneficiation processes, they are represented by more than
one entry in the table. Where possible, segregated as well
as combined waste streams were sampled at such operations.
Table V-37 also shows types of operations considered likely
in the U.S. in the future (marked with x's), as well as those
which represent likely recovery processes for ores not expected
to be worked soon (marked with o's). Characteristics of
wastes from the latter two groups of operations have been
determined, where possible, from historical data; probable
ore constituents and process characteristics; and examination
of waste streams expected to be similar (for example, gravity
processors of iron ore as indicators for gravity manganiferous-
ore operations).
Treatment of the individual process descriptions by ore
category, as adhered to previously in this report, is not
used here. Instead, because of the wide diversity of ores
encountered, the general character of mine and mill effluents
is discussed, followed by process descriptions and raw waste
characteristics of several representative operations.
General Waste Characteristics
Ferroalloy mining and milling wastewater streams are generally
characterized by:
(1) High suspended-solid loads
(2) High volume
(3) Low concentrations of most dissolved pollutants.
V-lll
DRAFT
-------
DRAFT
TABLE V-37. TYPES OF OPERATIONS VISITED AND ANTICIPATED-
FERROALLOY-ORE MINING AND DRESSING INDUSTRY
METAL ORE
MINED/MILLED
Chromium
Cobalt
Columbium and
Tantalum
Manganese
Molybdenum
Nickel
Tungsten
Vanadium
MINE
O
X
X
X
V(3)
V(D*
V<2)
V{1)
MILL
Category 1
(< 5.000 metric tons
[5.512 short tons] per year)
X
Category 2
(Physical
Concentration)
O
X
X
V
V
Category 3
(Flotation)
X
X
X
V(3)
X
V
Category 4
(Leaching)
O
X
X
V
V
( ) indicates number of operations visited
* seasonal mine discharge, not flowing during visit
X likely in the future; currently, not operating
O most likely process, if ever operated in the U.S.
V types of operations visited
V-112
DRAFT
-------
DRAFT
The large amounts of material to be handled per unit of
metal recovered, the necessity to grind ore to small particle
sizes to liberate values, and the general application of wet
separation and transport techniques result in the generation
of large volumes of effluent water bearing high concentrations
of finely divided rock, which must be removed prior to dis-
charge. In addition, the waste stream is generally contam-
inated to some extent by a number of dissolved substances,
derived from the ore processed or from reagent additions in
the mill. Total concentrations of dissolved solids vary
but, except where leaching is practiced, rarely exceed 2,500
mg/1, with Ca-H-, Na+, K+, Mg++, C03—, and SO^ accounting
for nearly all dissolved materials. Heavy metals and other
notably toxic materials rarely exceed 10 mg/1 in the untreated
waste stream.
The volume of effluent from both mines and mills may be
strongly influenced by factors of topography and climate and
is frequently subject to seasonal fluctuations. In mines,
the water flow depends on the flow in natural aquifers inter-
cepted and may be highly variable. Water other than process
water enters the mill effluent stream primarily by way of
the tailing ponds (and/or settling ponds), which are almost
universally employed. These water contributions direct
precipitation on the pond, from runoff from surrounding areas,
or even from seepage and are only partially amenable to
elimination or control.
A number of operations or practices common to many
milling operations in this category involve the use of contact
process water and contribute to the waste-stream pollutant
load. These include ore washing, grinding, cycloning and
classification, ore and tail transport as a slurry, and the
use of wet dust-control methods (such as scrubbers). In terms
of pollutants contributed to the effluent stream, all of these
processes are essentially the same. Contact of water with
finely divided ore, gangue, or concentrates results in the
suspension of solids in the waste stream, and in the solution
of some ore constituents in the water. In general, total
levels of dissolved material resulting from these processes
are quite low, but specific substances (especially, some
heavy metals) may dissolve to a sufficient degree to require
treatment. These processes may also result in the presence
V-113
DRAFT
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DRAFT
of oil and grease from machinery in the wastewater stream.
Good housekeeping and maintenance practice should prevent
this contribution from becoming significant.
Ore roasting may be practiced as a part of some processing
schemes to alter physical or chemical properties of the ore.
In current practice, it is used to change magnetic properties
in iron-ore processing in the U.S. and in the past was used
to alter magnetic/electrostatic behavior of columbium and
tantalum ores. Roasting is also used in processing vanadium
ores to render vanadium values soluble. Although a dry process,
roasting generally entails the use of scrubbers for air
pollution control. Dissolved fumes and ore components rendered
soluble by roasting which are captured in the scrubber thus
become part of the waste stream. This scrubber water may
constitute an appreciable fraction of the total plant effluent
and may contribute significantly to the total pollutant load.
One mill surveyed contributes 0.8 ton of contaminated scrubber
bleed water per ton of ore processed.
Effluents from some ferroalloy mining and milling operations
are complicated by other operations performed on-site. Thus,
smelting and refining at one site, and chemical purification
•at another, contribute significantly to the wastewater gen-
erated at two current ferroalloy-ore processing plants. Since
waste streams are not segregated, and the other processes
involve wastes of somewhat different character then those
normally associated with ore mining and beneficiation, such
operations may pose special problems in effluent limitation
development.
An additional component of the mill waste stream at some
sites which is not related to the milling process is sewage.
The use of the mill tailing basin as a treatment location for
domestic wastes can result in unusually high levels of a
number of pollutants in the effluent stream, including NH3_,
COD, BOD, and TOG. At other sites, effluent from separate
domestic waste-treatment facilities may be combined with mine
or mill effluents, raising levels of NH3^, BOD, TOC, or
residual chlorine.
Sources of Wastes - Mine Effluents
Factors affecting pollution levels in mine water flows include:
(1) Contact with broken rock and dust within the mine,
resulting in suspended-solid and dissolved-ore
constituents.
V-114
DRAFT
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DRAFT
(2) Oxidation of reduced (especially, sulfide) ores,
producing acid and increased soluble material.
(3) Blasting decomposition products, resulting in NH3_,
and COD loads in the effluent.
(4) Machinery operation, resulting in oil and grease.
(5) Percolation of water through strata above the mine,
which may contribute dissolved materials not found
in the ore.
As discussed previously, variable (and, sometimes, very high)
flow rates are characteristic of mine discharges and can
strongly influence the economics of treatment. Data for mine
flows sampled in the development of these guidelines are
presented in Table V-38. Observed mine flows in the industry
range from zero to approximately 36 cubic meters (9,510
gallons) per minute. Generally, total levels of dissolved
solids are not great, ranging from 10 to 1400 ppm in untreated
mine waters. Total levels of some metals, however, can be
appreciable, as the data below, show for some maximum observed
levels (in mg/1) .
Al 9.4 Mo 0.5
Cu 3.8 Pb 0.19
Fe 17 Zn 0.47
Mn 5.5
In addition, oil and grease levels as high as 14 mg/1, and
COD values up to 91 mg/1, were observed. Since simple settling
treatment greatly reduces most of the above metal values,
it is concluded that most of metals present were contributed
in the form of suspended solids. There is no apparent
correlation between waste content or flow volume and production
for mine effluents.
Sources £f Wastes - Mill Effluents
Physical Processing Mill Effluents. In general, mills practicing
purely physical ore beneficiation yield a minimal set of
V-115
DRAFT
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DRAFT
TABLE V-38. CHEMICAL CHARACTERISTICS OF RAW MINE WATER IN
FERROALLOY INDUSTRY
MINE
6102
6103
6104
6107
PRODUCT
Mo. W
Mo
W. Mo
V
FLOW
(m /mm (apm)l
2.66 (700)
6.43 (1.700)
34.06 (9.000)
11.3613.000)
pH
4.5
7.0
6.5
7.3
CONCENTRATION (mg/£)
Oil and
Grsau
14
1.0
2.0
Nitrate
-
0.15
0.12
-
Fluor Ida
443
4.5
0.52
-
As
<0.01
<0.01
<0.07
<0.07
Cd
007
<0.01
<001
<0005
Cu
3.8
0.06
<0.02
<0.02
Mn
5.3
5.5
0.21
63
Mo
05
<01
<0.1
<01
Pb
0.06
0.19
0.14
•
V
<0.5
<0.5
<0.5
Zn
70
047
0X15
0.09
V-116
DRAFT
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DRAFT
pollutants. Separation in jigs, tables, spirals, etc.,
contributes to pollution in the same fashion as the general
practices of grinding and transport—that is, through contact
of ore and water. Suspended solids are the dominant waste
constituent, although, as in mine wastes, some dissolved
metals (particularly, those with high toxicity) may require
treatment. Roasting may be practiced in some future opera-
tions to alter magnetic properties of ores. As discussed
previously, this could change the effluent somewhat, by
increasing solubility of some ore components, and by introducing
water from scrubbers used for dust and fume control on
roasting ovens. Since solubilization is generally undesirable
in such operations, the very high total dissolved solid values
observed at mill 6107 are not anticipated elsewhere.
No sites in the ferroalloy category actually practicing
purely physical beneficiation of ore using water were visited
and sampled in developing these guidelines, since none could
be identified. A mine/mill/smelter complex recovering nickel
(mill 6106) which was visited, however, produces an effluent
which is felt to be somewhat representative, since water
contacts ore in belt washing—and gangue in slag granulation—
operations at that site. Raw waste data for that operation
illustrate the generally low level of dissolved materials
in effluents from these operations. In general, these effluents
pose no major treatment problems and are generally suitable
for recycle to the process after minimal treatment to remove
suspended solids.
Flotation Mill Effluents. The practice of flotation adds a
wide variety of process reagents, including acids and bases,
toxicants (such as cyanide), oils and greases, surfactants,
and complex organics (including amines and xanthates). In
addition to finer grinding of ore than for physical separation,
and modified pH, the presence of reagents may increase the
degree of solution of ore components.
Flotation reagents pose particular problems in effluent limit-
ation and treatment. Many are complex organics used in small
quantities, whose fates and effects when released to the
environment are uncertain. Even their analysis is not simple
(References 26 and 27). Historically, effluent data are
widely available only for cyanide among the many flotation
reagents employed. Similarly, in the guideline-development
V-117
DRAFT
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DRAFT
effort, analyses were not performed for each of the specific
reagents used at the various flotation mills visited. The
presence of flotation reagents in appreciable quantities
may be detected in elevated values for COD, oil and grease,
or surfactants, as analytical data on mill effluents indicate.
The limitation of reagents individually appears unfeasible,
since the exact suite of reagents and dosages is nearly
unique to each operation and highly variable over time.
Current practice in the ferroalloy milling industry includes
flotation of sulfide ores of molybdenum, and flotation of
scheelite (tungsten ore). The ores floated are generally
somewhat complex, containing pyrite and minor amounts of lead
and copper sulfides. Reagents used in the sulfide flotation
circuits and reflected in effluents include xanthates, light
oils, and cyanide (as a depressant). Since the flotation
is performed at basic pH, solution of most metals is at a low
level. Molybdenum is an exception in that it is soluble
as the molybdate anion in basic solution and appears in
significant quantities in effluents from several operations.
Tungsten ore flotation involves the use of a quite different
set of reagents—notably, oleic acid and tall oil soaps—and
may be performed at acid pH. At one major plant, both sulfide
flotation for molybdenum recovery and scheelite flotation
are practiced, resulting in the appearance of both sets of
reagents in the effluent. Visit sites included plants recovering
both molybdenum (6101, 6102, and 6103) and tungsten (6104
and 6105) by flotation. Although flotation would almost
certainly be used in such cases, no currently active processors
of sulfide ores of nickel or cobalt could be Identified in
the U.S.
Ore Leaching• In many ways, ore leaching operations maximize
the pollution potential from ore beneficiation. Reagents are
used in large quantities and are frequently not recovered.
Extremes of pH are created in the process stream and generally
appear in the mill effluent. Techniques for dissolving the
material to be recovered are generally not specific, and
other dissolved materials are rejected to the waste stream
to preserve product purity. The solution of significant
fractions of feed ore, and the use of large quantities of
reagents, results in extremely high total-dissolved-solids
concentrations. Because of reagent costs, and the benefits
of increased concentration in the precipitation or extraction
of values from solution, the amount of water used per ton
V-118
DRAFT
-------
DRAFT
of ore processed by leaching is generally lower than that
for physical benefication or flotation. One ton of water
per ton of ore is a representative value.
Effluents for several mills in the ferroalloy industry which
employ leaching were characterized in this study. Visit
sites included a vanadium mill (mill 6107)(properly classed
in SIC 1094, but treated here because of lack of radioactives,
end use of product, and applicability of general process to
other ferroalloy ores) which practices leaching as the primary
technique for recovering values from ores, as well as two
tungsten mills which employ leaching in the process, though
not as the primary beneficiation procedure. One operation
(mill 6105) leaches a small amount of concentrate to reduce
lime and phosphorus content, and the other (mill 6104) leaches
scheelite flotation concentrates as part of a chemical refining
procedure. Data for samples from leaching plants in the
uranium and copper industries may also be examined for compar-
ison.
Process Descriptions and Raw-Waste Characterizations For
Specific Mines and Mills Visited
Mine/Mill 6101
At mine/mill 6101, molybdenum ore of approximately 0.2 percent
grade is mined by open-pit methods and is concentrated by
flotation to yield a 90 percent molybdenite concentrate.
The mine and mill are located in mountainous terrain, along
a river gorge. The mill is adjacent to and below the mine,
the elevation of which ranges from 2,550 meters (8,400 ft)
to 3,000 meters (10,000 ft) above MSL (mean sea level).
The local climate is dry, with annual precipitation amounting
to 28 cm (11 in.) and annual evaporation of 107 cm (42 in.).
Approximately 22,000 cubic meters (6 million gallons) of
water per day are used in processing 14,500 metric tons
(16,000 short tons) of ore. Reclamation of 10 percent of
the water at the mill site, evaporation, and retention in
tails reduce the daily discharge of water to 16,000 cubic
meters (4.3 million gallons). Process water is drawn from
wells on the property and from the nearby river. No mine
water is produced.
V-119
DRAFT
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DRAFT
Ore processing consists of crushing, grinding, and multiple
stages of froth flotation, followed by dewatering and drying
of concentrates. The complete process is illustrated in the
simplified flowsheet of Figure V-25. There are no recoverable
byproducts in the ore. Reagent use is summarized in Table
V-39.
Recovery of molybdenite averages 78 to 80 percent but varies
somewhat, depending on the ore fed to the mill. Recoveries
on ore which has been stockpiled are somewhat lower than
those achieved on fresh ore. This is, apparently, due to
partial oxidation of the molybdenite to (soluble) molybdenum
oxide and ferrimolybdite, which are not amenable to flotation.
Processing of these oxidized ores is also accompanied by an
increase in the dissolved molybdenum content of the plant
discharge. The final concentrate produced averages 90
percent
As the flowsheet shows, only one waste stream is produced.
Data for this stream, as sampled at the mill prior to any
treatment, are summarized in Table V-40.
High COD levels apparently result from the flotation reagents
used and provide some indication of their presence. The low
cyanide level found reflects significant decreases in cyanide
dosage over earlier operating modes and indicates almost
complete consumption of applied cyanide. Metal analyses
were performed in acidified samples containing the solid
tailings. High values may be largely attributed to metals
which were solubilized from the unacidified waste stream.
Mine /Mill 6102
Mill 6102 also recovers molybdenite by flotation, but mill
processing is complicated by the additional recovery of by-
product concentrates. Water use in processing approximately
39,000 metric tons (43,000 short tons) of ore per day amounts
to 90,000 cubic meters (25 million gallons) per day. Nearly
complete recycle of process water results in the daily use of
only 1,700 cubic meters (450,000 gallons) of makeup water.
Discharge from the mill tailing basin occurs only during spring
snow-melt runoff, when it averages as much as 140,000 cubic
meters (38.5 million gallons) per day.
V-120
DRAFT
-------
DRAFT
Figure V-25. MILL 6601 FLOWSHEET
CONCENTRATE-
•— MIDDLINGS •
SCAVENGER
FLOAT
14 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
-CONCENTRATE -
—MIDDLINGS-
CLEANER
FLOAT
(8 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
-TAILS-
TAILS
i
CONCENTRATE
OVERFLOW—
-UNDERFLOW
^UNDERFLOW
OVERFLOW
i
^UNDERFLOW
22.000 m3/
-------
DRAFT
TABLE V-39. REAGENT USE IN MOLYBDENUM MILL 6101
REAGENT
Lime
Vapor Oil
Pine Oil
Hypo (Sodium Thiosulfate)
(Na2 $203 • 5H20)
Phosphorus Pentasulfide (?2 SB)
MIBC (methyl-isobutyl carbinol)
Sodium Cyanide (Na CN)
DOSAGE
g/ metric ton ore
0.075
0.09
0.015
0.035
0.005
0.02
0.015
Ib/short
ton ore
0.15
0.18
0.03
0.07
0.01
0.04
0.03
TABLE V-40. RAW WASTE CHARACTERIZATION AND
RAW WASTE LOAD FOR MILL 6601
PARAMETER
TSS
TDS
Oil and Grease
COD
A*
Cd
Cu
Mn
Mo
Pb
Zn
fm
Total Cyanida
Fluonda
CONCENTRATION
(mat I )
IN WASTE WATER
500.000
2398
2.0
135
0.01
074
51
56.6
5.3
98
76.9
1.305
002
62
TOTAL WASTE
kg/day
14.000.000
42.000
32
2.200
016
12
820
900
85
160
1.200
21,000
032
99
Ib/day
32.000.000
92.000
70
4.800
0.35
26
14OO
2.000
190
350
2.600
46.000
070
220
RAW WASTE LOAD
par unit ore milled
kg/ metric ton
995
30
OO023
016
0000012
0.00086
0059
0064
0.0061
0011
0086
15
0000023
00071
Ib/short ton
1390
60
00046
032
0000023
00017
Oil
0.13
0012
0023
017
30
0000046
0014
par unit concentrate produced
kg/ metric ton
610.000
1330
1 4
96
0.0070
0.52
36
39
37
70
524
915
0014
43
Ib/short ton
1.200.000
3.670
28
190
0014
10
72
79
74
140
105
-1.830
0028
87
V-122
DRAFT
-------
DRAFT
Mining is both underground and open-pit, with underground
operations which began approximately 67 years ago, and the
first open-pit production in 1973. Recovery of molybdenite
is by flotation in five stages, yielding a final molybdenite
concentrate containing more than 93 percent MoS2^. Tungsten
and tin concentrates are produced by gravity and magnetic
separation, with additional flotation steps used to remove
pyrite and monazite. Recovered pyrite is sold as possible
(currently, about 20 percent of production), with the balance
delivered to tails. The monazite float product also reports
to the tailing pond, since recovery of monazite is not
profitable for this operation at this time.
The mill operation is located on the continental divide at
over 3,353 meters (11,000 feet) above MSL. The local terrain
is mountainous. Climate and topography have a major impact
on water-management and tailing-disposal practices, with a
heavy snow-melt runoff and the presence of major drainages
above tailing-pond areas posing problems.
Mill Description. Figure V-26 presents a greatly simplified
diagram of the flow of ore through the mill. Following crush-
ing and grinding, roughing and scavenging flotation are used
to extract molybdenite from the ore. Nearly 97 percent of
the incoming material—currently, about 39,000 metric tons
(43,000 short tons) per day—is thereby rejected and sent
directly to the byproduct recovery plant. The flotation
concentrate, averaging about 10 percent MoS2^ is fed to four
stages of further flotation. Reagents used in the primary
flotation step are summarized in Table V-41. Most are added
as the ore is fed to the ball mills for grinding.
Cleaner flotation in four stages and three regrinds yield a
final product averaging greater than 93 percent Moj32 content.
Reagent use in the cleaner grinding and flotation circuit is
summarized in Table V-A2.
Tailings from the rougher flotation are pumped to the by-
products plant, where heavy fractions are concentrated in
Humphreys spirals. Pyrite is removed from the concentrate
by flotation at pH 4.5, and the flotation tailings are then
tabled to further concentrate the heavy fractions. The pH
of the table concentrate is then adjusted to 1.5 and its temp-
erature raised to 70 degrees Celsius (158 degrees Fahrenheit),
and monazite is removed by flotation. The tailings from this
V-1123
DRAFT
-------
DRAFT
Figure V-26. SIMPLIFIED MILL FLOW DIAGRAM FOR MILL 6102
TO
TAILINGS
CRUSHING
(3 STAGES)
28% + 3 MESH
GRINDING IN
BALL MILLS
1
36% * 100 MESH
»|
1
FLOTATION
* I
CONCENTRATE |^ 1 FLOTATION
96% OF MILL FEED
CLEANER
FLOTATION +- T° .
(4 STAGES) TA"
1
« LIGHTS GRAVITY SEPARATION DRYING
~~ "~ (HUMPHREY'S SPIRALS) 1
*
PYRITE
FLOTATION
1
TAILS
© |
1
^ MONAZITE «- MONAZITE
~~ CONCENTRATE " FLOTATION
"*VV
MAGNETIC
SEPARATION
* *
f
CONCENTRATE
(93% + MoS2)
^_/Xx\ INDICATES I
V..X SAMPLING POINT \
„ 1
NONMAGNETIC MAGNETIC
TIN TUNGSTEN
CONCENTRATE CONCENTRATE
V-12A
DRAFT
-------
DRAFT
TABLE V-41. REAGENT USE FOR ROUGHER AND SCAVENGER
FLOTATION AT IV ILL 6102
REAGENT
Pine oil
Vapor oil
Syntax
Lima (Calcium oxide)
Sodium silicate
Nokes reagent
PURPOSE
Frother
Collector
Surfactant and Frother
Adjustment of pH to 8.0
Slime Dispersant
Lead Depressant
CONSUMPTION
kg/metric ton
ore milled
0.18
0.34
0.017
0.15
0.25
0.015
Ib/short ton
ore milled
0.35
0.67
0.034
0.30
0.50
0.03
TABLE V-42. REAGENT USE FOR CLEANER FLOTATION AT MILL 6102
REAGENT
Vapor oil
Sodium cyanide
Nokes reagent
Dowf roth 250
Valco 1801
PURPOSE
Collector
Pyrite and Chalco-
pyrite Depressant
Lead Depressant
Frother
Flocculant
CONSUMPTION
kg/metric ton
ore milled
0.45
0.13
0.45
0.015
0.003
Ib/short ton
ore milled
0.90
0.25
0.90
0.03
0.006
V-125
DRAFT
-------
DRAFT
flotation step are dewatered, dried, and fed to magnetic
separators, which yield separate tin (cassiterite) and tungsten
(wolframite) concentrates. Reagent use in the flotation of
pyrite and monazite is summarized in Table V-43.
Effluent samples were taken at three points in mill 6102
due to the complexity of the process. A combined tailing
sample was taken representative of the total plant effluent,
and, in addition, effluents were sampled from two points
in the process (marked 19 and 20 on the flowsheet, Figure
V-26). Although flows at these points are very small compared
to the total process flow, they were considered important
because of the acid conditions prevailing in monazite flota-
tion. Concentrations and total loadings in the mill effluent,
and concentrations in the effluents from pyrite flotation and
monazite flotation, are presented in Tables V-44 and V-45.
Considerably heavier use of cyanide than at mill 6101 (almost
ten times the dosage per ton of ore) is reflected in signifi-
cantly higher levels in the untreated mill waste. Total
metal contents are again elevated by leaching solid particles
in the tailing stream. The increase in solution of most
heavy metals as increasingly acid conditions prevail in
processing is evident in the data from the monazite and pyrite
flotation effluents.
Mine water is produced in the underground mine at mill 6102
at an average rate of 4,000 metric tons per day (700 gpm).
Its characteristics are summarized, along with those of other
mine waters, in Table V-38. At mill 6102, all mine water
is added to the mill tailing pond and then to the process
circuit.
Mine 6103
Mine 6103 is an underground molybdenum mine which is under
development. Ore from the mine will be processed in a mill
at a site approximately 16 kilometers (10 miles) from the mine
portal. The mill operation will produce no effluent, all
of the process water being recycled. Mine water flow presently
averages 9,800 cubic meters per day (1,700 gpm). Its quality
prior to treatment has been summarized in Table V-38.
V-126
DRAFT
-------
DRAFT
TABLE V-43. REAGENT USE AT BYPRODUCT PLANT OF MILL 6102
(Based on total byproduct plant feed)
REAGENT
I
PURPOSE
I
CONSUMPTION
kg/metric ton
ore milled
Ib/short ton
ore milled
PYRITE FLOTATION
Sulfuric acid
Z-3 Xanthate
Dowfroth 250
ARMAC C
Starch
Sulfuric acid
IpH Regulation
Collector
Frother
0.018
0.0005
0.0005
0.036
0.001
0.001
MONAZITE FLOTATION
Collector
WO2 Depressant
pH Regulation
0.0005
0.0005
0.0005
0.001
0.001
0.001
TABLE V-44. MILL 6102 EFFLUENT CHEMICAL CHARACTERISTICS
(COMBINED-TAILINGS SAMPLE)
PARAMETER
TSS
TDS
Oil and Grease
COD
As
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION
img/i) IN
WASTE WATER
150,000
2.254
4
23.8
<0.1
0.19
21.0
50
17.5
2.1
25.0
1.500
11.7
0.45
TOTAL WASTE
kg/day
200.000
360
2.100
<9
17
1.890
4.500
1.600
190
2.250
135.000
1.100
41
Ib/day
440,000
790
4,600
<20
37
4,200
9.900
3.500
418
4.950
300.000
2.400
90
RAW WASTE LOAD
per unit ore processed
kg/metric ton
998
4.7
0.0080
0.049
<0.0002
0.00040
0.047
0.10
0.037
0.0044
0.052
3.1
0.026
0.00095
Ib/shot ton
1996
9.3
0.016
0.098
O.0004
0.00080
0.088
0.21
0.074
0.0088
0.10
6.3
0.052
0.0019
per unit total
concentrate produced
kg/metric ton
2.700
4.6
28
<0.1
0.23
25
58
21
2.5
30
1,800
15
0.55
Ib/short ton
5,400
9.2
56
<0.2
0.46
50
120
43
5.0
60
3,600
30
1.1
V-L27
DRAFT
-------
DRAFT
TABLE V-45. CHEMICAL CHARACTERISTICS OF ACID-FLOTATION STEP
PARAMETER
PH
Cd
Cu
Fe
Mn
Mo
Pb
CONCENTRATION (mg/Jl ) AT INDICATED POINTS OF FIGURE V-26
PYRITE FLOAT (19)
45*
0.01
02
4.2
4.0
3.0
0.3
MONAZITE FLOAT (20)
1.5
0.042
0.5
490
53.3
4.0
1.34
•Value in pH units
V-128
DRAFT
-------
DRAFT
Mine/Mill 6104
This complex operation combines mining, beneficiation, and
chemical processing to produce a pure ammonium paratungstate
product as well as molybdenum and copper concentrates. A
total of 10,000 cubic meters (2.9 million gallons) of water
are used each day in processing 2,200 metric tons (2,425
short tons) of ore. The bulk of this water is derived from
the 47,000 cubic meters (13 million gallons) of water pumped
from the mine each day.
The mill process is illustrated in Figures V-27 and V-28,
which also show water flow rates. After crushing and grinding,
sulfides of copper and molybdenum are floated from the ore,
employing xanthate collectors and soda ash for pH modifica-
tion. This flotation product is separated into copper and
molybdenum concentrates in a subsequent flotation using sodium
bisulfide to depress the copper. Tailings from the sulfide
flotation are refloated using tall oil soap to recover a
scheelite concentrate, which is reground and mixed with
purchased concentrates from other sites. The scheelite is
digested and filtered, and the solution is treated for
molybdenum removal. Following solvent extraction and concen-
tration, ammonium paratungstate is crystallized out of
solution and dried.
Effluent streams from parts of the operation specifically
concerned with beneficiation were sampled and analyzed*
along with the combined discharge to tails for the complete
mill. Mine water was also sampled, and analyses have been
reported in Table V-38. Data for a composite effluent from
beneficiation operations, several individual beneficiation
effluents, and the combined plant discharge are presented
in Tables V-46, V-47, V-48, V-49, and V-50.
The combined-tails discharge characteristics are not truly
representative of raw waste from the leaching and chemical
processing parts of the operation, since advanced treatments
(including distillation and air stripping) are performed on
parts of the waste stream prior to discharge to tails. Total
dissolved solids and ammonia (not determined for the sample
taken), in particular, are greatly reduced by these treatments.
V-129
DRAFT
-------
DRAFT
Figure V-27. INTERNAL WATER FLOW FOR MILL 6104 THROUGH
MOLYBDENUM SEPARATION
WET
ORE
WATER
FROM
CREEK
121
(32
m3Ml
0000
MTen
1
<" CRUt
II— *• Al
GRIN
C781000gpd)
1
" nan
HINQ O14JC
WHO
40Sn3Aky
(131MO gpd)
^ 1
180 0*1 BULK MILFIDS
"^ FLOTATION
4KB ml/d«y 1121.000 0d)
WATER
FROM
MINE
'
f
IOMETER
OO-FTI
THICKENER
i/wzni-/an
MILOOOgpd
(29X100 gpd)
COPPER/
MOLYBDENUM
SEPARATION
. 1*082 PRODUCT
1 OJOSB m*jtov
1 (lOgpdl
TO STOCKPILE
110 IB3J^taV
4 1
1071
^ OCMEELITE O7'°
^ THICKENER
STI
(127400 gpd)
BLOW
46«il
I121J
AM ATMO
ERI
K7WN
23m3Aky
(eunogpd
t
,
»3/*w
00 gpdl
TO ^
IPHERE
B4B m /dw
(144.000 gpd)
ITIOM
'
DRVINQ
ROASTINO .- _
*1
L
toO, PRODUCT
TO
STOCKPILE
~|Mn3M«
|(700gpd) .
«.
SCRUBBER
,
U3Sm3Ahv
(360000 vdl
FLOTATION
1
UNDEI
4.380 m*/*y
lt.1Bl.000 B>d)
IFLOW
r
402 «3Atay (130400 gpdl
CONCEI
PER
ITRATE
ENER
96n3AHv
OSJOOgpdl
Cu CONCENTRATE PRODUCT
TO STOCKPILE
114 m3Aby
(30.000 gpd)
FILTI
WAS!
299m3/
(79.000
RINO
ID
KINO
290m3/
179,000
MOLYBDENUM
r SEPARATION
m.tae^ioogpd)
TO
TAILING POND
chv
gpd)
1.106 m'/dly
(292400 gpd)
dw
gpd)
400M3Mw
dOBjOOOtpd)
TO
' S> SOLVENT
EXTRACTION
(FIGURE V-2BI
V-130
DRAFT
-------
DRAFT
Figure V-28. INTERNAL WATER FLOW FOR MILL 6104
FOLLOWING MOLYBDENUM SEPARATION
TO ATMOSPHERE
FROM MOLYBDENUM
SEPARATION
(FIGURE V 27)
TO BOILERS
V-131
DRAFT
-------
DRAFT
TABLE V-46. COMPOSITE WASTE CHARACTERISTICS FOR BENEFICIATION
AT MILL 6104 (SAMPLES 6, 8, 9, AND 11)
PARAMETER
PH
COO
Oil and Create
At
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
Cyanide
CONCENTRATION
(mg/e) IN
WASTEWATER
10'
238
11 4
<0.07
004
49
22.5
190
0.22
6.3
4.8
02
TOTAL WASTE
kg/day
1.100
55
<034
0 19
24
110
91
1 1
30
23
096
It/day
2.400
120
<075
042
53
240
200
24
66
51
21
RAW WASTE LOAD
per unit ore processed
kg/metric ton
0.50
0.025
< 0.0002
0000086
0011
0.050
0041
000050
0014
0010
000044
Ib/short ton
10
0.050
<00003
000017
0022
010
0083
00010
0027
0021
000088
per unit total
concentrate produced
kg/metric ton
81
041
<0003
00014
0.18
081
067
00081
023
016
00072
Ib/short ton
16
0.81
<0007
0.0028
036
16
1 3
0016
046
0.32
0014
•Value in pH units
TABLE V-47. WASTE CHARACTERISTICS FROM COPPER-THICKENER OVERFLOW
FOR MILL 6104 (SAMPLE 5)
PARAMETER
PH
CO
Cu
Mn
Mo
Pb
Fe
CONCENTRATION
(mg/JJ) IN
WASTEWATER
IT
0.26
< 0.002
0.091
0.11
0.0064
2.4
Ib/day
_
0.053
<0.004
0.20
0.24
0.014
5.3
RAW WASTE LOAD
per unit ore milled
kg/metric ton
_
0.000011
<0.0000009
0.000041
0.000050
0.0000029
0.0011
Ib/short ton
0.000022
< 0.000002
0.000082
0.00010
0.0000058
0.0022
•Value in pH units
V-132
DRAFT
-------
DRAFT
TABLE V-48. SCHEELITE-FLOTATION TAILING WASTE CHARACTERISTICS
AND LOADING FOR MILL 6104 (SAMPLE 7)
PARAMETER
PH
Cd
Cu
Mn
Mo
Pb
Zn
Fe
CONCENTRATION
(mg/£) IN
WASTEWATER
10"
0.32
1.42
41
1.3
0.22
11.2
0.43
TOTAL WASTE
kg/day
—
1.3
5.9
170
5.5
.92
47
1.8
Ib/day
—
2.9
13
370
12
2.0
100
4.0
RAW WASTE LOAD
per unit ore mil led
kg/metric ton
—
0.00059
0.0027
0.077
0.0025
0.00042
0.021
0.00082
Ib/short ton
_
0.0012
0.0054
0.15
0.0050
0.00084
0.043
0.0016
•Value in pH units
TABLE V-49. 50-FOOT-THICKENER OVERFLOW FOR MILL 6104 (SAMPLE 10)
PARAMETER
PH
Cd
Cu
Mn
Mo
Pb
Zn
Fe
CONCENTRATION
(mg/i) IN
WASTEWATER
9"
<0.01
0.31
1.3
21.0
0.04
0.16
7.7
TOTAL WASTE
kg/day
—
< 0.005
0.15
0.61
9.9
0.019
0.075
3.6
Ib/day
—
<0.01
0.33
1.3
22
0.042
0.17
7.9
RAW WASTE LOAD
per unit ore milled
kg/metric ton
—
< 0.000002
0.000068
0.00028
0.0045
0.0000086
0.000034
0.0016
Ib/short ton
—
< 0.000005
0.00014
0.00055
0.0090
0.000017
0.000068
0.0033
•Value in pH units
V-133
DRAFT
-------
DRAFT
TABLE V-50. WASTE CHARACTERISTICS OF COMBINED-TAILING
DISCHARGE FOR MILL 6104 (SAMPLES 15,16, AND 17)
PARAMETER
TDS
Oil and Grease
COD
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Total Cyanide
CONCENTRATION
(mg/i) IN
WASTE WATER
2290
14.7
174
<0.07
0.03
0.03
0.52
50
2.2
<0.02
<0.5
< 0.01
TOTAL WASTE
kg/day
22.900
147
1.740
<0.7
0.30
0.30
5.2
500
22
<0.2
<5.0
<0.1
Ib/day
50.000
320
3.800
<1 5
0.66
0.66
11
1.100
480
< 0.4
<11
< 0.2
RAW WASTE LOAD
per unit ore processed
kg/metric ton
10.4
0.067
0.79
<0.0003
0.00014
0.00014
0.0024
0.23
0.010
< 0.00009
< 0.002
< 0.00005
Ib/short ton
21
0.13
1.6
< 0.0006
0.00027
0.00027
0.0047
0.45
0.020
< 0.0002
< 0.005
< 0.00009
per unit
concentrate produced
kg/metric ton
170
1.1
13
< 0.005
0.0023
0.0023
0.039
3.7
0.16
< 0.0015
<0.03
< 0.0008
tb/thort ton
340
2.2
26
<0.01
0.0046
0.0046
0.078
7.4
0.32
< 0.003
<007
< 0.002
V-134
DRAFT
-------
DRAFT
Mine/Mill 6105
Mill 6105, a considerably smaller operation than mine/mill
6104, also recovers scheelite. As shown In the mill flowsheet
of Figure 111-18, a combination of sulfide flotation, scheelite
flotation, wet gravity separation, and leaching is employed
to produce a 65 percent tungsten concentrate from 0.7 percent
mill feed. A total of 52 metric tons (57 short tons) per day
of water drawn from a well on site are used in processing
46 metric tons (51 short tons) of ore. Mill tailings are
combined prior to discharge, providing neutralization of
acid-leach residues by the high lime content of the ore.
Analytical data for a sample of the combined mill effluent
are presented in Table V-51.
The mine at this site intercepts an aquifer producing mine
water, which must be intermittently pumped out (for approxi-
mately % hour every 12 hours). Total effluent volume is
less than 4 cubic meters (1,000 gallons) per day. Samples
of this effluent were not obtained because of inactivity
during the site visit. It is expected to be essentially the
same as the mill water-source well, which drains the same
aquifer and which was sampled.
Mine/Mill 6106
Ferronickel is produced at this site by direct smelting of
a silicate ore (garnierlte) from an open-pit mine. Water
use is limited and is primarily involved in smelting, where
it is used for cooling and for slag granulation. Beneficia-
tion of the ore involves drying, screening, roasting, and
calcining but requires water for belt washing and for use
in wet scrubbers. Flow from all uses combined amounts to
approximately 28 cubic meters (7,700 gallons) per day.
This combined waste stream was sampled, and its analysis is
shown in Table V-52.
Mine water during wet-weather runoff through a creek bed to
an Impoundment used for mill water treatment results in
discharges as large as 21,000 cubic meters (576,000 gallons)
per day from the Impoundment. Since the mine was dry during
the site visit, no samples of this flow were obtained.
Company-furnished data for the Impoundment water quality,
however, reflect the impact of mine-site runoff.
V-135
DRAFT
-------
DRAFT
TABLE V-51. WASTE CHARACTERISTICS AND RAW WASTE LOAD AT MILL 6105
(SAMPLE 19)
PARAMETER
TDS
Oil and Grease
COD
NH3
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION
(mg/£) IN
WASTEWATER
1232
1
39.7
1.4
<0.07
<0.01
0.02
0.52
0.19
0.5
0.02
<0.5
<0.02
0.44
6.9
<0.01
TOTAL WASTE
kg/day
64
0.052
2.1
0.073
< 0.004
< 0.0005
0.0010
0.027
0.0099
0.026
0.0010
<0.03
< 0.001
0.023
0.36
< 0.0005
Ib/day
140
0.11
4.6
0.16
<0.01
< 0.001
0.0022
0.059
0.022
0.057
0.0022
<0.07
< 0.002
0.051
0.79
< 0.001
RAW WASTE LOAD
per unit ore processed
kg/ metric ton
1.4
0.0011
0.046
0.0015
C0.0001
<0.00001
0.000022
0.00058
0.00022
0.00057
0.000022
<0.0007
<0.00002
0.00050
0.0078
<0.00001
Ib/short ton
2.8
0.0022
0.092
0.0030
<0.0002
<0.00002
0.000045
0.0012
0.00043
0.0011
0.000045
<0.001
<0.00004
0.0010
0.016
< 0.00002
per unit total
concentrate produced
kg/metric ton
130
0.10
4.2
0.14
< 0.009
< 0.0009
0.002
0.053
0.020
0.052
0.0020
<0.06
< 0.002
0.045
0.71
< 0.0009
Ib/short ton
250
0.20
8.4
0.28
<0.02
< 0.002
0.010
0.11
0.040
0.10
0.010
<0.13
< 0.004
0.091
1.4
< 0.002
TABLE V-52. CHEMICAL COMPOSITION OF WASTEWATER, TOTAL WASTE, AND
RAW WASTE LOADING FROM MILLING AND SMELTER EFFLUENT
FOR MILL 6106
PARAMETER
PH
TSS
TDS
Oil and grease
As
Cd
Cu
Ml
Mo
Pb
Zn
Fe
Nl
CONCENTRATION
(mall)
IN WASTEWATER
8.6*
226.9
212
3.4
< 0.07
< 0.005
< 0.03
0.53
0.5
<0.1
0.06
24
0.4
TOTAL WASTE
kg/day
3.600
3.300
54
<1
<0.08
<2
a 79
380
6.3
Ib/day
7.900
7.300
120
< 2
<0.2
<1
18
17
<4
1.7
840
13.9
RAW WASTE LOAD
per unit ore milled
kg/1000
metric tons
790
730
12
<0.2
<0.02
<0.1
1.8
1.7
< 04
0.17
84
1.4
lb/1000 short ton*
1.600
1.6OO
24
<0.4
<0.04
< 0.2
3.7
3.5
<0.9
0.3S
170
2.8
par unit concentrate produced
kg/1000
metric tons
43.000
39.000
640
00
< 1
< 6
99
94
<20
9.4
4.500
75
Ib/IOOO short tons
86.000
79.000
1.300
<20
< 2
<10
200
190
<50
19
9.000
150
•Value in pH units.
V-136
DRAFT
-------
DRAFT
Mine/Mill 6107
At this operation, vanadium pentoxide, V205_, is produced from
an open-pit mine by a complex hydrometallurgical process
involving roasting, leaching, solvent extraction, and precipi-
tation. The process is illustrated in Figure 111-21 and also
in Figure V-29 (which shows system water flows). In the
mill, a total of 7,600 cubic meters (1.9 million gallons)
of water are used in processing 1,140 metric tons (1,250
short tons) of ore, including scrubber and cooling wastes and
domestic use.
Ore from the mine is ground, mixed with salt, and palletized.
Following roasting at 850 degrees Celsius (1562 degrees
Fahrenheit) to convert the vanadium values to soluble sodium
vanadate, the ore is leached and the solutions acidified to
a pH of 2.5 to 3.5. The resulting sodium decavanadate
(Ha6yi002_8) is concentrated by solvent extraction, and ammonia
Is added to precipitate ammonium vanadate, which Is dried
and calcined to yield a V205_ product.
The most significant effluent streams are from leaching and
solvent extraction, from wet scrubbers on roasters, and from
ore dryers. Together, these sources account for nearly 70
percent of the effluent stream, and essentially all of its
pollutant content. Analyses for these waste streams are
summarized in Tables V-53, V-54, and V-55. Effluents from
the solvent-extraction and leaching processes are currently
segregated from the roaster/scrubber effluent, although they
are both discharged at the same point, to avoid the genera-
tion of voluminous calcium sulfate precipitates from the
extremely high sulfate level in the SX stream and the high
calcium level in the scrubber bleed. Both of these waste
streams exhibit extremely high dissolved-solid concentra-
tions (over 20,000 mg/1) and are diluted approximately 10:1
immediately prior to discharge.
Mercury Ores
Water flow and the sources, nature, and quantity of the wastes
dissolved in the water during the processes of mercury-ore
mining and beneficlation are described in this section.
V-137
DRAFT
-------
Figure V-29. WATER USE AND WASTE SOURCES FOR VANADIUM MILL 6107
O
3)
f
oo
CI ""^>
.J*.
UjOOOgpT)
©1
t
4^20
MJO
?*)
i
' T
)«n)
f
*
SOLVENT
EXTRACTION
COOLING
1
U14£/mm 378tMn 114t/min 189 I/mm 871 C/mui IBSt/min
<400gpm) (IMgpml (30 gpm) (Mgpm) 1230 gpm) T(J (50 gpm)
1 i i I i AT"E |
WASHING. n_v.
WATER LEACH. AND ai™
SOLVENT EXTRACTION SCHUBI
SOLVENT , ,
ksKIM X
PONO^
1 RAIN 1 , (*)
114 I/mm AMMONIA
(30 gpm) TREATMENT
^U_lr —
^ — *->v
f 189^50.000-1(50.000.000^11) \
( WEST )
X, EFFLUENT POND ^^/
I ZJMUmm
1610 gpm)
, 3fi
NOTE "
RUNOFF FROM RAIN
IS NOT CONSIDERED
EXCEPT WHERE IT
ENTERS THE PROCESS
R SANITARV MISC ROASTER
IER USES USES OFF-GAS
SCRUBBER
*
SEWAGE
©TREATMENT
___^^
,__JL:
©
189
(SO
J MISC
j EVAPORATION UStS
t
®"
«P"I FROM S£~r
TAILING f "/.
DAM I **J
/"3.7aB.OOO-t(1.000.000«l)\ Xl1^66.000-t(3.000.000^ir\ 1 ^^~
l HOLDING J ( SCRUBBER ) 303 (/mm
X^^^ POND ^^r ^^^aLttD fOND^^S ISO gpm)
rt»rr — T "^ — i
j SOLIDS | j SOLIDS |
(M)
1
1
^~
f DIVERSION
•^N
B30l/mn
(140gpcn)
2^70t/min T 5301/mm
(800 gpm) 1140 gpm)
1 RECYCLE
"SJ,1" WATER
MISC USES TOWER
I 1.136 (/n*,
1300 gpm)
1 — '
_^ 1.13B(/min
AN>\ '*" ;•""'
FER W-l
wJ
CD) VERSION
J.OND__^
;
jT jT (^^.ff^^v— a^-n^
RECYCLE WA^E X^J^UENT POND^X ' '
^ ,- U49(/nitn
J* * (330 gpm)
40 gpm)
F®
O
3)
(»«) - SAMPLE NUMBER
SAMPLES (TOAND ftt^ARE
MINE WATER SAMPLES
-------
DRAFT
TABLE V-53. WASTE CHARACTERIZATION AND RAW WASTE LOAD FOR MILL 6107
LEACH AND SOLVENT-EXTRACTION EFFLUENT (SAMPLE 80)
PARAMETER
pH
TOS
Oil ind green
COO
NH3
Ai
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fa
Ca
Chloride
Fluoride
Sulfate
CONCENTRATION
(mo/ III
INWASTEWATER
3.5-
39.350
94
475
0.16
035
0.037
1.15
016
54
< 0.1
< 0.05
31
052
0.26
206
7.900
4.6
26.000
TOTAL WASTE
kg/day
-
83.000
200
1.000
0.34
074
0.078
2.4
032
110
<0.2
< 0.1
65
1 1
055
430
17.000
9.7
56.000
Ib/day
-
180.000
440
2.200
0.75
1.6
0.17
5.3
07
240
< 0.4
< 0.2
140
24
1.2
950
37.000
21
120.000
RAW WASTE LOAD
par unit ore milled
ke> metric ton
-
73
0.18
0.88
0.0003
0.00065
0.000068
0.0021
000028
O.OS6
< 00002
< 0.0001
0057
0.00096
00006
0.38
15
0.0085
48
Ib/ihoft ton
-
146
0.35
176
00006
0.0013
000014
0.0042
000056
0.19
< 0.0004
< 00002
Oil
0.0019
0001
0.75
30
0017
96
per unit concentrate produced
ho/metric ton
-
6.570
16
79
0.027
0.059
0.0061
0.19
0.025
8.6
<002
<0.01
5.1
0.086
0045
34
1.350
077
4.320
Ib/thon ton
-
13.100
32
160
0054
0.12
0.012
038
005
17
<0.04
<0.02
10
0.17
009
68
2.700
1 5
8.640
•Value in pH umti
V-139
DRAFT
-------
DRAFT
TABLE V-54. WASTE CHARACTERISTICS AND WASTE LOAD FOR DRYER
SCRUBBER BLEED AT MILL 6107 (SAMPLE 81)
PARAMETER
PH
TSS
TDS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Ca
Chloride
Fluoride
Sulfate
CONCENTRATION
(mg/U
IN WASTEWATER
7.8'
—
7,624
15
58.4
2
<0.07
< 0.005
0.25
0.06
4
<0.1
<0.05
29
0.33
27
118
4,220
1.35
255
TOTAL WASTE
kg/day
_
—
4,000
7.8
30.4
1.0
< 0.035
< 0.0025
0.13
0.03
2.1
<0.05
< 0.025
15
0.17
14
61
2,200
0.70
133
Ib/day
— _
8,800
17
67
2.2
<0.07
< 0.005
0.29
0.07
4.6
C0.1
<0.05
33
0.37
31
130
4,800
1.5
290
RAW WASTE LOAD
per unit ore milled
kg/metric ton
_
_
3.5
0.007
0.027
0.0009
<0.00003
<0.000002
0.00011
0.00003
0.0018
<0.00004
<0.00002
0.013
0.00015
0.012
0.054
1.9
0.0006
0.12
Ib/short ton
^
7.0
0.014
0.054
0.0018
< 0.00006
< 0.000004
0.00023
0.00006
0.0037
< 0.00009
< 0.00004
0.026
0.00030
0.025
0.11
3.9
0.0012
0.23
•Value in pH units
V-140
DRAFT
-------
DRAFT
TABLE V-55. WASTE CHARACTERISTICS AND LOADING FOR SALT-ROAST
SCRUBBER BLEED AT MILL 6107 (SAMPLE 77)
PARAMETER
PH
TSS
TDS
Oil and Grease
COD
Ammonia
As
Gd
Cr
Cu
Mn
Mo
Pb
V
Zn
Ca
Chloride
Fluoride
Sutfate
CONCENTRATION
(mg/£)
IN WASTEWATER
2.3*
2,000
80.768
5
1,844
0.04
0.08
< 0.006
0.9
<0.03
5.5
-
<0.05
-
< 0.003
78,400
59.500
7.5
780
TOTAL WASTE
kg/day
-
2,400
97,000
6.0
2,200
0.05
0.096
< 0.006
1.1
<0.04
6.6
—
<0.06
—
< 0.004
94,000
65,000
9.0
940
Ib/day
—
5,200
210,000
13
4300
0.11
0.21
<0.01
2.4
<0.09
15
-
< 0.009
210,000
140.000
20
2,100
RAW WASTE LOAD
per unit ore milled
kg/metric ton
—
2.1
85
0.0053
1.9
0.00004
0.00008
< 0.000005
0.00096
< 0.00004
0.0058
—
< 0.00005
—
< 0.000004
82
57
0.0079
0.82
Ib/short ton
—
4.2
170
0.011
3.9
0.00009
0.00017
< 0.00001
0.0019
< 0.00008
0.012
-
< 0.0001
-
< 0.000008
160
110
0.016
1.6
•Value in pH units
V-141
DRAFT
-------
DRAFT
Water Uses
Historically, water has had only limited use in the mercury-
ore milling industry. This is primarily because little,
if any, beneficiation of mercury ore is accomplished prior
to roasting the ore for recovery of mercury. In the past,
mercury ore was typically only crushed and/or ground to pro-
vide a properly sized kiln or furnace feed. However, because
high-grade ores are nearly depleted at present, lower-grade
ores are being mined, and beneficiation is becoming more
Important as a result of the need for a more concentrated
furnace or kiln feed.
Currently in the United States, one small operation (mine/mill
9201) is using gravity methods to concentrate mercury ore.
In addition, a large operation (mill 9202), due to open
during 1975, will employ a flotation process to concentrate
mercury ore. In both of these processes, water is a primary
material and is required for the process operating conditions.
Water is the medium in which the fine and heavy particles are
separated by gravity methods. In the flotation process,
water is introduced at the ore grinding stage to produce a
slurry which is amenable to pumping, sluicing, and/or classi-
fication for sizing and feed into the concentration process.
Water is not used in mercury mining operations and is dis-
charged, where it collects, only as an indirect result of a
mining operation. This water normally results from ground-
water seepage but may also include some precipitation and
runoff.
Water flows of the flotation mill and the operation employing
gravity beneficiation methods are presented in Figure V-30.
Sources of Wastes
There are two basic sources of effluents containing pollutants:
those from mines and the beneficiation process. Mines may
be either open-pit or underground operations. In the case of
an open pit, the source of the pit discharge, if any, is
precipitation, runoff and ground-water seepage into the pit.
Ground-water seepage is the primary source of water in under-
ground mines. However, in some cases, sands removed from
mill tailings are used to backfill stopes. These sands may
initially contain 30 to 60 percent moisture, and this water
may constitute a major portion of the mine effluent.
V-142
DRAFT
-------
DRAFT
Figure V-30. WATER FLOW IN MERCURY MILLS 9101 AND 9102
(NO DISCHARGE)
164m*Atav
(4.320 gpd)
GRAVITY-
MILL
t
^f TAILIMR
^V POMD
x^_
1.649 m3 diy (432.000 gpd)
(a) MINE/MILL 9201
(NO DISCHARGE!
3.8 m'/min 11.000 jpml
•DUE TO BEGIN OPERATION IN 1975
(b> MINE/MILL 9202
MILL
(NO DISCHARGE WATER NOT USED
9ENEFICIATION LIMITED TO
CRUSHING AND/OR GRINDING TO
PROVIDE FURNACE FEED.)
(e) OTHER MERCURY OPERATIONS
V-143
DRAFT
-------
DRAFT
The particular waste constituents present in a mine or mill
discharge are a function of the mineralogy and geology of the
ore body and the particular milling process employed, if any.
The rate and extent to which the minerals in an ore body
become solubilized are normally increased by a mining opera-
tion, due to the exposure of sulfide minerals and their
subsequent oxidization to sulfuric acid. At acid pH, the
potential for solubilization of most heavy metals is greatly
increased.
Wastewater emanating from mercury mills consists almost
entirely of process water. High suspended-solid loadings
are the most characteristic waste constituent of a mercury
mill waste stream. This is primarily due to the necessity
for fine grinding of the ore to make it amenable to a parti-
cular beneflciation process. In addition, the increased
surface area of the ground ore enhances the possibility for
solubilization of the ore minerals and gangue. Although
the total dissolved-solid loading may not be extremely high,
the dissolved heavy-metal concentration may be relatively
high as a result of the highly mineralized ore being pro-
cessed. These heavy metals, the suspended solids, and process
reagents present are the principal waste constituents of a
mill waste stream. In' addition, depending on the process
conditions, the waste stream may also have a high or low pH.
The pH is of concern, not only because of its potential
toxiclty, but also because of its effect on the solubility
of the waste constituents.
Quantities of Wastes
The few mercury operations still active in late 1974 were,
for the most part, obtaining their ore from open-pit mines.
In the past, however, more than 2/3 of the domestic production
was from ore mined from underground mines. No discharge
exists from the open-pit mines visited or contacted during
this study. Also, no specific information concerning discharges
from underground mercury mines was available during the period
of this study. However, it is expected that, where discharges
occur from these underground mines, the particular metals
present and the extent of their dissolution depend on the
particular geology and mineralogy of the ore body and on
the oxidation potential and pH prevailing within the mine.
V-144
DRAFT
-------
DRAFT
Silica and carbonate minerals are the common introduced
gangue minerals in mercury deposits, but pyrite and marcasite
may be abundant in deposits formed in iron-bearing rocks.
Stibnite is rare but is more common than orpiment. Other
metals, such as gold, silver, or base metals, are generally
present in only trace amounts.
Process Description - Mercury Mining
Mercury ore is mined by both surface and underground methods.
Prior to 1972, underground mining accounted for about 60
percent of the ore and 70 percent of the mercury production
in the U.S. Currently, with market prices of mercury falling,
only a couple of the lower-cost open-pit operations remain
active.
The mode of occurrence of the mercury deposit determines
the method of mining; yet, with either type, the small
irregular deposits preclude the large-scale operations
characteristic of U.S. mining.
Process Description - Mercury Milling
Processes for the milling of mercury which require water and
result in the waste loading of this water are:
(1) Gravity methods of separation
(2) Flotation
One mercury operation (mill 9201) visited employs gravity
separation methods of beneficiation; the volume of the waste
stream emanating from this mill is approximately 1,679 cubic
meters (440,000 gallons) per day. In addition, another new
plant (mill 9202) due to begin production during early 1975
was contacted. This mill will use a flotation process and
expects to discharge 5.5 cubic meters (1,430 gallons) of
water per minute. These waste streams function to carry large
quantities of solids out of the mill. While the coarser
material is easily settled out, the very fine particles of
ground ore (slimes) are normally suspended to some extent in
V-145
DRAFT
-------
DRAFT
the waste water and often present removal problems. The
quantity of suspended solids present in a particular waste
stream is a function of the ore type and mill process, as
these factors determine how finely ground the ore will be.
In addition to suspended solids, solubilized and dispersed
colloidal or adsorbed heavy metals may be present in the
waste stream. Metals most likely to be present at relatively
high levels are mercury; antimony; and, possibly, arsenic,
zinc, cadmium, and nickel. The levels at which these metals
are present depend on the extent to which they occur in the
particular ore body. Calcium, sodium, potassium, and mag-
nesium normally are found at concentrations of 10 to 200
parts per million.
In the past, little beneficiation of mercury ores was accom-
plished and typically was limited to crushing and/or grinding.
In a few cases, gravity methods were used to concentrate the
ore. These practices require no process reagents. However,
the operation (mill 9202) due to open during 1975 will use
a flotation process, which will require the use of flotation
reagents. These reagents add to the waste loading of the
mill effluent as they are consumed in the process. The
reagents which are expected to be used at this mill are
listed in Table V-56.
Mill 9201 currently beneficiates mercury ore by gravity
methods. The ore is first crushed, washed, and screened
to provide a feed suitable for gravity separation. The ore
is concentrated by tabling, which essentially Involves washing
the crushed ore slurry across a vibrating table which has
ridges and furrows formed in parallel on its surface. As
the ore slurry is washed across this surface, the heavy ore
minerals collect in the furrows, while the fines are carried
across the ridges and discarded. The vibrating action causes
the heavy minerals to travel along the furrows to the end
of the table, where they are collected.
Sometime during the spring or early summer of 1975, mill
9202 is to begin operation for the concentration of mercury
sulfide ore by a froth flotation process.
Waste characteristics of mill effluents of the operation
visited and of a pilot-plant operation using the flotation
process are presented In Table V-57.
V-146
DRAFT
-------
DRAFT
TABLE V-56. EXPECTED REAGENT USE AT MERCURY-ORE FLOTATION
MILL 9202
REAGENT
Dowfroth 250 (Polypropylene glycol methyl ethers)
Z-11 (Sodium isopropyl xanthate)
Lime (Calcium oxide)
Sodium silicate
PURPOSE
Frother
Collector
Depressing
Agent
Depressing
Agent
CONSUMPTION
kg/metric ton
ore milled
0.15
0.13
0.05
0.10
Ib/short ton
ore milled
0.30
0.25
0.10
020
V-147
DRAFT
-------
DRAFT
TABLE V-57. WASTE CHARACTERISTICS AND RAW WASTE LOADINGS
AT MILLS 9201 AND 9202
MILL
9201
9202
IPUot Operation!
MILL
9201
9202
IPUoiOpornion)
MILL
9201
9202
(Pilot OporMionl
MINE
9201
9202
(Pilot Operation)
jnHi
6.6
-
Hg
WASTE LOAD
CONCEN.
r"AJIOM In kg/1000 mnric ton
Img/ZI (rb/IOOOehortMral
of conamreie produad
-
00072 11
122)
ki kg/1000 rn.tr* Mm
lfe/1 000 draft Mm)
of oramUled
-
0084
10.188)
Sb
CONCEN-
TRATION
<06
0.03
WASTE LOAD
In kg/1000 metric Mm
db/IOOOriurtMra)
of ooncMifraw produced
< e.900
K13JOO)
BO
1100)
In kg/1000 metric 1om
(ft/1000 ehontoml
oforenHIMd
<006
K010I
04
(08)
Te
CONCEN
TRATION
<0.08
WASTE LOAD
in kg/1000 metric torn
llb/1 000 riiort torn)
of eonantrete produced
< 1.100
(C2.200)
-
m kg/1000 metric torn
IB/I 000 dran torn)
of ore milled
<0008
K001W
:
Zn
CONCEN-
TRATION
014
WASTE LOAD
ui kg/1000 metric Mm
Ub/IOMdioriMinl
ol eonamrele produad
1.930
13.8601
-
m kg/1000 metric ton
iB/IODOfhorttam)
at ore milled
0014
(00281
-
Fi
CONCEN-
TRATION
<06
OM
WASTE LOAD
m kg/1000 metric Mm
lB/1000ihorttom)
of oonoMitratB produced
103300)
80
1160)
HI kg/1000 metric torn
(Ib/IOOOctontom)
olonmUled
CONCEN-
TRATION
InVtl
002
0.3S
WASTE LOAD
hi kg/1000 metric Mm
ID/1000 dan lam)
of oonamrete produced
270
(540)
600
11.200)
in kg/ 1000 metric torn
(lb/1000 dwn torn)
of ore millid
0002
10004)
5
(10)
Mn
CONCEN-
TRATION
(mg/t)
500
0.06
WASTE LOAD
m kg/1000 metric Mm
(fc/1 000 dwn lorn)
of conantrpto produad
688.000
11.376,0001
79
lisa)
in kg/1000 metric torn
lib/1000 short tontl
ol ore mill*)
5
110)
065
(1301
SULFIOE
CONCEN
TRATION
Img/t)
,06
~
WASTE LOAD
» kg/1000 metric torn
IB/1000 diort lorn)
of concerrlrele prodiiad
< 6000
KI3400)
-
m kg/1000 metric torn
lib/1000 dwn tontl
ol ore rrultod
<005
K010I
-
V-148
DRAFT
-------
DRAFT
Uranium, Radium, and Vanadium Ores
Water use; flow; and the sources, nature, and quantity of
wastes during the processes of uranium, radium, and vanadium
ore mining and beneficiation are described in this section.
For vanadium-ore mining and beneficiation, only those opera-
tions beneficiating ores containing source material (i.e. ,
uranium and thorium) at concentrations exceeding 0.05 percent
as defined by the Nuclear Regulatory Commission, and which
are subject to NRG licensing, are considered here.
Water Use. Uranium ores often are found In arid climates,
and water is conserved as an expensive asset in refining or
milling uranium, vanadium, and radium ores. Some mines yield
an adequate water supply for the associated mill, and a water-
use pattern as shown in part (a) of Figure V-31 can be employed.
Here, all or part of the mine water is used in the mill and
then rejected to an impoundment, from which it is removed
by evaporation and, possibly, seepage. Mine water—or at
least, that portion not needed in the mill—is treated to
remove values and/or pollutants. Sometimes the treated water
is reintroduced to the mine for in-sltu leaching of values.
Waste water from the impoundment is recycled to the mill when
conditions warrant, and additional recycle loops (not shown
in the figure) may be attached to the mill itself.
When mines are dry or too far from the mill to permit economi-
cal utilization of their effluents, the mill derives water
from wells or, rarely, from a stream (part (b) of Figure
V-31). In these instances, any mine water discharge may
be treated to remove uranium values and/or pollutants, and
these are then shipped to the mill (part (c) of Figure V-31).
There are completely dry underground mines and open-pit mines
that lose more water by evaporation than they gain by seepage
from aquifers. All known mills in this industry segment
use a hydrometallurgical process.
The quantity of water used in milling is approximately equal
in weight to that of the ore processed. The quantity of
makeup water is equal to or less than this in the presence
of recycling. From these considerations, it is apparent that
the entire uranium milling industry uses a fairly small quan-
tity of water, about 30,000 cubic meters (7.9 million gallons)
per day.
V-149
DRAFT
-------
DRAFT
Figure V-31. TYPICAL WATER-USE PATTERNS
F IN-SITU LEACH ~~\
I ~~
MINE
H
TREATMENT
MILL
i
»f IMPOUNDME
^^^^^ ^—
i
(a) WET MINE/MILL COMPLEX
TREATMENT
MILL
RECYCLE
(b) SEPARATED MILL
n IN-SITU LEACH
MINE
TREATMENT
DISCHARGE
(e) SEPARATED WET MINE
V-150
DRAFT
-------
DRAFT
Waste Constituents
Radioactive Waste Constituents. Radium is the most potent
cumulative poison known, and its dissemination has been care-
fully supervised since the 1920's. The chemistry of radium
is similar to that of calcium, barium, and strontium. The
amount of radium that may be dissolved in water which can
be used for human consumption has been set by the International
Commission on Radiological Protection (ICRP) at three picocuries
(or picograms) per liter—i.e., about six orders of magnitude
below other typical pollutant concentrations.
Radium, with a half-life of 1,620 years, is generated by the
radioactive decay of uranium, which has the very long half-
life of 4.51 billion years. In uranium ores that are in place
for billions of years, an equilibrium could be established
between the rate of decay of uranium into radium and the
rate of decay of radium into its daughters. Once this equili-
brium is established, the ratio of uranium to radium equals
the ratio of the half-lives—i.e., 2.7 million. An equili-
brated ore with a typical grade of 0.22 percent uranium would
contain 0.82 microgram of radium per kilogram. If the milling
process were to permit all of this radium to go into solution
in the water used to leach (or otherwise treat) the ore, and
if water were used in the typical ratio of one ton to each
ton of ore, then, without recycling, the concentration of
radium in tailings would also be 0.82 microgram per liter.
Although it is one of the least soluble substances, radium
sulfate is soluble to 20 micrograms per liter, and the above
concentration could be accommodated as a solute. Geological
redeposition reduces the amount of radium in the ore. Because
milling processes preferentially dissolve uranium and leave
radium in solid tailings, actual concentrations of radium in
tailing-pond solutions range from 0.000001 to 0.0021 microgram
per liter (1 picogram to 2,100 picograms per liter). These
concentrations are often quoted in curies (Ci)—i.e., 1 to
2,100 picocuries per liter (pCi/1)—since the radioactive
source strength of a quantity of radium in curies is essentially
equal to its content of radium by weight in grams. (Source
strength unit for radionuclides has been defined as that
quantity of radioactive material that decays at a rate of
37 billion (3.7 x 10 exp 10) disintegrations per second.)
V-151
DRAFT
-------
DRAFT
The radium content of tailing-pond solutions must, therefore,
be reduced by a factor of up to 600 to achieve ICRP drinking
water quality of 3 p'Cl/1.
Thorium. There are other radioactive species that result
from the decay of uranium. Only thorium 230, with a half-
life of 80,000 years, Is usually considered In addition to
radium 226. It Is observed In concentrations of 1 to 100,000
pCi/1 In tailing-pond solutions. A maximum concentration
for Th 230 of 2,000 pCl/1 Is recommended In an early radiological-
exposure standard (10CFR 20) that also sets a Ra 226 limit
of 30 pCi/1 (which is ten times the IRCP potable-water limit).
Chemical and Physical Waste Constituents. Chemical contami-
nants of milling wastewaters derive from compounds Introduced
in milling operations or are dissolved from ore in leaching.
The common physical pollutants—primarily, suspended solids—
figure prominently in discharges from wet mines, and In the
management of deep-well disposal and recycle systems.
Seventy percent or more of radium 226 usually ends up in
solid tailings, including the suspended solids of effluents.
Additional pollutants (particularly, metals) are expected
to appear in the waste streams of specific plants that might
be using unusual ores. Certain compounds, particularly organics,
are expected to undergo changes and are not identifiable
individually but would appear in waste-stream analysis under
class headings (e.g., as TOC, oils and greases, or surfactants).
In one specific example, It has been observed that oils and
greases that are known to enter alkaline leach processes
disappear and are replaced by approximately equivalent quanti-
ties of surfactants—presumably, by saponlfication (the process
involved in soap manufacture). Table V-58 shows waste
constituents expected from mills based upon the process-
chemical consumption and the ore mineralogies which are
commonly encountered. These substances are shown in three
groups: those expected from acid leach processes, those
expected from alkaline leach processes, and metals expected
to be leached from the ore during milling processes.
Table V-59 shows two groups of constituents (among the sets
of parameters which were analyzed both in background waters
and waste streams): (1) Constituents that were found to
exceed background by factors from three to ten; and (2)
Constituents that were found to exceed background by a factor
V-152
DRAFT
-------
DRAFT
TABLE V-58. WASTE CONSTITUENTS EXPECTED
ACID LEACH PROCESS
ACID-LEACH CIRCUIT:
Sulfuric acid
Sodium chlorate
LIQUID/SOLID-SEPARATION CIRCUIT:
Polyacrylamides
Guar gums
Animal glues
ION-EXCHANGE CIRCUIT:
Strong base anionic resins
Sodium chloride
Sulfuric acid
Sodium bicarbonate
Ammonium nitrate
SOLVENT-EXTRACTION CIRCUIT:
Tertiary amines
(usually, alamine-336)
Alky I phosphoric acid
(usually. EHPA)
Isodecanol
Tri butyl phosphate
Kerosene
Sodium carbonate
Ammonium sulfate
Sodium chloride
Ammonia gas
Hydrochloric acid
PRECIPITATION CIRCUIT:
Ammonia gas
Magnesium oxide
Hydrogen peroxide
ALKALINE LEACH PROCESS
ALKALINE-LEACH CIRCUIT:
Sodium carbonate
Sodium bicarbonate
ION-EXCHANGE CIRCUIT:
Strong base anionic resins
Sodium chloride
Sulfuric acid
Sodium bicarbonate
Ammonium nitrate
PRECIPITATION CIRCUIT:
Ammonia gas
Magnesium oxide
Hydrogen peroxide
METALS LEACHED FROM ORE
BY MILLING PROCESSES
Magnesium
Copper
Manganese
Barium
Chromium
Molybdenum
Selenium
Lead
Arsenic
Vanadium
Iron
Cobalt
Nickel
SOURCE: Reference 28
V-153
DRAFT
-------
DRAFT
TABLE V-59. CHEMICAL AND PHYSICAL WASTE CONSTITUENTS
OBSERVED IN REPRESENTATIVE OPERATIONS
MINE/
CATEGORY
9401 /
ALKALINE
9402/
ACID
9403/
ALKALINE
9404
ACID
9405/
ACID
9406/
MINE
CONSTITUENTS THAT EXCEED BACKGROUND*
BY FACTORS BETWEEN THREE AND TEN
Color. Cyanide. Nitrogen n Ammonia, Phosphate.
Total Solids. Sulfata, Surfactants
Pb
Acidity. COD. Color. Dissolved Solids. Phosphate.
Total Solids
Ag. B. Ba. Hg. Zn
Color. Dissolved Solids. Fluoride. Sulfate. Total
Solids. Turbidity
Chloride. Color. Dissolved Solids. Total Solids.
Turbidity
Ag. Hg. K. Mg. Na
Color. Conductivity. Fecal Coliforrn. Hardness.
Phosphate. Suspended Solids. Total Solids.
Turbidity
Al. As. B. Ba. Be. Ca. Cd. Cr. Cu. Fe. Hg. Mg, Mo.
Ni. Pb, Sb. Se. Zn
Ammonia. Chloride. Hardness, Nitrete. Nitrite. Oil
and Grease. Organic Nitrogen. Sulfate, Total Solids,
Turbidity
As. B, Be. Ca. Mg. Na
CONSTITUENTS THAT EXCEED BACKGROUND*
BY A FACTOR OF MORE THAN TEN
Alkalinity. COD. Fluoride, Nitrate
As, Mo. V
Ammonia. Chloride, Sulfate
Al. As. Be. Cr. Cu. K. Mg. Mn. Mo. Na. Ni. Pb, V
Chloride, COD, Nitrate, Surfactants, Suspended
Solids, TOC
As, Mo. Na, Ti, V
Acidity. Ammonia. Sulfata, Suspended Solids
Al. As. Cr. Fe, Mn. Ni, Pb. Ti, V. Zn
Chloride, COD. Dissolved Solids. Kieldahl Nitrogen,
Nitrate. Volatile Solids
Co, K. Mn. Na
(None among the analyzed items)
•"Background" is defined in text.
V-154
DRAFT
-------
DRAFT
of more than ten. Comparison of Tables V-58 and V-59
illustrates that more, rather than fewer, pollutants are
observed to be "added" by the operation than are predicted
from process chemistry and ore characteristics. Observed
pollutant increases in conjunction with toxicant lists
were, therefore, used to select the parameters on which field
sampling programs were to concentrate. (See also Section VI.)
Table V-59 also illustrates some specific differences among
the subcategories of SIC 1094 that are further explored in
the following discussion.
Constituents Introduced in Acid Leaching. Acid leaching
(discussed in Section III) dissolves numerous ore constituents,
between one and three percent of the ore, that appear in
the process stream; upon successful extraction of uranium
and vanadium values, these ore constituents are rejected to
tailing solutions. In plants using a sulfuric-acid
leach, calcium, magnesium, and iron form sulfates directly.
Phosphates, molybdates, vanadates, sulfides, various oxides,
and fluorides are converted to sulfates with the liberation
of phosphoric acid, molybdlc acid, hydrogen sulfide, and
other products. The presence of a given reaction product
depends on the type of ore that is being used; since this is
variable, pollutant parameters must be selected from an
inclusive list. The major pollutant in an acid leach opera-
tion is likely to be the sulfuric acid Itself, since a free
acid concentration of one to one hundred grams of acid per
liter is maintained in the leach.
Excess free acid remaining in the leach liquors and in solvent-
extraction raffinates (nonsoluble portions) can be recycled
to advantage. In some operations, this acid is used to condi-
tion incoming ores by reaction with acid-consuming gangue.
Although this step aids in controlling pH of raw wastes, it
does not reduce the amount of sulfates therein.
Oxidants are added to the acid leach liquor following Initial
contact with ore and after reducing gases, such as hydrogen
and H2S, have been driven from the slurry. They act in
conjunction with an iron content of about 0.5 g/1 to assure
that uranium is in the U(VI) valence state. Sodium chlorate
(NaC103) and manganese dioxide (Mn02) serve this purpose in
quantities of 1 to 4 g/1. The species of a pollutant in the
effluent will normally be one of the more oxidized forms—
e.g., ferric rather than ferrous iron.
V-155
DRAFT
-------
DRAFT
Constituents Introduced in_ Alkaline Leaching. Alkaline
leaching is less likely to solubilize compounds of iron
and the light metals and has no effect on the common carbonates
of the gangue. Sulfates and sulfldes, in the oxidizing
conditions required for conversion of U(IV) to U(VI), consume
sodium carbonate and, together with the sulfate ion generated
In the common method of sodium removal, pollute waste waters.
The wastewater of an alkaline leach mill is largely derived
from two secondary processes (Figure V-32): tailing repulping,
and purification (or sodium removal). The leach itself is
recycled via the recarbonatlon loop. The wastes discarded
to tailings often contain organic compounds derived from
the ores. Oxidizing agents are used in leaching, but air
and oxygen gas under pressure have been found to serve as
well as more expensive oxldants and to reduce pollutant
problems. The concentrations used in alkaline leach are only
of academic interest because of recycling. Sodium carbonate
concentration varies from 40 to 50 g/1; sodium bicarbonate
concentration, from 10 to 20 g/1.
An ammonium carbonate process that leads directly to a sodium-
free uranium trioxide product has been investigated. It is
more selective for uranium than the sodium carbonate process,
but vanadium, while not being recovered, interferes with
uranium recovery. The process does not require bicarbonate
and could produce ammonium sulfate, a needed fertilizer, as
a byproduct (Section III). A flow chart of an ammonium car-
bonate process is shown in Figure V-33.
Constituents Introduced in Concentration Processes. Ion-
exchange (IX) resins are ground into small particles that
appear among suspended solids in raw waste streams. Solvents
are not completely recovered in the phase-separation step of
solvent-exchange (SX) concentration. The extent of the contri-
butions of each of these pollutants is difficult to judge by
observation of the waste stream, since there are no specific
analysis procedures for these contaminants. Some prediction
of the concentration is possible from the observable loss of
(IX) resin and SX solvents. Only a small fraction of IX
V-156
DRAFT
-------
DRAFT
Figure V-32. ALKALINE-LEACH WATER FLOW
GROUND ORE
TO ATMOSPHERE
FRESH WATER
OR
TAILING-SOLUTION
RECYCLE
ALKALINE
LEACH
t
FILTERING
i
uatu
STACK
, OMO
LEACH ,
RECYCLE 1
1 1
RECARBONATION
PREGNANT
LEACH
PRECIPITATION
CRUDE
PRODUCT
STACK
'GAS
t
BARREN
LEACH
FRESH
WATER
PURIFICATION
(SODIUM REMOVAL)
REPULPED
'TAILINGS
WASTE
WATER'
. TO TAILING
POND
END
PRODUCT
I
TO
STOCKPILE
V-157
DRAFT
-------
DRAFT
Figure V-33. AMMONIUM CARBONATE LEACHING PROCESS
MINING
TO ATMOSPHERE
ORE
*
I
WASTE GAS
AIR
WATER
—&•
—*•
GRINDING
1
PRESSURE LEACHING
1
COUNTERCURRENT
DECANTATION
^ LEACH
^SOLUTION"
PREGNANT
SOLUTION
AMMONIA AND
CARBON DIOXIDE
ABSORPTION TOWERS
t
' t
1
I
STEAM
1
1
1
GAS
1
1
STEAM S
PRECIP
i
1
TRIPPING
RANIUM
TATION
SLURRY
•FILTRATE-
TO TAILING
POND
SLURRY
FILTRATION
PRODUCT
TO
STOCKPILE
V-158
DRAFT
-------
DRAFT
resin is actually lost by the time it is replaced because
of breakage; in one typical operation, the loss amounts
to about 100 kg (220 Ib) per day at a plant that has
an inventory of about 500 metric tons (551 short tons)
of resin and handles 3,000 metric tons (3,307 short tons)
of resin per day of ore and about as much water. The
raw waste concentration of IX resin can thus be estimated
as about 30 ppm. Standard tests for water quality would
measure this contribution as total organic carbon (TOC)
due to other sources (for example, organic ore constituents).
Most of this contribution is in suspended solids; this
is illustrated by the fact that TOC is only about 6 mg/1
in the supernatant of the raw waste stream discussed
above.
Solvents are lost at a rate of up to 1/2000 of the water
usage in the SX circuit. This ratio is set by the solubil-
ities of utilized solvents, which range from 5 to 25 mg/1,
and by the fact that inadequate slime separation can
lead to additional loss to tailing solids. TOC of the
raw waste supernatant at mills using SX was found to
be 20 to 24 mg/1. It is, again, impossible to determine
what part of this measurement should be ascribed to SX
solvents—particularly, in view of highly carbonaceous
ores.
The most objectionable constituents present in mill effluents
may be the very small amounts (usually, less than 6 ppm)
of the tertiary amines or alkyl phosphates employed in
solvent extraction. In some cases, these compounds have
been found to be toxic to fish. Dilutions of up to 1,500:
1 have been used (Reference 28) to lower the organic
concentrations of effluents of this type to acceptable values.
An analytic procedure for the entire class of these mate-
rials and their decay products is not available, and they
must be identified in specific instances.
Difficulties in distinguishing among solvents, ion-exchange
resins, carbonaceous ore constituents, and their degradation
products made it impossible to discriminate between the
wastes of mills using SX or IX processes. Since some
of the solvents have structures with potential for toxic
effects in their degradation products, it would be desirable
to trace their fates as well as those of ion-exchange
resins. Future research in this field could lead to
better characterization and improved treatment of waste-
water.
V-159
DRAFT
-------
DRAFT
Process Descriptions. Water Use, and Waste Characteristics for
Uranium. Radium, and Vanadium Ore Mining and Milling
Four mine/mill complexes in the licensed segment of the SIC
1094 category were visited to collect data on the utilization
of water and the characteristics of raw and treated wastes.
Water use in the mines and mills is listed in Table V-60, and
treatment systems employed are listed in Table V-61.
The consumption of water is seen to vary from 0.75 to 4.3
cubic meters per metric ton (180 to 1,030 gal per short ton)
of ore capacity, with an average of 1.35 cubic meters per metric
ton (323 gal per short ton) . Two of the operations (9401
and 9404) derive their water supply from wells, and one (9403)
obtains its water from a stream, in the manner shown in Figure
V-34c. The fourth operation (9402) utilizes mine water.
Where mine water is available, at least some of it is treated
by ion exchange to recover uranium values. Water use in
representative operations is illustrated in Figure V-34, and
the water-flow configurations of these operations are illus-
trated in Figures V-35, V-36, V-37, and V-38. While an attempt
was made to obtain a water balance in each case, there are some
uncertainties. In Figure V-35, for example, the loss from
tailings by seepage and evaporation is probably not quite
equal to the raw waste input from the plant, and expansion of
the tailing-pond area may be necessary. Similarly, it proved
difficult to account for the rain water entering the open-
pit mine of the operation in Figure V-38. If and when it rains
into this mine, some water evaporates immediately from the
surface, while the rest runs into a central depression or
seeps into underground aquifiers. The first and last effects
in combination are clearly dominant; less than ten percent of
the calculable water input is seen to evaporate from the
central depression (Figure V-38).
V-160
DRAFT
-------
DRAFT
Figure V-34. WATER FLOW IN MILLS 9401, 9402, 9403, AND 9404
•t*<** «• n*ir
MINE I . »| IONEXCHANOE I I*ni100yfdl „ I DISCHARGE I
, 1 8J3»m3Miy I , 1 H I
lao;ioo-di \^7^
(1.320.000 gpd)
1.636 oi3Miy
(431^00 gpd)
(a) MILL 9401
. 4.3J8 m3M«
' (1.140.000 gpd)
3.800 m:._.
(061.000 gpd)
,3««1
'-'I I
« 6.307 m3Miy
I < (1.400.000 gpd)
-»4 LOSS
(b) MILL 9402
TO .
ATMOSPHERE A
EVAPORATION
r— ""^ M80
(i.80o.t
B.400 m3Miy
11.430.000 gpd)
.3«W
DO gpd)
DISCHARGE
MILL
f 620 m3Miy
1 (137.400 gpdl
I— — I RECARBONATION
I_
1
P
_ s^^.
(174.400 gpdl
1
3POND%-
ITATION
(c) MILL 9403
gpd)
OPEMOT
M1Mi ' 1J81mW
TO ATMOSPHERE
TAILING
POMD -X CAPACITY OF
1.636 m3Mn
1432/WOgpdl
(d) MILL 9404
V-161
DRAFT
-------
DRAFT
TABLE V-60. WATER USE AND FLOWS AT MINE/MILLS 9401, 9402, 9403,
AND 9404
WATER CATEGORY
WATER USED
MINE/MILL 9401
m3/day
gpd
MINE/MILL 9402
m3/day
gpd
MINE/MILL 9403
m3/day
gpd
MINE/MILL 9404
m3/day
gpd
MINE PORTION
Water Supply
Discharge
Supplied to Mill
Recycled to Mill
Loss (Evaporation, etc.)
8,339
3,339
0
6.000
estO
2,203,000
882.100
0
1.321.000
estO
11.552
4,325
S.307
0
1,920
3,052.000
1.143.000
1.402.000
0
507,200
N/A
N/A
N/A
N/A
N/A
N/A
N/A
N/A
N/A
N/A
est 1.530
0
0
0
est 1,630
est404.200
0
0
0
est 404,200
MILL PORTION
Makeup Water
Water in Circuit
Discharge
Evaporation and
Seepage
2.700
3.200
0
2.700
713.300
845.400
0
713,300
6,307
8,900
0
6.307
1,402,000
2.351.000
0
1,402.000
6.060
6^80
5,400
660
1.601,000
1.738.000
1,427.000
174,400
5300
5,300
0
6,300
1.400.000
1.400,000
0
1,400.000
N/A - Not available
TABLE V-61. WATER TREATMENT INVOLVED IN U/Ra/V OPERATIONS
FEATURE
Smling Baun
Evaporating Pond
Ion-Exchange Plant
Tailing Pondlil
Ion-Exchange Plant
Recarbonitar
OacpWall
Utitiution of
Or« Handling
PARAMETER
S MINE/MILL
9401
9402
MINE PORTION
Area in hecterai lecrei )
Retention Time in houn
Area in hectares lacret)
U3 Og Concentration in mg/l
U3 OB Removal in *
OJ (0 74)
eft 20
N/A
25
96
O.J 11.7)
en 80
N/A
2 to 12
98
9403
N/A
N/A
N/A
N/A
N/A
MILL PORTION
Area in hectare* (acres)
Number term-connected
Daily Water UM in metric ton (inort tons)
Daily Water UM in metric torn (ihori lont)
Capacity in metric com Miort torn) water per day
Send/Slime Separator*
Decant Facilities
Filter.
Coprecipitation
21 (SI 8)
1
490 (540)
1.635(1 .8021
0
YM
V«
100(247)
6
N/A
N/A
0
Vei
TOTAL OPERATION
Capacity in metric ton hhon torn) per day || 3.200(3.527)
6.40017.055)
24 (59.3)
3
N/A
520(573)
0
Yea
1.400(1343)
9404
N/A
N/A
2 14.91
N/A
N/A
107 (264)
1
N/A
N/A
1.635(1.602)
Ves
Ves
Yes
2.70012.976)
N/A - Not available
V-162
DRAFT
-------
DRAFT
Figure V-35. FLOWCHART OF MILL 9401
TO ATMOSPHERE
L
TO ATMOSPHERE
TO STOCKPILE
V-163
DRAFT
-------
DRAFT
Figure V-36. FLOW CHART FOR MILL 9402
MINING
5.300 m3/day
(1,400,000 gpd)
ORE
CRUSHING
AND
GRINDING
2S°4
Nad0
3 \.
NaCI
NH,
LEACHING
THICKENERS
SOLVENT
EXTRACTION
PRECIPITATION
YELLOW CAKE
TO STOCKPILE
EVAPORATION
AND SEEPAGE
5,300 m3/day
(1.400.000 gpd)
V-164
DRAFT
-------
DRAFT
Figure V-37. FLOW CHART OF MILL 9403
0 8 m'Atey
Ulgpml
FLOAT TAILS
ALTERNATIVE
WASH EQUIVALENT TO MOISTURE
RETAINED IN CAKE (ELIMINATED
IN DRVER)
FROM ALKALINE
TAILING SUMP
6B4m3Miv
(120 opm)
FROM ACID
TAILING SUMP
2jM3iMnc Ion/Ay
(3.134 ikon um/diy)
LIQUID
44-HECTARE 1108-ACRE)
TAILING POND WITH 23-HECTARE
(SB-ACRE) LIQUID POOL
TO STOCKPILE
V-165
DRAFT
-------
DRAFT
Figure V-38. FLOW CHART OF MILL 9404
MILL WATER SUPPLY = 5,300
RAIN
1,530 m3/day
< i (404,200 gpd)
m3 (1.400.000 gal) per day
TO ATMOSPHERE
EVAPORATION
.137 m3/day
(36.200 gpd)
C320-hectare (790-acre) /2-hectare (4.9-acreKA
OPEN-PIT MINE I EVAPORATION ) )
Vy^POND^X^/
TOTAL LOSS OF
5,300 m3/day
(1,400,000 gpd) F
ii L
( TAILING ^
V POND A
\ FILTER ]
CAPACITY OF
1,635 m3/day
(431,900 gpd)
i i
( DEEP WELL
( WELL 2 ^
REPULP ] « SANDS
i i
)
BARREN
SOLUTION
V-166
DRAFT
MINING |
ORE
GRINDING •*— |
1
ACID LEACH
BIN
"1 ••
HYDROCYCLONES j
SLIMES
RESIN-IN-PULP
ION EXCHANGE ^_,
1
PREGNANT
SOLUTION
1ELLUE
SE
NT
(HIGH CD
PRECIPITATION
AND
FILTERING
~I}_
— WASHING
-LOW c rJ
-------
DRAFT
Waste Characteristics Resulting From Mining and Milling
Operations. Two of the operations visited use alkaline
leaching, and two use acid leaching, for extraction of ura-
nium values. Only one operation discharges from the mill,
while two others discharge from mines. Among the five NRC-
licensed subcategories listed in Section IV, only mills
employing a combination process of acid-and-alkaline leach-
ing are not represented by the plants visited. An operation
representing this subcategory was not visited because its
processes were changed recently. During the visits to these
mills, industry plans that change water use by factors of up
to ten, and which will take place within a year, were pre-
sented. The data on raw wastes presented in the following
discussion are based mostly on analyses of samples obtained
during site visits.
The data obtained are organized into several broad waste
categories:
1. Radioactive nuclides.
2- Organics, including TOC, oil and grease, surfac-
tants, and phenol.
3. Inorganic anions, including sulfide, cyanide,
fluoride, chloride, sulfate, nitrate, and
phosphate.
4. Light metals, relatively nontoxic, including
sodium, potassium, calcium, magnesium, aluminum,
titanium, beryllum, and the ammonium cation
(NH4+).
5. Heavy metals. some of which are toxic, including
silver, aluminum, arsenic, barium, boron, cadmium,
chromium, copper, iron, mercury, manganese, moly-
bdenum, nickel, lead, selenium, strontium, tellur-
ium, titanium, thallium, uranium, vanadium, and
zinc.
This class is further subdivided into the metals forming
primarily cationic species and those forming anionic species
in the conditions characteristic of raw SIC 1094 wastes (in
particular, chromium, molybdenum, uranium, and vanadium).
6. Other pollutants (general characteristics), includ-
ding acidity, alkalinity, COD, solids, color, odor,
turbidity and hardness.
V-167
DRAFT
-------
DRAFT
Radioactive Nuclides. Decay products of uranium include
Isotopes of uranium, thorium, proactinium, radium, radon,
actinium, polonium, bismuth, and lead. (Fission products,
generated by bombardment with cosmic or manmade neutrons,
are undoubtedly also present but are too rare to deserve
further consideration.) These decay products respond to
mining and milling processes in accordance with the chemistries
of the various elements and, with the exception of the bulk
of uranium isotopes, appear in the wastes. A fair fraction
of the most-toxic isotope, radium 226, remains with solid
tailings and sediment in mine-water settling basins. Raw
waste, including these solids, shows concentrations of
radium and uranium that should not be released to the environ-
ment. The amounts that have been observed under this program
are shown in Table V-62, where it is seen that alkaline
mills are highest, mines are second highest, and acid mills
are lowest in the radium content of wastes. The high levels
encountered at mines are alarming, although partially
explainable by buildup in the recycle accompanying ion-exchange
recovery of uranium. Recycle also explains the high radium
loads found at alkaline mills. The low concentrations
observed at acid mills are partially due to the low solubility
of radium sulfate (formed by reaction with sulfuric acid
leach) and to the lack of recycle, but concentrations—
shown in parentheses—for an evaporation and seepage pond
illustrate that such Impoundments may become a pollution
hazard to ground-water supplies.
Organics. Organics derived from carbonaceous ores and from
chemicals added in processing are measured as TOC and, occasion-
ally, are distinguishable as oils or greases, surfactants, or
phenol. The small amounts of organlcs that are observed are
reviewed in Table V-63.
Inorganic Anions. These may be distinguished into two classes:
(1) Sulfldes, cyanides, and fluorides, for which technically
and economically feasible treatments (e.g., oxidation and lime
precipitation) are readily available; and (2) Chlorides, sul-
fates, nitrates, and phosphates, which are present in fairly
large concentrations in mill wastes and cannot be removed
economically. Distillation and reverse osmosis, while tech-
nically feasible, raise the cost of recovered water to the
level of $5 per cubic meter ($20/1000 gal) and require a
large energy expenditure. Impoundment, in effect, results
in distillation in regions like the southwestern states.
V-168
DRAFT
-------
DRAFT
TABLE V-62. RADIONUCLIDES IN RAW WASTEWATERS FROM
URANIUM/RADIUM/VANADIUM MINES AND MILLS
RADIONUCLIDE
and units of measurement
RADIUM 226
in picocuries/£
THORIUM
in mg/£
URANIUM
in mg/£
CONCENTRATION
MINES
200 to 3,200
<0.1
4 to 25
ACID MILLS
200 to 700
(4.100)*
(1.1)"
30 to 40
ALKALINE MILLS
100 to 19,000
N/A
4 to 45
•Parentheses denote values measured in wastewater concentrated by evaporation
N/A = Not available
TABLE V-63. ORGANIC CONSTITUENTS IN U/Ra/V RAW WASTE WATER
PARAMETER
Total Organic Carbon (TOG)
Oil and Grease
MBAS Surfactants
Phenol
CONCENTRATION (mg/'l
MINES
16 to 45
3 to 4
0.001 to 7
<0.2
ACID MILLS
6 to 24
1
0.5
<0.2
ALKALINE MILLS
1 to 450
3
0.02
<0.2
V-169
DRAFT
-------
DRAFT
Other anlons are grouped together in conjunction with the
light-metal cations as total dissolved solids and are found
in the levels shown in Table V-64.
Light Metals. The ions of sodium, potassium, and ammonium
found in wastewaters are subject to inclusion in the cate-
gory of total dissolved solids. Calcium, titanium, magnesium,
and aluminum respond to some treatments (e.g., lime neutrali-
zation) and are shown separately. Table V-65 shows concen-
trations of aluminum, beryllium, calcium, magnesium, and
titanium found in wastewater effluents of mines and mills
covered in this ore category.
Heavy Metals. The leach processes in the uranium/vanadium
industry involve highly oxidizing conditions that leave a
number of ore metals—specifically, arsenic, chromium, moly-
bdenum, uranium, and vanadium—in their most oxidized states,
often as arsenates, chromates, molybdates, uranates and vana-
dates. These anionic species are, typically, much more soluble
than cations of these metals that precipitate as hydroxides
or sulfldes in response to lime and sulfide precipitation
treatments. Most of these anions can be reduced to lower
valences by excess sulfide and will then precipitate (actually,
copreclpitate with each other) and stay in solid form if
buried by sediment. The observed range of concentrations
for the anionic heavy metals for mines and mills visited is
shown in Table V-66. One or more of the heavy metals is
observed in high concentrations in each type of operation.
The catlonic heavy metals that had been expected to occur
from data on ores and processes include lead, manganese,
iron, and copper. Field sampling results added nickel, sil-
ver, strontium, and zinc to this list. The observed concen-
trations of these metals are shown in Table V-67. Cadmium
was found in a concentration above the lower detection limit
(20 micrograms per liter) at one alkaline mill discharge.
Other Pollutants. Acid leach mills discharge a portion of
the acid leach; alkaline leach mills discharge sodium car-
bonate; and mine water is found to be well buffered with
measurable acidity and alkalinity. Chemical oxygen demand
is occasionally high, and raw wastes, reslurried only to
the extent needed for transport to tailings, carry a high
load of total solids. These factors are reflected in the
data shown in Table V-68. These measures indicate the need
V-170
DRAFT
-------
DRAFT
TABLE V-64. INORGANIC ANIONS IN U/Ra/V RAW WASTEWATER
PARAMETER
Sulfide
Cyanide
Fluoride
Total Dissolved Solids (TDS)
CONCENTRATION (mg/l)
MINES
<0.5
<0.01
0.45
1.400 to 2,000
ACID MILLS
<0.5
<0.01
< 0.01
15.000 to 36.000
ALKALINE MILLS
< 0.5
< 0.01 to .04
1.4 to 2.1
5,000 to 13.000
TABLE V-65. LIGHT-METAL CONCENTRATIONS OBSERVED IN U/Ra/V
RAW WASTEWATER
PARAMETER
Aluminum
Beryllium
Calcium
Magnesium
Titanium
CONCENTRATION (mg/l]
MINES
0.4 to 0.5
0.01
90 to 120
35 to 45
0.8 to 1.1
ACID MILLS
700 to 1.600
0.08
220
550
7
ALKALINE MILLS
0.2 to 20
0.006 to 0.3
5 to 3,200
10 to 200
2 to 15
TABLE V-66. CONCENTRATIONS OF HEAVY METALS FORMING
ANIONIC SPECIES IN U/Ra/V RAW WASTEWATER
PARAMETER
Arsenic
Chromium
Molybdenum
Uranium
Vanadium
CONCENTRATION (rnqfZ }
MINES
0.01 to 0.03
<0.02
0.5 to 1.2
2 to 25
as to 2.1
ACID MILLS
0.1 to 2.5
2 to 9
0.3 to 16
30 to 180
120
ALKALINE MILLS
0.3 to 1.5
<0.02
<0.3
4 to 50
0.5 to 17
V-171
D«AFT
-------
DRAFT
TABLE V-67. CONCENTRATIONS OF HEAVY METALS FORMING
CATION 1C SPECIES IN U/Ra/V RAW WASTEWATER
PARAMETER
Silver
Copper
Iron
Manganese
Nickel
Lead
Zinc
CONCENTRATION (mg/£ 1
MINES
<0.01
<0.5
0.2 to 15
< 0.2 to 0.3
< 0.01
0.07 to 0.2
0.02 to 0.03
ACID MILLS
<0.01
0.7 to 3
300
100 to 210
1.4
0.8 to 2
3
ALKALINE MILLS
0.1
<0£to1
0.9 to 1.6
< 0.2 to 40
05
<0.5 to 0.7
OA
TABLE V-68. OTHER CONSTITUENTS PRESENT IN RAW WASTEWATER
IN U/Ra/V MINES AND MILLS
PARAMETER
Acidity
Alkalinity
Chemical Oxygen
Demand (COD)
Total Solids
CONCENTRATION (mg/£)
MINES
2
200 to 230
<10to750
200 to 10.000
ACID MILLS
4.000
0
30
300.000 to 500.000
ALKALINE MILLS
0
1,000 to 5.000
10
100.000 to 300,000
V-172
DRAFT
-------
DRAFT
for settling, neutralization, and aeration of the wastes
before discharge. Those treatments also effect significant
reductions in other pollutants; for example, neutralization
depresses heavy metals, and aeration reduces organics.
Waste Loads in Terms of_ Production. The loads of those
pollutants that indicated conditions warranting treatment
at the exemplary plants were related to ore production to
yield relative waste loads. The data for three subcategories
of the SIC 1094 segment are presented in Table V-69 (mines)
and Tables V-70, V-71, V-72, and V-73 (mills).
Occasional large ratios between the parameters observed at
differing operations are believed to be due to ore quality.
The point is illustrated by TOC at mills 9401 and 9403:
The operators of mill 9401 had contracted to run an ore
belonging to mine 9404 on a toll basis. The ore carried a
high carbonaceous material content that caused water at
the 9401 mill to turn brown and may have adversely affected
the concentration process at mill 9404. Mill 9403, in con-
trast, was concentrating its own, much cleaner, ore. The
ratio of 200:1 in TOC is, therefore, expected.
Radium, uranium, and thorium will be control parameters in
this segment on the basis of the code of Federal Regulation
10, part 20 (10CFR20) and 10CFR25, regardless of their con-
centrations in effluents. This decision is dictated by the
subcategorization and the assumption that concentrations
of source material (uranium and thorium) in the ore exceed
0.05 percent.
Metal Ores - Not Elsewhere Classified (SIC 1099)
This section discusses the water uses, sources of wastes,
and waste loading characteristics of operations engaging
in the mining and milling of ores of antimony, beryllium,
platinum-group metals, rare earth-metals, tin, titanium, and
zirconium. The approach used in discussion of waste character-
istics of these (SIC 1099) metal processes includes a
general discussion of water uses and sources of wastes in
the entire group, followed by a description of the character
and quantity of wastes generated for each individual metal
listed above.
V-173
DRAFT
-------
DRAFT
TABLE V-69. CHEMICAL COMPOSITION OF WASTE WATER AND RAW WASTE LOAD
FOR URANIUM MINES 9401 AND 9402
PARAMETER
TSS
COD
TOC
Alkalinity
Ca
MB
Fe
Mo
V
Ra
Th
U
MINE 9401
CONCENTRATION
(mO/ 8.)
IN WASTEWATER
_
242
16.8
224.4
93
46
0.47
0.6
1.0
3,190*
-
12.1
RAW WASTE LOAD
kg/day
—
2.300
150
2.100
860
420
4
6
9
29,700t
-
113
to/day
—
5.200
320
4.600
1.900
920
10
11
20
-
-
248
MINE 9402
CONCENTRATION
(mg/£)
IN WASTEWATER
299
600
25
-
117
36
0.23
0.53
<0£
2.710*
<0.1
11.6
RAW WASTE LOAD
kg/day
640
7 ,OOO
290
-
1.300
410
3
6
<6
31.100*
Owt
134
to/day
1,400
15.000
640
-
3.000
910
6
13
<13
—
<2.5
294
•Valua In plcoeiirtes/£
Value in picocurln/day
TABLE V-70. CHEMICAL COMPOSITION OF RAW WASTEWATER AND RAW WASTE
LOAD FOR MILL 9401 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COD
TOC
Alkalinity
Cu
Fe
Mn
Pb
At
Mo
V
Ra
U
Fluoride
CONCENTRATION
(rug/ >
IN WASTEWATER
294.000
65.6
460
12.200
<03
<0.3
17
19.000T
43.9
2.1
TOTAL WASTE
kg/day
3.200.000
160
1,216
32.940
<1.4
2JB
<0£4
<0.14
0.89
<0.81
46
51.300"
118
5.7
Ib/day
7.000.000
331
2.680
72.620
<3
6.5
<1.2
< 0.3
2
< 1.8
101
-
261
13
RAW WASTE LOAD
par unit ora milled
kg/ metric ton
1.000
0.047
0.38
10
< 0.00042
0.00078
< 0.0001 7
< 0.000042
0.00028
< 0.00025
0.014
16"
0.041
00018
Ib/fhort ton
2,000
0.094
0.76
21
<0.00084
0.0016
<0.00034
< 0.000084
0.00056
< 0.00051
0.029
33***
0.081
0.0035
par unit concentrate produced
kg/metric ton
550.000
26
211
5.720
<0.23
0.43
< 0.094
< 0.023
0.15
<0.14
8
8.870"
22
0.98
Ib/ihort ton
1.100.000
52
422
11,440
<047
036
<0.19
< 0.047
0.31
<0.28
16
20.100*"
45
2.0
•On the bant of 1973 production of 94.6% U,O8 and 5.5% VjOg
,*Value In picocuno/2.
Value in mierocuriai/day
Valua in mierocuriei/metric ton
Value in mfcrocunas/riiort ton
V-174
DRAFT
-------
DRAFT
TABLE V-71. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
LOAD FOR MILL 9402 (ACID- OR COMBINED ACID/ALKALINE-
MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Acidity
Al
Cu
Mn
Pb
At
Cr
Mo
V
Ra
U
COMCENTHATION
img/i )
IN WASTEWATER
525.000
63.5
24.0
35.000
1.594
2.7
105
2.1
23
9.0
16.0
125
234'
31.1
TOTAL WASTE
kg/day
4.100.000
337
127
185.700
8.460
14
557
11
12
48
85
663
1.240*
165
Ib/day
9.000.000
743
281
409.500
18.600
32
1.228
25
27
105
187
1.462
i ..
364
RAW WASTE LOAD
per unit ore milled
kg/metric ton
1.000
0.082
0.031
45
21
0003
0.14
0003
0.003
0.012
0.021
0.16
0.30"
0.040
Ib/short ton
2,000
0 16
0.062
91
4.1
0.007
0.27
0.005
0006
0023
0041
0.32
0.27"'
0.080
per unit concentrate produced*
kg/metric ton
450.000
37
14
20.400
930
1.6
61
1.2
1.3
5.2
9.3
73
136tr
18
Ib/short ton
900.000
74
28
40,800
1.860
31
122
24
27
10
187
146
124***
36
•On the basis of 1973 production of 98.2% U.O- and 1 8% MO,
A J O 3
Value in picocunes/2
"Value in microcuriei/day
tt
Value in microcuries/metnc Ion
•"Value in microcuries/shorl ton
V-175
DRAFT
-------
DRAFT
TABLE V-72. CHEMICAL COMPOSITION OF WASTE WATER AND RAW WASTE
LOAD FOR MILL 9403 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Alkalinity
Ca
MB
Ti
Al
Cu
ft
Mn
Ni
Pb
Zn
Al
Mo
V
Ra
Th
U
Fluoride
CONCENTRATION
(mg/ 1 1
IN WASTE WATER
111.000
278
<1
1.160.6
3.200
100
0.396
18
1 1
1 6
38
052
0.69
< 05
1 4
< 0.3
<05
111'
<0.1
39
1 4
TOTAL WASTE
kg/day
1.4OO.OOO
146
< 5.2
5.980
16.640
990
2.1
94
57
8.3
198
2.7
3.6
< 2.6
73
< 1 6
< 2.6
580"
•.05
20
7 3
Ib/day
3.100.000
319
< 11
13.190
36.680
2.180
45
206
13
18
436
6
79
< 57
16
< 34
< 57
-
<1
45
16
RAW WASTE LOAD
par unii on millad
kg/ metric ton
1.000
01
< 0.0037
43
12
071
00015
0.07
00041
0.0059
014
0.0019
00026
< 0.0019
00052
< 00011
< 0.0019
0.41"
< 00004
0032
00052
Ib/ihort ton
2.000
02
< 00074
85
24
14
00029
013
00081
0.012
028
0.0039
00051
< 00037
001
< 00022
< 0.0037
037"'
•.00008
0.064
001
par unit eoncantrata produced*
kg/matnc ton
1.050.000
109
39
4300
1.3
743
1.5
70
4.3
63
149
20
27
2
6.6
< 12
2
431 "
C0.4
34
55
Ib/ahort ton
2.100.000
217
78
9.000
25
1.486
31
141
86
13
297
41
5.4
4
11
< 23
4
392'"
-------
DRAFT
TABLE V-73. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
LOAD FOR MILL 9404 {ACID- OR COMBINED ACID/ALKALINE-
MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Acidity
C*
M»
Ti
Al
Cu
Fe
Mn
Ni
R»
Zn
At
Cr
Mo
V
Ra
U
CONCENTRATION
<0.5
0.13
2
< 0.3
120
690'
1745
TOTAL WASTE
kg/d.y
2.700X100
3,330
33
21,400
1.190
2320
16
3,920
36
1,720
1.110
73
45
<27
069
11
< 1.6
640
3.860"
925
Ib/day
6.000.000
7.350
72
47.200
2.620
6,430
35
8.660
79
3.800
2.450
16
98
< 58
15
23
< 35
1.400
-
2.035
RAW WASTE LOAD
per unit or* milled
kg/nwtric Ion
1.000
12
0.012
7.9
044
1 1
00059
1.5
00013
064
041
00027
OO016
< 000098
0.00026
00039
< 000059
0.24
135"
038
Ib/fhort Ion
2.000
2.5
0.024
ISA
0.88
2.2
0.012
29
0.0026
128
082
00054
00033
<0002
000051
00079
< 00012
047
1.23"*
0.75
(Mr unit concmtrM* produced*
kg/mMric ton
630,000
651
6.4
418
232
569
31
766
0.7
336
217
1 4
09
052
013
2.1
0.31
124
718"
180
Ib/thort ton
1 .060,000
1.300
12.8
836
464
1.139
62
1.532
1.4
673
435
29
1 7
10
0.27
4 1
062
248
652"-
361
•On trw brut ol 1973 production of 100X U-jOg
Vatue in picocimei/£
"Vilue in microcurMS/day
Value in nticrocuriet/mnric ton
"Value m microcunes/thon ton
V-177
DRAFT
-------
DRAFT
Water Uses. The primary use of water in each of these
Industries is in the beneficiation process, where it is
required for the operating conditions of the process. Water
is a primary material in the flotation of antimony, titanium,
and rare-earth minerals; in the leaching of beryllium ore;
in the concentration of titanium, zirconium, and rare-earth
minerals (monazite) from beach-sand deposits; and in
the extraction of platinum metals from placers by gravity
methods. No primary tin ore deposits of any commercial
significance are currently being mined in the U.S. However,
a small amount of tin is recovered as a byproduct of a
molybdenum operation through the use of flotation and magnetic
methods.
Water is Introduced into flotation processes at the ore grinding
stage to produce a slurry which is amenable to pumping,
sluicing, or classification for sizing and feed into the
flotation circuit. In leaching processes, water is the solvent
extraction medium. Water also serves as the medium for gravity
separation of heavy minerals.
In underground mining of antimony ore and in open-pit mining
of titanium and beryllium ores, water is not used directly
but, rather, is present (if at all) only as an Indirect con-
sequence of these mining operations. The mining of sand placer
deposits for titanium, zirconium, and rare-earth minerals is
done by dredging, in which a pond is required for flotation
of the barge. In mining a placer for platinum-group minerals,
a barge may be floated either in the stream or on an on-shore
pond, depending on the location of the ore.
Water flows of the antimony, beryllium, platinum, rare-earth
titanium, and zirconium mineral operations visited are presented
in Figures V-39, V-AO, and V-41.
Sources of Wastes. There are two basic sources of effluents
containing pollutants: those from mines or dredging operations
and the beneficiation process. Mines may be either open-pit
or underground operations. In the case of an open pit, the
source of the pit discharge (if any) is precipitation, runoff,
and ground-water seepage into the pit. Only one underground
mine was encountered in the SIC 1099 ore mining industry—an
antimony mine—and no existing discharges have been reported
V-178
DRAFT
-------
DRAFT
Figure V-39. WATER FLOWS AND USAGE FOR MINE/MILLS 9901 (ANTIMONY) AND
9902 (BERYLLIUM)
INO DISCHARGE)
m TO ma ™s««
MjOOO TO 100.000 gpdl
FLOTATION
MILL
288 TO 343 m3/d«y
TAILING-
POND
IMPOUNDMENT
EVAPORATION
AND
SEEPAGE
(NO
DISCHARGE)
(a) ANTIMONY MINE/MILL 9901
(b) BERYLLIUM MINE/MILL 9902
V-179
DRAFT
-------
DRAFT
Figure V-40. WATER FLOWS AND USAGE FOR MINE/MILLS 9903 (RARE EARTHS)
AND 9904 (PLATINUM)
TO ATMOSPHERE
to08m3/nUnim A
(20gpm) I
IAROE)
0.08 m3/mlnuti
(Zlgprnl
0 M m3/mlniiU
(06 gpm)
LEACH CONCENTRATION
t
f
1 6 m'/minuti
(412 gpm)
EVAPORATION
-./EVAPORATION M
EVAPORATION
I
0.08m3/ml«rt."V TOND J
(20 gpm) ^-
^1 TAILING \
1.88 m'/minut.
(618 gpm)
RECLAIM
TANK
^V POND )
0.2 m3/ffllnuu
61 gpm)
NOTE. FOR BYPRODUCT RECOVERY. SEE PART (b) OF FIGURE V-41 (MINE/MILL 89061
(a) RARE-EARTH MINE/MILL 9903
1
O.r •
2«.730 tn3/dty V^
— (6.480.000 gpd) ^^
24.730 m3/d«y
(8.480.000 gpd)
>REDQE
POND
J ^
^S *9-BO° "i3***
*"^ (12^80XXW gpd)
49^00 m3Miv
(12.960.000 gpd)
DREDG
BENEFI
EWITH
CIATION
(b) PLATINUM MINE/MILL 9904
V-180
DRAFT
-------
DRAFT
Figure V-41. WATER FLOWS AND USAGE FOR TITANIUM MINE/MILLS 9905 AND 9906
OPEN-PIT
MINE
2.668 m3/day
(699,000 gpd)
DISCHARGE
TO
RIVER
FLOTATION AND
MAGNETIC-SEPARATION
MILL
36,069 m3/day
(9,460.000 gpd)
INTERMITTENT
DISCHARGE (SEASONAL)
35.19 m3/day
(9.220,000 gpd)
PUMP
BASIN
o
878 m°/day
(230,000 gpd)
(a) TITANIUM MINE/MILL 9905
TO ATMOSPHERE
WET
MILL
(GRAVITY
SEPARATION)
1—OVERFLOW •
12.099 m3/day
(3,170.000 gpd)
BULK
'CONCENTRATE"
DRY
MILL
(ELECTROSTATIC
AND
MAGNETIC METHODS)
12,099 m3/day
7.633 m3/day
(2.000.000 gpd)
EVAPORATION
fuar >
)gpd)
12.595
(3.300,0
^^ OTO
1
m3/day
00 gpd)
17,175 m3/rfay
(4,500.000 gpd) I
RAIN AND TO
RUNOFF STREAM
(b) TITANIUM/ZIRCONIUM/MONAZITE MINE/MILL 9906
V-181
DRAFT
-------
DRAFT
at this time. Effluents from beach-sand dredging operations
orglnate as precipitation, runoff, and groundwater seepage.
In addition, effluents result from the fresh water used in
wet mill gravity beneficiation of the sands and, subsequently,
are usually discharged into dredge ponds.
The waste* constituents present in a mine or mill discharge
are functions of the mineralogy of the ores exploited and of
the milling or extraction processes and reagents employed.
Acid conditions prevailing at a mine site also affect the
waste components by Influencing the solubility of many
metallic components.
Wastewater from a placer or sand mining operation is primarily
water that was used in a primary or secondary gravity separation
process. Also, where a placer does not occur in a stream,
water is often used to fill a pond on which the barge is
floated. The process water is generally discharged into either
this pond or an on-shore settling pond. Effluents of the
settling pond usually are combined with the dredge-pond dis-
charge, and this comprises the final discharge. The principal
wastewater constituents from these operations are high suspended-
solid loadings and coloring due to high concentrations of
humlc acids and tannic acid from the decay of organic matter
Incorporated into former beach sands and gravels being mined.
Wastewater emanating from mills processing lode ores consists
almost entirely of process water. High suspended-solid loadings
are the most characteristic waste constituent of a mill waste
stream. This is primarily due to the necessity for fine
grinding of the ore to make it amenable to a particular benefi-
ciation process. In addition, the increased surface area of
the ground ore enhances the possibility for solubillzation of
the ore minerals and gangue. Although the total dissolved-
solld loading may not be extremely high, the dissolved heavy-
metal concentration may be relatively high as a result of the
highly mineralized ore being processed. These heavy metals,
the suspended solids, and process reagents present are the
principal waste constituents of a mill waste stream. In addition,
depending on the process conditions, the waste stream may
also have a high or low pH. The pH is of concern, not only
because of its potential toxlcity, but also because of its
effect on the solubility of the waste constituents.
Wastewater emanating from a beach-sand dredging pond consists
of water in excess of that needed to maintain the pond at the
V-182
DRAFT
-------
DRAFT
proper level. This water also originates as wet mill effluent
and, as a result, contains suspended solids. However, the
primary waste constituents from these milling operations are
the humic and tannlc acids which are Indigenous to the ore
body and which result in coloring of the water.
Description o£ Character and Quantity of_ Wastes
The quantity of wastes resulting from mining and milling
activities is discussed below individually for each of the SIC
1099 metals.
Antimony
Process Description - Antimony Mining. Currently, only one mine
exists which is operated solely for the recovery of antimony
ore (mine 9901). This ore is mined from an underground mine
by drifting (following the vein).
As indicated in Figure V-39, no discharge currently exists
from the mine.
Process Description - Antimony Milling. Only one mill is
operating for the recovery of antimony ore as the primary
product. This mill (9901) employs the froth flotation
process to concentrate the antimony sulfide mineral, stibnite
(Figure 111-28). The particular flotation reagents used by
this mill are listed in Table V-74. Water in this operation
is added between the crushing and grinding stages at the rate
of 305 to 382 cubic meters (80,000 to 100,000 gallons) per
day. There is no discharge, but flow to an impoundment totals
286 to 343 cubic meters (75,000 to 90,000 gallons) per day.
Quantities of_ Wastes. Waste constituents originate from two
sources: solubilizatlon and dispersion of ore constituents
and consumption of the milling reagents.
In metal mining and milling effluents, heavy-metal constituents
are of primary concern, due to their potentially toxic nature.
Metallic minerals known to occur with antimony in the commer-
cially valuable ore body of mine 9901 are:
V-183
DRAFT
-------
DRAFT
TABLE V-74. REAGENT USE AT ANTIMONY-ORE FLOTATION MILL 9901
REAGENT
Dowfroth 250 (Polypropylene glyool
methyl ethers)
Aerofloat 242 (Essentially Aryl
dithiophosphorie acids)
Ume (Calcium oxide)
Lead nitrate
PURPOSE
Frother
Collector
Depressant
Activating
Agent
CONSUMPTION
kg/metric ton
ore milled
0.4
0.1
2 to 3
05
Ib/short ton
ore milled
0.8
0.2
4 to 6
1.0
V-184
DRAFT
-------
DRAFT
Stibnite (Sb2S3)
Pyrite (FeS2)
Arsenopyrite (FeAsS)
Sphalerite (ZnS)
Argentite (Ag2S)
Cinnabar (HgS)
Galena (pbS)
The metals in these minerals are the ones which would be
expected to occur at highest concentrations in the waste
stream, and results of raw-waste analysis support this
conclusion (Table V-75). The raw-waste characterization
presented in Table V-75 is based upon the analysis of
samples collected during the mill visit. As would be
expected on the basis of the mineralization of the ore body,
the metals present at relatively high concentrations in the
raw waste are antimony (64.0 mg/1), zinc (4.35 mg/1), and
iron (18.8 mg/1). Arsenic is not as high as was expected
but is about an order of magnitude greater than mean back-
ground levels reported in surface waters of the Pacific North-
west Basin. Waste loadings for important constituents of
wastewaters from mill 9901 are listed in Table V-76.
Beryllium
Process Description - Beryllium Mining. Beryllium ore is
mined on a large scale at only one domestic operation. At
mine 9902, bertrandite (H2Be4Si209) is recovered by open-pit
methods. A small amount of beryl is also mined in the U.S.
by crude open-cut and hand-picking methods. As indicated
in Figure V-39, no discharge currently exists at mine 9902.
Process Description - Beryllium Milling. Currently, only
one domestic beryllium operation uses water in a beneficia-
tion process. This operation is identified as mill 9902 and
employs a proprietary acid leach process to concentrate
beryllium oxide from the ore.
Quantities of Wastes. As indicated in Figure V-31, approxi-
mately 3,061 cubic meters (802,000 gallons) per day of waste-
water are discharged from mill 9902. Waste constituents
V-185
DRAFT
-------
DRAFT
TABLE V-75. CHEMICAL COMPOSITION OF RAW WASTEWATER DISCHARGED FROM
ANTIMONY FLOTATION MILL 9901
PARAMETER
PH
Acidity
Alkalinity
Color
Turbidity (JTU)
TSS
TDS
Hardness
Chloride
COD
TOC
Al
As
Be
Be
B
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
Total Mn
CONCENTRATION (mg/£l
BJ*
8.6
11.0
113*
170
149
68
40
14
43
7JB
6.2
0.23
<0402
<041
0.103
0.67
OM
0.12
184
0.13
143
040
PARAMETER
Hg
Ni
n
v
K
Se
Afl
Na
Sr
Ta
Ti
Zn
Sb
Mo
Oil and Graaw
MBAS Surfactants
Cyanide
Phenol
Fluoride
Total Kjehbhl N
Sulflde
Sulfata
Nitrate
Phosphate
CONCENTRATION (mg/JU
0.0038
0.10
<046
<0.2
0.036
<0.02
2A
0.11
4J5
64JO
14
041
0422
164
246
045
•Value In pH unto
f Value in cobalt units
V-186
DRAFT
-------
DRAFT
TABLE V-76. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD
AT ANTIMONY MILL 9901
PARAMETER
PH
TSS
COD
TOC
Fa
Pb
Sb
Zn
Cu
Mn
Mo
CONCENTRATION
(mg/e) IN
WASTEWATER
8.3»
997
43
7.8
18.8
0.13
64.0
4.35
0.12
0.40
<0.2
RAW WASTE LOAD
per unit concentrate produced
kg/metric ton
—
74.78
3.22
0.585
1.41
0.0097
4.8
0.366
0.009
0.03
< 0.01 5
Ib/thort ton
—
149.56
6.44
1.170
2.82
0.0194
9.6
0.652
0.018
0.06
<0.030
per unit ore milled
kg/metric ton
—
7.48
0.0322
0.059
0.141
0.00097
0.48
0.033
0.0009
0.003
< 0.001 5
Ib/short ton
—
14.96
0.0644
0.118
0.282
0.00194
0.96
0.066
0.0018
0.006
< 0.0030
•Value in pH units
V-187
DRAFT
-------
DRAFT
originate from two sources: solubilization and dispersion
of ore constituents and consumption of milling reagents.
However, because this process Involves acid leaching, high
solubilization is observed in the waste constituents (Table
V-22).
The mineralization of the ore body from which bertrandite is
obtained is essentially that presented in the tabulation
given below for mine 9902 (beryllium) .
Quartz 8102^
Feldspar Al silicates with Ca, K, and Na
Fluorite CaF2_
Carbonates
Iron Oxide Minerals
Tourmaline (XY3A16_(B03)3_(Si6018) (OH)4)
where X = Na, CaT Y = Al, Fe(+3), Li, Mg
Constituents of-these minerals are also expected to be the
main constituents in the mill waste, and results of waste
analysis support this (Table V-77). As indicated, the waste
stream from this leaching process is exceptionally high in
dissolved solids (18,380 mg/1), consisting largely of sulfate
(10,600 mg/1). Fluoride (45 mg/1) is also present at rela-
tively high concentration, as are aluminum (552 mg/1),
beryllium (36 mg/1), and zinc (19 mg/1). Raw waste loads
are not presented because of the proprietary nature of the
process and production and because no effluent results at
this unique operation.
Rare Earths
Process Description - Rare-Earth Metals Mining. The rare-
earth mineral monazite (Ce, La, Th, Y)P04_) is recovered
predominantly as a byproduct from sand placers mined by
dredging—primarily, for their titanium mineral content.
(Refer to information on mill 9906, as described for titanium.)
The rare-earth mineral bastnaesite Is also currently recovered,
as the primary product, by an operation mining the ore from
an open-pit mine (mine 9903).
As indicated in Figure V-40, no discharge currently exists
at mine 9903.
V-188
DRAFT
-------
DRAFT
TABLE V-77. CHEMICAL COMPOSITION OF RAW WASTEWATER FROM
BERYLLIUM MILL 9902 (NO DISCHARGE FROM TREATMENT)
PARAMETER
Conductivity
Color
Turbidity MTU)
TDS
Acidity
Alkalinity
Hardnm
COO
TOC
OllandGrean
MBAS Surfactant*
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
Pb
Mg
CONCENTRATION (mg/£)
17.000*
88'
1.3
18,380
3J03B
0
4.000
22
55
<1
0.78
552
0.16
36.0
<6X>
OJK
0.047
43.0
0.20
QJ07
<0.1
320X1
PARAMETER
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Tl
Zn
Mo
Chloride
Fluoride
Sulfate
Nitrate
PtuT.B.T.I.Otlll
rnoifHiavo
Cyanide
Phenol
Total Kfeldahl N
Sulflde
CONCENTRATION (mg/JK)
49JO
0.15
< OXK
<02
77 JO
OJJ62
OJ04
270.0
022
<0£
19X1
< 02
170
46
10,600
126
03
< 0.01
< 0X11
0.19
< 0.5
•Value In mlcromhot/cm
fValue In cobalt unttt
V-189
DRAFT
-------
DRAFT
Process Description - Milling. Monazite is concentrated by
the wet gravity and electrostatic and magnetic separation
methods, discussed in the titanium segment of this section.
A single mill (9903) is currently beneficiating rare-earth
minerals mined from a lode deposit. These rare-earth minerals,
bastnaesite and some monazite, are initially concentrated by
the froth flotation process (Figure V-42). Flotation of
rare-earth minerals requires rigidly controlled conditions and
a pH of 8.95, and temperature-controlled reagent addition is
critical to the successful flotation of these minerals.
Rare-earth oxides (REO) in the mill- heads range from 6 to 11
percent and are upgraded in the flotation circuit to a con-
centrate that averages 57 to 65 percent REO, depending upon
the heads. This concentrate is leached with hydrochloric
acid to remove calcium and strontium carbonates, increasing
the REO content in the leached concentrate by as much as 5
to 10 percent. This concentrate is processed In a solvent
extraction plant to produce high-purity europium and yttrium
oxides; a cerium hydrate product; a concentrate of lanthanum,
praesodymium and neodymium; and a concentrate of samarium and
gadolinium (Figure V-43).
In the solvent extraction plant, the flotation concentrate is
initially dried and then roasted to remove carbon dioxide and
to convert the rare-earths to oxides. These oxides, with
the exception of cerium oxide, are converted to soluble chlorides
in a hydrochloric-acid leaching circuit. Following leaching,
the acid slurry is passed through a countercurrent decanta-
tlon circuit. The primary thickener overflow containing the
chlorides is fed into the europium circuit, while the leached
solids from the countercurrent decantatlon circuit make up
the feed for the cerium process.
The leach liquor (primary thickener overflow) is clarified
in a carbon filter and adjusted to a pH of 1.0 and a tempera-
ture of 60 degrees Celsius (140 degrees Fahrenheit) prior to
countercurrent extraction of europium with organic solvent
(90 percent kerosene and 10 percent ethyl/hexyl phosphoric
acid) . The raffinate from the extraction circuit makes up the
feed for the lanthanum circuit, which is discussed later.
V-190
DRAFT
-------
DRAFT
Figure V-42. BENEFICIATION OF BERTRANDITE, MINED FROM A LODE DEPOSIT,
BY FLOTATION (MILL 9903)
FRESH
WATER
OM «3/mfe 1M mj/mtmm
g1S»»il f
EQUIPMENT
CRUSHING
GRINDING
CYCLONE
UNDERFLOW
t
_ I ,
1 'G
JHJOH4NTEKSITY
CLONE!
FIRST C
1 FLOTATI
T5
'
SECOND. THIRD
FLOTATIC
UNDEI
tCAVE
FLOTATIO
i
-FROTH— '
i
—FROTH 1 FIRS1
»— FROTH
:LEANER
ON CELLS
TH
'
. AND FOURTH «
HER — w
N CELL*
IFLOW
NOER
N CELLS
LASBIFICATION AND MOLYBDENUM
I
CONDITIONING P* 1 CONDITIONING l*~1 O
1 *C
L«»»FI
I '
AMD SECOND ROUOHEB .. ,
FLOTATION CELLS
UNDERFLOW
* ^~-
iF^TATlSNTErLh0"0"'10— K. ™
[HYDROCHLORIC 1
ACID |
|0™ irL-.H
MCbNIHAlb I"™* AmTATQRS
1
ALTERNATIVE
1
I
f
1
i • f SODIUM 1
IORZAN | | CARBONATE |
t » t
»NO STAGE OF 1 ^ 1 FIRST STAGE OF 1
DNDITIONING f*~~| CONDITIONING |
1 0»C i
||1SO»FI J
'
i 1
v
ING >
NO J
—--^
1M n'/mlniti
(SlSgpml
1 fc LEACHED CONCENTRATE _».TO
THICKENER WAS
1 DRYER 1 1
LJL-J ,
ALTERNATIVE
1
FINAL
PRODUCT
SEPARATION OF RARE
EARTH METALS BY
SOLVENT EXTRACTION
(SEE FIGURE V43)
TO STOCKPILE
V-191
DRAFT
-------
o
a
>
•n
VO
tVJ
Figure V-43. BENEFICIATION OF RARE-EARTH FLOTATION CONCENTRATE BY
SOLVENT EXTRACTION (MILL 9903}
ILT
4
"t
!»« OVERFLOW
FILTER *~l THICKENER 1
1 1
i
, f
DRYER | ""'SiV'0" •*— | AMMONIA ]
JAR
Fl
NU
YDS
STW
. I 1
IE OR f
ANUM CARBON
«« F'IT" CUROP.UM
J PRODUCT
LE
| THICKENER «,« 1
1
SODIUM HVDROSULFIDE
AND AMMONIA
*
PRECIPITATION ^ R«flNAT[
FILTRATE
DRUM
FILTER
t
I
GA
SODIUM CARBON
1
GADOIIMIUM/SAMA R IUM
PRECIPITATION TANK
EU
PURI
J
'
tOPIUM
•(CATION
J
-------
DRAFT
After loading the organic with europium, the europium is
stripped in the solvent extraction strip circuit with AN hydro-
chloric acid. The pregnant strip solution contains iron,
which is removed in precipitation tanks by the addition of soda
ash to lower the pH to 3.0 to 3.5. This causes ferric hydrox-
ide to precipitate, and the precipitate is removed in a
pressure filter. Following removal of the iron, the europium-
bearing solution goes through another solvent extraction and
stripping circuit, similar to the previous one. The pregnant
strip is pumped to a purification circuit, where europium
oxide is prepared for the market.
Solutions from the purification circuit are neutralized with
sodium carbonate to produce gadolinium and samarium carbonates,
which are collected by a drum filter.
Returning to the countercurrent decantation circuit, the
solids remaining from leaching are filtered and repulped.
The cerium solids are then thickened, filtered, and dried to
produce the final concentrate.
As mentioned previously, the raffinate from the first solvent
extraction circuit provides the feed for the lanthanum circuit.
This raffinate is clarified in a carbon filter, and ammonia
is added to precipitate lanthanum hydrate. The precipitate is
thickened and filtered to produce the final concentrate.
Quantities of Wastes. As indicated in Figure V-40, raw wastes
are discharged at a rate of 1.96 cubic meters (518 gallons)
per minute from the flotation circuit and at a rate of 0.08
cubic meter (21 gallons) per minute from the leach/solvent
extraction plant. These waste streams are not combined, and
both are characterized in Table V-78. These data are based
upon the analysis of raw-waste samples collected during the
mill visit. Table V-79 presents the results of chemical
analyses for the rare-earth metals.
Reagents used in the flotation, leach, and solvent extraction
processes of mill 9903 are identified below.
V-193
DRAFT
-------
DRAFT
TABLE V-78. CHEMICAL COMPOSITION OF RAW WASTEWATER
FROM RARE-EARTH MILL 9903
PARAMETER
PH
Acidity
Alkalinity
Color
Turbidity (JTU)
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
s.o2
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/JU
FLOTATION
9.02"
-
m
•
14/476
380.000
-
-
3,100
.
-
•
-
-
.
-
-
0.35
.
-
LEACH/
SOLVENT
EXTRACTION
8-23*
345
2.125
80t
52
76.162
786
7,220
>1.500
47
<1
212
126
<0.1
0.01
0.009
<0.01
< 0.005
2.910
0.04
<0.03
0.03
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Chloride
Fluoride
Sulfate
Nitrate
Phosphsto
Cyanide
Phenol
CONCENTRATION (mg/£)
FLOTATION
.
•
0.6
-
.
< 0.3
-
.
-
-
-
-
•
-
-
.
365
.
-
-
-
-
LEACH/
SOLVENT
EXTRACTION
<0.05
6.6
3jO
0.85
<0.1
<03
94
0.015
OJ09
650
4.6
3.36
7 JO
<0.1
54.000
<0.1
23
1.60
0.09
<0.01
<0.01
•Value in pH units
Value In cobalt units
V-194
DRAFT
-------
DRAFT
TABLE V-79. RESULTS OF CHEMICAL ANALYSIS FOR RARE-
EARTH METALS (MILL 9903-NO DISCHARGE)
PARAMETER
Y
La
Ce
Pr
Nd
Sm
Eu
Gd
Th
CONCENTRATION (mg/ £ )
LEACH WASTEWATER
_
442
24
6.2
9.6
0.27
< 0.001
< 0.001
< 0.001
FLOTATION RECLAIM WATER
0.014
1.32
2.75
027
0.51
0.041
< 0.001
0.006
< 0.001
V-195
DRAFT
-------
DRAFT
Flotation Circuit
Frother Methylisobutylcarbinol
Collector N-80 Oleic Acid
pH Modifier Sodium Carbonate
Depressants Orzan, Sodium Silicofluroide
Conditioning Agent Molybdenum Compound
Leach Circuit
Leaching Agent Hydrochloric Acid
Solvent-Extraction Circuit
Leaching Agent Hydrochloric Acid
Precipitants Sodium Carbonate, Ammonia, Sodium Hydrosulfide
Solvents Kerosene, Ethyl/Hexyl Phosphoric Acid
In rare-earth metal mining and milling, effluent constituents
expected to be present are a function of the mineralogy of the
ore and the associated minerals. The principal minerals
associated with the ore body of mine 9903 are: bastnaesite
(CeFCOS., with La, Nd, Pr, Sm, Gd, and Eu); barite (BaS04_);
calcite (CaC03^; and strontianite (SrC03).
The dissolved-solid content of the leach/solvent-extraction
waste stream are extremely high (76,162 mg/1) and are due
largely to chlorides (54,000 mg/1). The metals present at
highest concentrations are those which would be expected on
the basis of known mineralization and use in the process.
These are strontium (4.5 mg/1) and barium (less than 10 mg/1).
The high concentration of tellurium (3.36 mg/1) is unexplained
on the basis of known mineralization, but mineralization is
assumed to be the source of this element. Waste characteristics
and raw waste loading for the rare-earth flotation and concentrate
leaching/solvent extraction processes are given in Table V-80.
Platinum-Group Metals
Process Description - Platinum Mining. Production of platinum-
group metals is largely as a byproduct of gold and copper
refining, and primary ore mining is limited to a single dredging
operation (mine 9904) , which is recovering platinum-metal
alloys and minerals from a placer deposit.
V-196
DRAFT
-------
DRAFT
TABLE V-80. CHEMICAL COMPOSITION AND RAW WASTE LOAD FROM RARE-EARTH
MILL 9903
PARAMETER
CONCENTRATION
{mg/£) IN
WASTEWATER
RAW WASTE LOAD f
per unit of concentrate
kg/metric ton
Ib/thort ton
per unit ore milled
kg/metric ton | Ib/short ton
(a) Flotation Mill
pH
TSS
TOC
Cr
Mn
V
Fluoride
9.02*
360.000
3,100
0.35
0.5
<0.3
365
—
9,335
80.4
0.009
0.013
< 0.0078
9.46
—
18,670
160.8
0.018
0.026
< 0.01 6
18.93
—
933.5
8.04
0.0009
0.0013
<0.0008
0.95
—
1.867.0
16.08
0.0018
0.0026
< 0.001 6
1.89
(b) Leach/Solrant-Exchange Mill
pH
TSS
TOC
Si02
Cr
Mn
V
Te
Ni
8.23*
786
47
1.25
0.04
3.0
<0.3
3.36
0.85
_
0.833
0.047
0.00125
0.00004
0.003
C0.0003
0.003
0.001
_
1.67
0.094
0.00250
0.00008
0.006
< 0.0006
0.006
0.002
_
—
—
—
—
—
—
—
—
_
—
—
—
—
—
—
—
-
Value in pH units
Based upon maximum production achievable (part a) or estimated amount of flotation concentrate
produced (part b)
V-197
DRAFT
-------
DRAFT
Process Description - Milling. Mill 9904 employs a physical
separation process to beneficiate the placer gravels (Figure
111-20). The dredged gravels are intially screened, jigged,
and tabled to separate the heavy minerals from the nonmineral
lights, which are discarded. Chromite and magnetite are sep-
arated from the platinum-group metal alloys and minerals by
magnetic separation. The final platinum-group metal concen-
trate is produced from the magnetic-separation product by dry
screening and passing the resultant material through a blower
to remove the remaining lights.
Quantities of Wastes . Wastes resulting from the mining and
milling activities of this operation cannot be considered
separately, since the wet mill discharges to the dredge pond.
No reagents are required in the milling process, and, as a
result, the principal waste constituent from this operation
is suspended solids (30 mg/1). Table V-81 lists the chemical
composition of the wastewater and waste loads from mine/mill
9904.
As indicated in Figure V-40, 24,700 cubic meters (6.5 million
gallons) per day of water are discharged from the dredge pond
to the river. The wet milling process utilizes 49,500 cubic
meters (12.96 million gallons) per day.
The principal associated minerals in this placer (mine 9904)
are:
Chromite (FeCr2_04)
Ferroplatinum (Fe, Ft, Ir, Os, Ru, Rh, Pd, Cu, Ni) alloy
Iridium/ruthenium/osmium alloy
Taurite (Ru, Ir, Os)S2_
Unnamed mineral (Ir, Rh, Pd)S
Mertieite (Pt^(Sb, As)2_)
Sperrylite (PtAs2_)
Gold (Au)
Tin
Tin is recovered in the U.S. only as a byproduct of a molyb-
denum operation. At this mine (6102), the ore is mined by
glory-hole methods, in which the sides of an open hole are
V-198
DRAFT
-------
DRAFT
TABLE V-81. CHEMICAL COMPOSITION AND LOADING FOR PRINCIPAL WASTE
CONSTITUENTS RESULTING FROM PLATINUM MINE/MILL 9904
(INDUSTRY DATA)
PARAMETER
Alkalinity
Conductivity
Hardness
COD
BOD
TS
TDS
TSS
(N) NH3
Kjaldahl Nitrogen
Al
Cd
Cr
Cu
Total Fe
Pb
Zn
Chloride
Fluoride
Nitrate
Sulfate
Sulfide
CONCENTRATION
(mg/U
WASTEWATER
83
109*
35.6
7.6
35
82
52
30
0.18
0.28
0.337
< 0.001
<1.0
<1.0
0.166
0.010
0.028
11.0
0.95
4.5
5.5
1.2
RAW WASTE LOAD
per unit ore milled
kg/1000 metric tons
1.20
-
0.51
0.11
0.05
1.18
0.75
0.43
0.003
0.004
0.005
< 0.00001
<0.01
<0.01
0.002
0.0001
0.0004
0.16
0.01
0.06
0.08
0.02
lb/1000 short tons
2.39
-
1.03
0.22
0.10
2.36
1.50
0.86
0.006
0.008
0.010
< 0.00002
<0.03
<0.03
0.005
0.0003
0.0008
0.32
0.01
0.13
0.16
0.03
•Value in micromhos/cm
TS = Total Solids
V-199
DRAFT
-------
DRAFT
caved and the broken rock trammed out through a tunnel at
the bottom of the hole. No specific waste characteristics
and water uses can, therefore, be assigned to tin mining and
milling.
Titanium
Process Description - Mining. Titanium minerals are recovered
from lode and sand deposits. The single lode deposit being
exploited in the U.S. is mined by open-pit methods at mine
9905. Ancient beach-sand placers are mined at several opera-
tions by dredging methods. In these operations, a pond is
constructed above the ore body, and a dredge is floated on
the pond. The dredges currently used normally are equipped
with suction head cutters to mine the mineral sands. Wastes
from dredge ponds and wet mills are combined; therefore,
these operations are discussed under one heading: Dredging
Operations.
Quantities of_ Wastes: Mine 9905. This is the only existing
mine from which titanium lode ore is mined. Water is discharged
from this open pit at a rate of 2,668 cubic meters (699,000
gallons) per day. The chemical composition of this waste is
presented in Table V-82. As these data show, oils and grease
(3.0 mg/1), fluorides (3.20 mg/1), total KJeldahl nitrogen
(2.24 mg/1), and nitrates (15.52 mg/1) are present at relatively
high concentrations. The oils and greases undoubtably result
from the heavy equipment used in the mining operations, and
the fluorides are probably indigenous to the ore body. How-
ever, the reason for the high concentrations of nitrogen and
nitrates may be explained in part by the use of nitrate-based
blasting agents.
Process Description - Titanium Milling; Mill 9905. Ore
brought to this mill is beneficiated by a combination of
the magnetic-separation and flotation processes (Figure V-44).
The ore is initially crushed and then screened. Both the
undersize and the oversize screened ores are magnetically
cobbed to remove the nonmagnetic rock, which is discarded.
Oversize magnetic rock undergoes further crushing and
screening, while undersize material is fed into the grinding
V-200
DRAFT
-------
DRAFT
TABLE V-82. CHEMICAL COMPOSITION OF RAW WASTE WATER
FROM TITANIUM MINE 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hardnest
COD
TOG
Oil and Create
MBAS Surfactants
Total Kjaldahl N
Al
A»
Be
Ba
a
Cd
Ca
Cr
Cu
Total Fa
CONCENTRATION (mg/£)
1.000-
11.3*
037
1.240
14
BA
1382
6464
6.4
10J
3.0
0.32
2.24
0.1
0.1
0.003
<1
0.01
< 0.002
94.6
<0.03
0.33
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Sr
Ao
Ha
Sa
Ta
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfate
»!••• tn
HIlrBlv
ft I.. .•..•. n *iii
rnoXfirian
CONCENTRATION (mg/e)
<0.1
<041
140 JO
0.76
<0i
OJ007
<0.1
<0.1
-------
DRAFT
Figure V-44. BENEFICIATION AND WASTE WATER FLOW OF ILMENITE
MINE/MILL 9905 (ROCK DEPOSIT)
MAKEUP ^V _
WAIfcH J 878n,3/dav*
' (232.000 gpd)
36.069 m3/doy
(9.450.000 gpd)
i— MAGN
MAGNETITE
, * ,
DEWATERER
•-FILTRA1
c
WATER
"RETURN TO MILL"
35.191 m3/day
(8 220 000 and) l"~
MINING
ORE
CRUSHING
\
GRINDING
\
CLASSIFICATION
1
MAGNETIC
SEPARATION
ETICS Z-NONMA
re-J
^
TAILINGS
1
SNETICS-i
ILMENITE
AND GANGUE
1
FLOTATION
CIRCUIT
1
THICKENER
-1-. \
TAILING A
POND )
^^ ^^^
FILTER
-=T 1
i
*
DRIER
SEASONAL I -1
DISCHARGE 1
CONCENTRATE
TO SHIPPING
V-202
DRAFT
-------
DRAFT
circuit. The latter utilizes grinding in rod mills, which are
in circuit with "Ty Hukki" classifiers. Final grinding of the
undersize material is done in a ball mill.
The magnetite and ilmenite fractions are magnetically sepa-
rated, with the magnetite further upgraded by additional
magnetic processing. The Ilmenite sands are then upgraded in
a flotation circuit consisting of roughers and three stages
of cleaners. The iloenite concentrate is filtered and dried
prior to shipping.
Quantities of_ Wastes: Mill 9905. Wastes are discharged
from this mill at a rate of 35,191 cubic meters (9,220,000
gallons) per day. The results of a chemical analysis of this
wastewater are presented in Table V-83. These data are based
on analysis of raw waste samples collected during the mill
visit.
Reagents consumed in the flotation circuit of mill 9905 are
identified in Table V-84. The principal associated minerals
in the ore body of mine 9905 are listed in Table V-85. These
reagents and constituents of the ore body comprise the princi-
pal constituents of the waste stream.
As indicated in Table V-84, relatively high levels of iron,
titanium, zinc, nickel, vanadium, chromium, and selenium were
observed in the wastes of mill 9905. Table V-86 is a compila-
tion of the concentrations and the raw waste loading of the
principal constituents of raw wastewater from mill 9905.
Titanium
Dredging Operations; Mill 9906 and 9907. These operations
are representative of the operations which recover titanium
minerals from beach-sand placers. Operations 9906 and 9907
utilize a dredge, floating on a pond, to feed the sands to a
wet mill (Figure V-45). The sands are beneficiated in the
wet mill by gravity methods, and the bulk concentrate is sent
to a dry mill for separation and upgrading of the heavy
minerals. As Indicated in Figure V-41, for mill 9906, no
discharge exists from the dry mill. Water used in the wet
mill is discharged to the dredge pond, which subsequently
discharges at a rate of 12,099 cubic meters (3.17 million
gallons) per day. Raw waste characterization of the combined
V-203
DRAFT
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TABLE V-83. CHEMICAL COMPOSITION OF RAW WASTE WATER
FROM TITANIUM MILL 9905
PARAMETER
Conductivity
Color
Turblditv (JTU)
TOS
TSS
Acidity
Alkalinity
Hardnen
COD
TOC
Oil and Greats
IMBAS Surfactants
Total Kjeldahl N
Al
At
Ba
B
Cd
Ca
Cr
Cu
Total Fa
CONCENTRATION (mull)
650*
18.0*
13.
518
26,300
6.0
81.4
344.8
< 1.6
9.0
2.0
0.04
0.65
210
< 0.01
< 0.002
< 0.01
< 0.002
360
0.58
0.43
500
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Phenol
Chlorida
Fluoride
Sulfata
Nitrate
Phosphate
CONCENTRATION (mg/£)
< 0.05
187.5
5.9
1.19
<0.1
2.0
23.7
0.132
0.015
41
0.29
< OJ06
2.08
7.6
< 0.1
< 0.1
< 041
19.1
325
213
ojea
< 0.05
•Value in mlcromhot/cm
* Value in cobalt units
TABLE V-84. REAGENT USE IN FLOTATION CIRCUIT OF MILL 9905
REAGENT
Tall oil
Fuel oil
Methyl amyl alcohol
Sodium bifluoride
Sulfuric acid
PURPOSE
Frother
Frother
Frother
Depressant
pH Modifier
CONSUMPTION
kg/metric ton
ore milled
1.33
0.90
0.008
0.76
1.775
Ib/short ton
ore milled
2.66
1.80
0.016
1.52
3.55
V-204
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TABLE V-85. PRINCIPAL MINERALS ASSOCIATED WITH
ORE OF MINE 9905
MINERAL
llmenite
Magnetite
Pyroxene
Feldspar
COMPOSITION
FeTiOa
FeaOA
Complex Ferromagnesium Silicate
Aluminum Silicates with Calcium,
Sodium, and Potassium
TABLE V-86. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD AT MILL 9905
PARAMETER
TSS
TOC
Ni
Ti
Fe
V
Cr
Mn
Se
Cu
Zn
Fluoride
CONCENTRATION
(mB/K) IN
WASTEWATER
26,300
9.0
1.19
2.08
500
2.0
0.58
5.9
0.132
0.43
7.6
32.5
RAW WASTE LOAD
per unit concentrate produced
kg/metric ton
462.8
0.158
0.021
0.036
8.8
0.035
0.010
0.103
0.0002
0.008
0.133
0.569
Ib/short ton
925.8
0.316
0.042
0.072
17.6
0.070
0.020
0.206
0.0004
0.016
0.266
1.14
per unit ore milled
kg/metric ton
210.4
0.072
0.01
0.017
4.0
0.016
0.005
0.048
0.001
0.0003
0.061
0.26
Ib/short ton
420.8
0.144
0.02
0.034
8.0
0.032
0.01
0.096
0.002
0.0006
0.122
0.52
V-205
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Figure V-45. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS (RUTILE, ILMENITE,
ZIRCON, AND MONAZITE) AT MILL 9906
TO
ATMOSPHERE
ORE + WATER
t
20,100 m/day
(5,310.000 gpd)
ORE FED
FROM DREDGE
i
EVAPORATION
POND
WATER
RECYCLE
H2O
_*_
VIBRATING
SCREENS
7,570 mj/day
(2,000,000 gpd)
—OVERSIZE-
WASH
'WATER'
SPIRALS OR LAMINAR
FLOWS (ROUGHERS AND CLEANERS)
11.000 m3/day
(3.170,000 gpd)
T
—TAILINGSn
12,000 m3/day
(3.170,000 gpd)
TO DRY MILL
(FIGURE 111-30)
rWASTE (DREDGE)]
POND
15.615 m3/day
(4,500,000 gpd)
DISCHARGE
V-206
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wet-mill and dredge-pond discharge is presented in Table V-87.
These data are based on analysis of raw waste samples collected
during the visits to these operations.
No reagents are used in the beneficiation of the sands, as
gravity methods are employed in the wet mill, and magnetic and
electrostatic methods are used in the dry mill. Therefore,
the principal waste constituents, with the exception of waste
lubrication oil from the dredge and wet mill, are influenced
primarily by the ore characteristics. The ore bodies of opera-
tions 9906 and 9907 contain organic material which, upon
disturbance, forms a colloidal slime of high coloring capacity.
This organic colloid—primarily, humates and tannic acid—and
the wasted oil are the principal waste constituents of the pond
discharges. This is reflected in the high carbon oxygen
demand (COD) and total organic carbon (TOC) values detected in
the waste streams of operations 9906 and 9907 (Table V-87).
High levels of phosphate and organic nitrogen are present in
these waste streams also. The phosphate and nitrogen are
undoubtedly associated with the sediments in the ore body.
Raw waste loads of principal wastewater constituents discharged
from the milling operations at mills 9906 and 9907 are given
in Table V-88.
Zirconium
Zirconium is recovered as a byproduct of the mining and milling
of sand placer deposits, which have been described under Waste
Characteristics of Titanium Ores. No operations for zirconium
alone are known in the United States. The waste characteristics
and water uses accompanying mining and milling to obtain
zircon concentrate are, therefore, identical to those of
the previously described operations.
V-207
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TABLE V-87. CHEMICAL COMPOSITION OF RAW WASTE WATER
AT MILLS 9906 AND 9907
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
T8S
Acidity
Alkalinity
COD
TOC
Total KMdahl N
Oil and Greaw
MBAS Surfactant*
Al
Aa
Be
Be
B
Cd
Ca
Cr
Cu
CONCENTRATION (ma/Jll
MILL 9906
200*
51,400*
<0.1
1,644
11,000
47.2
47.6
1^38
972
0.66
400
<0.01
69.0
0.06
< 0.002
<0.6
0.10
< 0:002
0.10
QJ03
<0.01
16.0
0.03
<0.002
<0.05
<0.01
<0.03
PARAMETER
Total Fe
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Chloride
Fluoride
Phosphate
Phenol
CONCENTRATION imgfU
MILL 9906
4A
1
<0.1
<0in
27.0
<0.06
<0i
0.014
<0.1
<0.1
30.0
0.03
0.35
<0.01
MILL ((07
OA3
COJOS
0.66
0.01
<0£1
<0.1
<0£
13
<0j05
<0.01
6A
<0.1
<0.1
15X)
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DRAFT
TABLE V-88. RAW WASTE LOADS FOR PRINCIPAL WASTEWATER CONSTITUENTS
FROM SAND PLACER MILLS 9906 AND 9907
PARAMETER
TSS
TOC
COD
Oil and Grwu
Ti
Fe
Mn
Cr
Phaplun
MILL 0906
CONCENTRATION
(mo/ 11
IN WASTEWATER
11.000
872
1.337
400
<0.2
4.9
0.36
0.03
036
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
330
28.2
40.13
12
< 0.006
0.15
0.0011
00009
0.011
In/then ton
660
68.4
80.26
24
< 0.012
0.30
0.0022
0.0018
0022
MILL 9907
CONCENTRATION
(mg/ei
IN WASTEWATER
209
321
361.6
40
0.4
0.93
<0.01
<0.01
0.4
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
6.01
7.71
8.68
036
0.01
0.022
< 0.0024
< 0.0024
0X11
Ih/ihortton
10.02
15.42
17.36
1.92
0.02
0.044
< 0.0048
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DRAFT
SECTION VI
SELECTION OF POLLUTANT PARAMETERS
INTRODUCTION
The water-quality investigation which preceded development
of recommended effluent guidelines covered a wide range of
potential pollutants. After considerable study, a list of
tentative control parameters was prepared for each category
and subcategory represented in this study. The wastewater
constituents finally selected as being of pollution signifi-
cance for the ore mining and dressing industry are based upon
(1) those parameters which have been Identified as known con-
stituents of the ore-bearing deposits and overburden, (2)
chemicals used in processing or extracting the desired
metal(s), and (3) parameters which have been identified as
present in significant quantities in the untreated wastewater
from each subcategory of this study. The wastewater constit-
uents are further divided into (a) those that have been
selected as pollutants of significance (with the rationale
for their selection), and (b) those that are not deemed
significant (with the rationale for their rejection). This
Section is concluded with a summary list of the pollution
parameters selected for each category.
GUIDELINE PARAMETER-SELECTION CRITERIA-
Selection of parameters for use in developing effluent limita-
tion guidelines was based primarily on the following criteria:
(1) Constituents which are frequently present in mine
and mill discharges in concentrations deleterious
to human, animal, fish, and aquatic organisms (either
directly or indirectly).
(2) The existence of technology for the reduction or
removal, at an economically achievable cost, of the
pollutants in question.
(3) Research data Indicating that excessive concentrations
may be capable of disrupting an aquatic ecosystem.
(4) Substances which result in sludge deposits, produce
unsightly conditions in streams, or result in undesir-
able tastes and odors in water supplies.
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SIGNIFICANCE AND RATIONALE FOR SELECTION OF POLLUTION
PARAMETERS
£H, Acidity, and Alkalinity
Acidity and alkalinity are reciprocal terms. Acidity is
produced by substances that yield hydrogen ions upon hydro-
lysis, and alkalinity is produced by substances that yield
hydroxyl ions. The terms "total acidity" and "total alka-
linity" are often used to express the buffering capacity
of a solution. Acidity in natural waters is caused by
carbon dioxide, mineral acids, weakly dissociated acids,
and the salts of strong acids and weak bases. Alkalinity
is caused by strong bases and the salts of strong alkalies
and weak acids.
The term pH is a logarithmic expression of the concentration
of hydrogen ions. At a pH of 7, the hydrogen and hydroxyl
ion concentrations are essentially equal, and the water is
neutral. Lower pH values indicate acidity, while higher
values indicate alkalinity. The relationship between pH
and acidity or alkalinity is not necessarily linear or
direct.
Haters with a pH below 6.0 are corrosive to water works
structures, distribution lines, and household plumbing
fixtures and can thus add such constituents to drinking
water as iron, copper, sine, cadmium, and lead. The
hydrogen ion concentration can affect the "taste" of the
water. At a low pH, water tastes "sour." The bactericidal
effect of chlorine is weakened as the pH increases, and it
is advantageous to keep the pB close to 7. This is very
significant for providing safe drinking water.
Extremes of pH or rapid pB changes can exert stress condi-
tions or kill aquatic life outright. Dead fish, associated
algal blooms, and foul atrenches are aesthetic liabilities
of any waterway. Even moderate changes from "acceptable"
criteria limits of pB are deleterious to some species.
The relative toxicity to aquatic life of many materials
is increased by changes In the water pB. Hetalocyanide
complexes can increase a thousand-fold la toxicity with a
drop of 1.5 pB units. The availability of many nutrient
substances varies with the alkalinity and acidity. Ammonia
is more lethal with a higher pB.
VX-2
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The lacrimal fluid of the human eye has a pH of approxi-
mately 7.0, and a deviation of 0.1 pH unit from the norm
may result in eye irritation for the swimmer. Appreciable
irritation will cause severe pain.
Acid conditions prevalent in the ore mining and dressing
industry may result from the oxidation of sulfides in mine
waters or discharge from acid-leach milling processes.
Alkaline-leach milling processes also contribute waste load-
ing and adversely affect effluent receiving waters.
Total Suspended Solids
Suspended solids include both organic and inorganic materials.
The Inorganic compounds include sand, silt, and clay. The
organic fraction includes such materials as grease, oil,
tar, animal and vegetable fats, various fibers, sawdust,
hair, and various materials from sewers. These solids may
settle out rapidly, and bottom deposits are often a mixture
of both organic and inorganic solids. They adversely affect
fisheries by covering the bottom of the stream or lake with
a blanket of material that destroys the fish-food bottom
fauna or the spawning ground of fish. Deposits containing
organic materials may deplete bottom oxygen supplies and
produce hydrogen sulfide, carbon dioxide, methane, and other
noxious gases.
In raw water sources for domestic use, state and regional
agencies generally specify that suspended solids in streams
shall not be present in sufficient concentration to be
objectionable or to interfere with normal treatment processes.
Suspended solids in water may interfere with many industrial
processes and cause foaming in boilers or encrustation on
equipment exposed to water, especially as the temperature
rises. Suspended solids are undesirable in water for textile
industries; paper and pulp; beverages; dairy products; laun-
dries; dyeing; photography; cooling systems; and power plants.
Suspended particles also serve as a transport mechanism for
pesticides onto clay particles.
Solids may be suspended in water for a time and then settle
to the bed of the stream or lake. These settleable solids
discharged with man's wastes may be inert, slowly biodegrad-
able materials, or rapidly decomposable substances. While
VI-3
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In suspension, they Increase the turbidity of the water,
reduce light penetration, and impair the photosynthetic
activity of aquatic plants.
Solids in suspension are aesthetically displeasing. When
they settle to form sludge deposits on the stream or lake
bed, they are often much more damaging to the life in water,
and they retain the capacity to displease the senses. Solids,
when transformed to sludge deposits, may do a variety of
damaging things, including blanketing the stream or lake bed
and thereby destroying the living spaces for those benthlc
organisms that would otherwise occupy the habitat. When
of an organic (and, therefore, decomposable) nature, solids
use a portion or all of the dissolved oxygen available in the
area. Organic materials also serve as a seemingly Inexhaust-
ible food source for sludgeworms and associated organisms.
Turbidity is principally a measure of the light-absorbing
properties of suspended solids. It is frequently used as
a substitute method of quickly estimating the total suspended
solids when the concentration is relatively low.
High suspended-solid concentrations are contributed as part
of the mining process, as well as the crushing, grinding,
and other processes commonly used in the milling industry
for most milling operations. High suspended-solid concen-
trations are also characteristic of dredge-mining and gravity-
separation operations.
Oil and Grease
Oil and grease exhibit an oxygen demand. Oil emulsions may
adhere to the gills of fish or coat and destroy algae or
other plankton. Deposition of oil in the bottom sediments
can serve to exhibit normal benthic growths, thus Interrupt-
ing the aquatic food chain. Soluble and emulsified material
ingested by fish may taint the flavor of the fish flesh.
Water-soluble components may exert toxic action on fish.
Floating oil may reduce the re-aeration of the water surface
and, in conjunction with emulsified oil, may interfere with
photosynthesis. Water-insoluble components damage the plumage
and coats of water animals and fowls. Oil and grease in
water can result in the formation of objectionable surface
slicks, preventing the full aesthetic enjoyment of the water.
Oil spills can damage the surface of boats and can destroy
the aesthetic characteristics of beaches and shorelines.
VI-4
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Levels of oil and grease which are toxic to aquatic organisms
vary greatly, depending on the type and the species suscepti-
bility. However, it has been reported that crude oil in
concentrations as low as 0.3 mg/1 is extremely toxic to
fresh-water fish. There is evidence that oils may persist
and have subtle chronic effects.
This parameter is found in discharges of the ore mining and
dressing industry as a result of the contribution from lubri-
cants and spillage of fuels, as well as the usage of reagents
in many milling processes.
Chemical Oxygen Demand (COD) and Total Organic Carbon (TOC)
The chemical oxygen demand (COD) determination provides a
measure of the oxygen equivalent of that portion of the
organic matter in a sample that is susceptible to oxidation
by a strong chemical oxidant. With certain wastes contain-
ing toxic substances, this test—or a total organic carbon
determination—may be the only method for obtaining the
organic load.
Chemical oxygen demand will result in depletion of dissolved
oxygen in receiving waters. Dissolved oxygen (DO) is a
waterwquality constituent that, in appropriate concentrations,
is essential, not only to keep organisms living, but also
to sustain species reproduction, vigor, and the development
of populations. Organisms undergo stress at reduced DO con-
centrations that makes them less competitive and able to
sustain their species within the aquatic environment. For
example, reduced DO concentrations have been shown to inter-
fere with fish populations through delayed hatching of eggs,
reduced size and vigor of embryos, production of deformities
in young, interference with food digestion, acceleration of
blood clotting, decreased tolerance to certain toxicants,
reduced food efficiency and growth rate, and reduced maximum
sustained swimming speed. Fish food organisms are likewise
affected adversely in conditions with suppressed DO. Since
all aerobic aquatic organisms need a certain amount of oxygen,
the total lack of dissolved oxygen due to a high COD can kill
all inhabitants of the affected area.
VI-5
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The total organic carbon (TOG) value generally falls below
the true concentration of organic contaminants because other
constituent elements are excluded. When an empirical rela-
tionship can be established between the total organic carbon,
the biochemical oxygen demand, and the chemical oxygen demand,
the TOC provides a rapid, convenient method of estimating the
other parameters that express the degree of organic contamina-
tion. Forms of carbon analyzed by this test, among others,
are: soluble, nonvolatile organic carbon; insoluble, par-
tially volatile carbon(e.g., oils); and insoluble, particulate
carbonaceous materials (e.g., cellulose fibers).
The final usefulness of the two methods is to assess the
oxygen-demanding load of organic material on a receiving
stream. The widespread use of oil-based compounds, organic
acids, or other organic coumpounds in the flotation process,
as well as the absence of accurate, reproducible tests which
can be routinely performed, points to the use of these tests
as indicators of the levels of particular reagent groups
which are being discharged.
COD reflects the presence of a variety of materials which
may be present in the effluent from ore dressing operations.
Many flotation reagents exert a chemical oxygen demand, and
the presence of excessive levels of these materials in the
effluent stream will be reflected in elevated COD values.
Higher COD values are generally observed for flotation
effluent streams than for those where flotation is not
practiced. In addition, elevated COD values reflect the
release of significant quantities of chemicals whose environ-
mental fates and effects are largely unknown.
Cyanide
Cyanides in water derive their toxiclty primarily from undis-
sociated hydrogen cyanide (HCN), rather than from the cyanide
Ion (CN-) . HCN dissociates in water into H+ and CN- In a
pH-dependent reaction. At a pH of 7 or below, less than 1
percent of the cyanide is present as CN-; at a pH of 8, 6.7
percent; at a pH of 9, 42 percent; and at a pH of 10, 87 percent
of the cyanide Is dissociated. The toxicity of cyanides is
also increased by increases in temperature and reductions in
oxygen tensions. A temperature rise of 10 degrees Celsius (14
degrees Fahrenheit) produces a two- to three-fold increase in
the rate of the lethal action of cyanide.
Cyanide has been shown to be poisonous to humans, and
amounts over 18 ppm can have adverse effects. A single
dose of about 50 to 60 mg is reported to be fatal.
VI-6
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Trout and other aquatic organisms are extremely sensitive
to cyanide. Amounts as small as 0.1 part per million can
kill them. Certain metals, such as nickel, may complex with
cyanide to reduce lethality—especially, at higher pH values—
but zinc and cadmium cyanide complexes are exceedingly toxic.
When fish are poisoned by cyanide, the gills become consid-
erably brighter in color than those of normal fish, owing
to the inhibition by cyanide of the oxidase responsible
for oxygen transfer from the blood to the tissues.
The presence of cyanide in the effluents of the mining and
milling industry is primarily due to the use of cyanide as a
depressant in flotation processes and as a leaching reagent—
particularly, in the gold and silver ore milling categories.
Ammonia
Ammonia is a common product of the decomposition of organic
matter. Dead and decaying animals and plants, along with
human and animal body wastes, account for much of the ammonia
entering the aquatic ecosystem. Ammonia exists in its non-
ionized form only at higher pH levels and is the most toxic
in this state. The lower the pH, the more ionized ammonia
is formed, and its toxicity decreases. Ammonia, in the pres-
ence of dissolved oxygen, is converted to nitrate (N03^ by
nitrifying bacteria. Nitrite (N02), which is an intermediate
product between ammonia and nitrate, sometimes occurs in
quantity when depressed oxygen conditions permit. Ammonia
can exist in several other chemical combinations, including
ammonium chloride and other salts.
Nitrates are considered to be among the poisonous ingredients
of mineralized waters, with potassium nitrate being more
poisonous than sodium nitrate. Excess nitrates cause irri-
tation of the mucous linings of the gastrointestinal tract
and the bladder; the symptoms are diarrhea and diuresis, and
drinking one liter (1.06 quart) of water containing 500 mg/1
of nitrate can cause such symptoms.
Infant methemoglobinemia, a disease characterized by certain
specific blood changes, and cyanosis may be caused by high
nitrate concentrations in the water used for preparing feed-
ing formulae. While it is still impossible to state precise
VI-7
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concentration limits, it has been widely recommended that
water containing more than 10 mg/1 of nitrate nitrogen
(N03_-N) not be used for infants. Nitrates are also
harmful in fermentation processes and can cause disagreeable
tastes in beer. In most natural water, the pH range is such
that ammonium ions (NH4+) predominate. In alkaline waters,
however, high concentrations of un-ionized ammonia in undlsso-
ciated ammonium hydroxide increase the toxicity of ammonia
solutions. In streams polluted with sewage, up to one half
of the nitrogen in the sewage may be in the form of free
ammonia, and sewage may carry up to 35 mg/1 of total nitrogen.
It has been shown that, at a level of 1.0 mg/1 of un-ionized
ammonia, the ability of hemoglobin to combine with oxygen is
impaired, and fish may suffocate. Evidence indicates that
ammonia exerts a considerable toxic effect on all aquatic
life within a range of less than 1.0 mg/1 to 25 mg/1, depending
on the pH and the dissolved oxygen level present. Ammonia
can add to the problem of eutrophication by supplying nitrogen
through its breakdown products. Some lakes in warmer climates,
and others that are aging quickly, are sometimes limited by
the nitrogen available. Any Increase will speed up the plant
growth and the decay process. In leaching operations, ammonia
may be used in leaching solutions (as in the 'Dean-Leute'
ammonium carbamate process, for precipitation of metal salts,
or for pH control. In the ore mining and dressing industry,
high levels at selected locations may thus be encountered.
Aluminum
Aluminum is one of the most abundant elements on the face of
the earth. It occurs in many rocks and ores, but never as
a pure metal. Although some aluminum salts are soluble,
aluminum is not likely to occur for long in surface waters
because it precipitates and settles or is absorbed as alum-
inum hydroxide, carbonate, etc. The mean concentration of
soluble aluminum is approximately 74 micrograms per liter,
with values ranging from 1 to 2,760 micrograms per liter.
Aluminum can be found in all soils, plants, and animal tissues.
The human body contains about 50 to 150 mg of aluminum, and
aluminum concentrations in fruits and vegetables range up
to 37 mg/kg. The total aluminum in the human diet has been
estimated at 10 to 100 mg/day; however, very little of the
aluminum is absorbed by the alimentary canal. Aluminum is
not considered a problem in public water supplies. Note, how-
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DRAFT
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ever, that excessively high doses of aluminum may interfere
with phosphorus metabolism. Aluminum present in surface waters
can be harmful to aquatic life—particularly, marine aquatic
life. Marine organisms tend to concentrate aluminum by a
factor of approximately 10,000. Administration of 0.10 mg/1
of aluminum nitrate for 1 week proved lethal to sticklebacks.
Approximately 5 mg/1 of aluminum is lethal to trout when
exposed for 5 minutes, but the presence of only 1 mg/1 over
the same time period produces no harmful effects.
Aluminum is generally a minor constituent of irrigation waters.
In addition, most soils are naturally alkaline and, as such,
are not subject to the toxic effects of relatively high con-
centrations of aluminum. Where soils are quite acidic (pH
below 5.0), aluminum toxicity to plants becomes very signifi-
cant. Aluminum presence is primarily observed in wastewaters
from the bauxite-ore mining industry. At pH 4.5, 1.0 mg/1
of aluminum reduces the yield of corn 25 percent; at concentrations
of 2.28 and 4.56 mg/1 of aluminum, yields are decreased by
39 percent and 59 percent, respectively.
Antimony
Antimony is rarely found pure in nature, its common forms
being the sulfide, stibnite (Sb2S3) and the oxides cervantite
(Sb2_04) and valentinite (Sb203). Any antimony discharged to
natural waters has a strong tendency to precipitate and be
removed by sedimentation and/or adsorption.
Antimony compounds are toxic to man and are classified as acutely
moderate or chronically severe. A dose of 97.2 mg of antimony
has reportedly been lethal to an adult. Antimony potassium
tartrate, once in use medically to treat certain parasitic
diseases, is no longer recommended because of the frequency
and severity of toxic reactions, including cardiac disturbances.
Various marine organisms reportedly concentrate antimony to
more than 300 times the amount present in the surrounding
waters. Few of the salts of antimony have been tested in
bioassays; as a result, data on antimony toxicity to aquatic
organisms are sketchy. Antimony is commonly found associated
with sulfide ores exploited in the silver and lead industry,
as well as in operations operated for antimony primary or by-
product recovery.
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Arsenic
Arsenic is found to a small extent in nature in the elemental
form. It occurs mostly in the form of arsenites of metals
or as arsenopyrlte (FeS2_.FeAs2_).
Arsenic is normally present in sea water at concentrations
of 2 to 3 micrograms per liter and tends to be accumulated
by oysters and other shellfish. Concentrations of 100 mg/kg
have been reported in certain shellfish. Arsenic is a cumu-
lative poison with long-term chronic effects on both aquatic
organisms and mammalian species, and a succession of small
doses may add up to a final lethal dose. It is moderately
toxic to plants and highly toxic to animals—especially, as
arsine (AsH3) .
Arsenic trioxide, which also is exceedingly toxic, was studied
in concentrations of 1.96 to 40 mg/1 and found to be harmful
in that range to fish and other aquatic life. Work by the
Washington Department of Fisheries on pink salmon has shown
that a level of 5.3 mg/1 of As2_03_ for 8 days is extremely
harmful to this species; on mussels, a level of 16 mg/1 is
lethal in 3 to 16 days.
Severe human poisoning can result from 100-mg concentrations,
and 130 mg has proved fatal. Arsenic can accumulate in the
body faster than it is excreted and can build to toxic levels,
from small amounts taken periodically through lung and intes-
tinal walls from the air, water, and food. Arsenic is a normal
constituent of most soils, with concentrations ranging up to
500 mg/kg. Although very low concentrations of arsenates may
actually stimulate plant growth, the presence of excessive
soluble arsenic in irrigation waters will reduce the yield of
crops, the main effect appearing to be the destruction of
chlorophyll in the foliage. Plants grown in water containing
one mg/1 of arsenic trioxides show a blackening of the vascular
bundles in the leaves. Beans and cucumbers are very sensitive,
while turnips, cereals, and grasses are relatively resistant.
Old orchard soils in Washington that contain 4 to 12 mg/kg
of arsenic trioxide in the topsoll were found to have become
unproductive.
Arsenic is known to be present in many complex metal ores—
particularly, the sulfide ores of cobalt, nickel and other
ferroalloy ores, antimony, lead, and silver. It may also be
solubillzed in mining and milling by oxidation of the ore and
appear in the effluent stream.
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Beryllium
Beryllium is a relatively rare element, found chiefly in the
mineral beryl. In the weathering process, beryllium is con-
centrated in hydrolyzate and, like aluminum, does not go into
solution to any appreciable degree. Beryllium is not likely
to be found in natural waters in greater than trace amounts
because of the relatively insolubility of the oxide and hydrox-
ide at the normal pH range of such waters.
Absorption of beryllium from the alimentary tract is slight,
and excretion is fairly rapid. However, as an air pollutant,
it is responsible for causing skin and lung diseases of variable
severity.
Concentrations of beryllium sulfate complexed with sodium
tartrate up to 28.5 mg/1 are not toxic to goldfish, minnows,
or snails. The 96-hour minimum toxic level of beryllium sulfate
for fathead minnows has been found to be 0.2 mg/1 in soft water
and 11 mg/1 in hard water. The corresponding level for beryllium
chloride is 0.15 mg/1 in soft water and 15 mg/1 in hard water.
In nutrient solution, at acid pH values, beryllium is highly
toxic to plants. Solutions containing 15 to 20 mg/1 of beryllium
delay germination and retard the growth of cress and mustard
seeds in solution culture. The presence of beryllium in
wastewaters was detected only in raw-waste effluents from the
mining and milling of bertrandite.
Cadmium
Cadmium in drinking water supplies is extremely hazardous to
humans, and conventional treatment, as practiced in the United
States, does not remove it. Cadmium is cumulative in the liver,
kidney, pancreas, and thyroid of humans and other animals.
A severe bone and kidney syndrome in Japan has been associated
with the ingestion of as little as 600 micrograms per day of
cadmium.
Cadmium is an extremely dangerous cumulative toxicant, causing
insidious progressive chronic poisoning in mammals, fish,
and (probably) other animals because the metal is not excreted.
Cadmium can form organic compounds which may lead to mutagenic
or teratogenic effects. Cadmium is known to have marked acute
and chronic effects on aquatic organisms also.
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Cadmium acts synergistically with other metals. Copper and
zinc substantially increase its toxicity. Cadmium is concen-
trated by marine organisms—particularly, mollusks, which
accumulate cadmium in calcareous tissues and in the viscera.
A concentration factor of 1000 for cadmium in fish muscle has
been reported, as have concentration factors of 3,000 in marine
plants, and up to 29,600 in certain marine animals. The eggs
and larvae of fish are, apparently, more sensitive than adult
fish to poisoning by cadmium, and crustaceans appear to be
more sensitive than fish eggs and larvae.
Cadmium, in general, is less toxic in hard water than in soft
water. Even so, the safe levels of cadmium for fathead minnows
and bluegills in hard water have been found to be between 0.06
and 0.03 mg/1, and safe levels for coho salmon fry have been
reported to be 0.004 to 0.001 mg/1 in soft water. Concentrations
of 0.0005 mg/1 were observed to reduce reproduction of Daphnia
magna in one-generation exposure lasting three weeks.
Cadmium is present in minor amounts in the effluents from
several ferroalloy-ore and copper mining and milling operations.
Chromium
Chromium, in its various valence states, is hazardous to man.
It can produce lung tumors when inhaled and induces skin sensi-
tizations. Large doses of chromates have corrosive effects
on the intestinal tract and can cause inflammation of the
kidneys. Levels of chromate ions that have no effect on man
appear to be so low as to prohibit determination to date.
The toxicity of chromium salts toward aquatic life varies
widely with the species, temperature, pH, valence of the
chromium, and synergistic or antagonistic effects—especially,
that of hardness. Fish are relatively tolerant of chromium
salts, but fish-food organisms and other lower forms of aquatic
life are extremely sensitive. Chromium also inhibits the
growth of algae.
In some agricultural crops, chromium can cause reduced growth
or death of the crop. Adverse effects of low concentrations
of chromium on corn, tobacco, and sugar beets have been docu-
mented .
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Chromium is present at appreciable concentrations In the
effluent from mills practicing leaching. It is also present
as a minor constituent in many ores, such as those of plat-
inum, ferroalloy metals, lead, and zinc.
Copper
Copper salts occur in natural surface waters only in trace
amounts, up to about 0.05 mg/1, so their presence generally
is the result of pollution. This is attributable to the
corrosive action of the water on copper and brass tubing,
to Industrial effluents, and—frequently—to the use of
copper compounds for the control of undesirable plankton
organisms.
Copper is not considered to be a cumulative systemic poison
for humans, but it can cause symptoms of gastroenteritis, with
nausea and intestinal irritations, at relatively low dosages.
The limiting factor in domestic water supplies is taste.
Threshold concentrations for taste have been generally reported
in the range of 1.0 to 2.0 mg/1 of copper, while as much as
5 to 7.5 mg/1 makes the water completely unpalatable.
The toxicity of copper to aquatic organisms varies significantly,
not only with the species, but also with the physical and
chemical characteristics of the water, including temperature,
hardness, turbidity, and carbon dioxide content. In hard
water, the toxicity of copper salts is reduced by the precipi-
tation of copper carbonate or other insoluble compounds. The
sulfates of copper and zinc, and of copper and cadmium, are
synergistic in their toxic effect on fish.
Copper concentrations less than 1 mg/1 have been reported to
be toxic—particularly, in soft water—to many kinds of fish,
crustaceans, mollusks, insects, phytoplankton, and zooplankton.
Concentrations of copper, for example, are detrimental to some
oysters above 0.1 ppm. Oysters cultured in sea water con-
taining 0.13 to 0.5 ppm of copper deposit the metal in their
bodies and become unfit as a food substance.
Besides, those used by the copper mining and milling industry,
many other ore minerals in the ore mining and dressing industry
contain byproduct or minor amounts of copper; therefore, the
waste streams from these operations contain copper.
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Fluorides
As the most reactive non-metal, fluorine is never found free
in nature, but rather occurs as a constituent of fluorite or
fluorspar (calcium fluoride) in sedimentary rocks and also as
cryolite (sodium aluminum fluoride) in igneous rocks. Owing
to their origin only in certain types of rocks and only in a
few regions, fluorides in high concentrations are not a common
constituent of natural surface waters, but they may occur
in detrimental concentrations in ground waters.
Fluorides are used as insecticides, for disinfecting brewery
apparatus, as a flux in the manufacture of steel, for preserving
wood and mucilages, for the manufacture of glass and enamels,
in chemical industries, for water treatment, and for other
uses.
Fluorides in sufficient quantity are toxic to humans, with
doses of 250 to 450 mg giving severe symptoms or causing death.
There are numerous articles describing the effects of fluoride-
bearing waters on dental enamel of children; these studies
lead to the generalization that water containing less than 0.9
to 1.0 mg/1 of fluoride will seldom cause mottled enamel in
children; for adults, concentrations less than 3 or 4
mg/1 are not likely to cause endemic cumulative fluorosis and
skeletal effects. Abundant literature is also available
describing the advantages of maintaining 0.8 to 1.5 mg/1 of
fluoride ion in drinking water to aid in the reduction of
dental decay—especially, among children.
Chronic fluoride poisoning of livestock has been observed in
areas where water contains 10 to 15 mg/1 fluoride. Concentra-
tions of 30 to 50 mg/1 of fluoride in the total ration of
dairy cows are considered the upper safe limit. Fluoride
from waters, apparently, does not accumulate in soft tissue
to a significant degree, and it is transferred to a very small
extent into milk and, to a somewhat greater degree, into eggs.
Data for fresh water indicate that fluorides are toxic to fish
at concentrations higher than 1.5 mg/1.
High fluoride levels in the effluents from mines may result from
high levels in intercepted aquifers or from water contact from
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rock dust and fragments. The use of mine water in milling,
as well as extended contact of water with crushed and ground
ore, may yield high fluoride levels in mill effluents. Levels
may also be elevated by chemical action in leaching operations.
Iron
Iron is one of the most abundant constituents of rocks and
soils and, as such, is often found in natural waters. Although
many of the ferric and ferrous salts, such as the chlorides,
are highly soluble in water, ferrous ions are readily oxidized
in natural surface waters to insoluble ferric hydroxides.
These precipitates tend to agglomerate, flocculate, and settle
or be absorbed in surfaces; hence, the concentration of iron
in well-aerated waters is seldom high. Mean concentrations
of iron in U.S. waters range from 19 to 173 tnicrograms per liter,
depending on geographic location. When the pH is low, however,
appreciable amounts of iron may remain in solution.
Standards for drinking water are not set for health reasons.
Indeed, some iron is essential for nutrition, and larger quantities
of iron are taken for therapeutic reasons. The drinking-water
standards are set for esthetic reasons.
In general, very little iron remains in solution; but, if the
water is strongly buffered and a large enough dose is supplied,
the addition of a soluble iron salt may lower the pH of the
water to a toxic level. In addition, a fish's respiratory
channel may become irritated and blacked by depositions of
iron hydroxides on the gills. Finally, heavy precipitates of
ferric hydroxide may smother fish eggs.
The threshold concentration for lethality to several types
of fish has been reported as 0.2 mg/1 of iron. Concentrations
of 1 to 2 mg/1 of iron are indicative of acid pollution and
other conditions unfavorable to fish. The upper limit for
fish life has been estimated at 50 mg/1. At concentrations of
iron above 0.2 mg/1, trouble has been experienced with populations
of the iron bacterium Crenothrix.
Iron is very common in natural waters and is derived from
common iron minerals in the substrata. The iron may occur
in two forms: suspended and dissolved. The iron mining
and processing industry inherently increases iron levels present
in process or mine waters. The aluminum-ore mining and dressing
industry also contributes elevated iron levels through mine
drainage.
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Lead
Lead sulfide and lead oxide are the primary forms of lead found
in rocks. Certain lead salts, such as the chloride and the
acetate, are highly soluble; however, since the carbonate and
hydroxide are insoluble and the sulfide is only slightly
soluble, lead is not likely to remain in solution long in
natural waters. In the U.S., lead concentrations in surface
and ground waters used for domestic supplies average 0.01
tog/1. Some natural waters in proximity to mountain limestone
and galena contain as much as 0.4 to 0.8 mg/1 of lead in solu-
tion.
Lead is highly toxic to human beings and is a cumulative poison.
Typical symptoms of advanced lead poisoning are constipation,
loss of appetite, anemia, abdominal pain, and gradual paralysis
in the muscles. Lead poisoning usually results from the cumu-
lative 'toxic effects of lead after continuous ingestion over
a long period of time, rather than from occasional small doses.
The level at which the amount of bodily lead intake exceeds
the amount excreted by the body is approximately 0.3 rag/day.
A total intake of lead appreciably in excess of 0.6 mg/day
may result in the accumulation of a dangerous quantity of
lead during a lifetime.
The toxic concentration of lead for aerobic bacteria is reported
to be 1.0 mg/1; for flagellates and infusoria, 0.5 mg/1.
Inhibition of bacterial decomposition of organic matter occurs
at lead concentrations of 0.1 to 0.5 mg/1. Toxic effects
of lead on fish include the formation of a coagulated mucus
film over the gills—and, eventually, the entire body—which
causes the fish to suffocate. Lead toxicity if very dependent
on water hardness; in general, lead is much less toxic in
hard water. Some data indicate that the median period of
survival of rainbow trout in soft water containing dissolved
lead is 18 to 24 hours at 1.6 mg/1. The 96-hour minimum toxic
level for fathead minnows to lead has been reported as 2.4
mg/1 of lead in soft water and 75 mg/1 in hard water. Toxic
levels for fish can range from 0.1 to 75 mg/1 of lead, depending
on water hardness, dissolved oxygen concentration, and the
type of organism studied. Sticklebacks and minnows have not
been visibly harmed when in contact with 0.7 mg/1 of lead in
soft tap water for 3 weeks. However, the 48-hour minimum toxic
level for sticklebacks in water containing 1,000 to 3,000
mg/1 of dissolved solids is reported to be 0.34 mg/1 of lead.
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The U.S. Public Health Service Drinking Water Standard specifies
a rejection limit of 0.05 ppm (mg/1) for lead.
Elevated concentrations of lead are discharged from lead and
zinc mines and mills, as well as from mining and milling
operations exploiting other sulfide ores, such as tetrahedrite
(for silver and lead); copper ores; ferroalloy ore minerals;
or mixed copper, lead, and zinc ores.
Manganese
Pure manganese metal is not naturally found in the earth, but
its ores are very common. Similar to iron in its chemical
behavior, it occurs in the bivalent and trivalent forms. The
nitrates, sulfates, and chlorides are very soluble in water,
but the oxides, carbonates, and hydroxides are only sparingly
soluble. The background concentration of manganese in most
natural waters is less than 20 micrograms per liter.
Manganese is essential for the nutrition of both plants and
animals. The toxicological significance of manganese to
mammals is considered to be of little, although some cases
of manganese poisoning have been reported due to unusually
high concentrations. Manganese limits for drinking water
have been set for esthetic reasons rather than physiological
hazards.
As with most elements, toxicity to aquatic life is dependent
on a variety of factors. The lethal concentration of man-
ganese for the stickleback has been given at 40 mg/1. The
threshold toxic concentration of manganese for the flatworm
Polycelis nigra has been reported to be 700 mg/1 when in the
form of manganese chloride and 660 mg/1 when in the form of
manganese nitrate. Trench, carp, and trout tolerate a manganese
concentration of 15 mg/1 for 7 days; yet, concentrations of
manganese above 0.005 mg/1 have a toxic effect on some algae.
Manganese In nutrient solutions has been reported to be toxic
to many plants, the response being a function of species and
nutrient-solution composition. Toxic levels of manganese in
solution can vary from 0.5 to 500 mg/1.
On the basis of the literature surveyed, it appears that the
concentrations of manganese listed below are deleterious to
the stated beneficial uses.
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a. Domestic water supply 0.05 mg/1
b. Industrial water supply 0.05 mg/1
c. Irrigation 0.50 mg/1
d. Stock watering 10.0 mg/1
e. Fish and aquatic life 1.0 mg/1
Elevated manganese concentrations are found in the effluents
of iron-ore, lead, and zinc mining and milling operationsand
would be expected from any future operations exploiting mangan-
ese ores.
Mercury
Elemental mercury occurs as a free metal in certain parts of
the world; however, since it is rather inert and insoluble
in water, it is not likely to be found in natural waters.
Although elemental mercury is insoluble in water, many of the
mercuric and mercurous salts, as well as certain organie
mercury compounds, are highly soluble in water. Concentrations
of mercury in surface waters have usually been found to be
much less than 5 micrograms per liter.
The accumulation and retention of mercurial compounds in the
nervous system, their effect on developing tissue, and the
ease of their transmittal across the placenta make them parti-
cularly dangerous to man. Continuous intake of methyl mercury
at dosages approaching 0.3 mg Hg per 70 kg (154 Ib) of body
weight per day will, in time, produce toxic symptoms.
Mercury's cumulative nature also makes it extremely dangerous
to aquatic organisms, since they have the ability to absorb
significant quantities of mercury directly from the water as
well as through the food chain. Methyl mercury is the major
toxic form; however, the ability of certain microbes to synthe-
size methyl mercury from the inorganic forms renders all mercury
in waterways potentially dangerous. Fresh-water phytoplankton,
macrophytes, and fish are capable of biologically magnifying
mercury concentrations from water 1,000 times. A concentration
factor of 5,000 from water to pike has been reported, and factors
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of 10,000 or more have been reported from water to brook trout.
The chronic effects of mercury on aquatic organisms are not
well-known. The lowest reported levels which have resulted
in the death of fish are 0.2 micrograms per liter of mercury,
which killed fathead minnows exposed for six weeks. Levels
of 0.1 microgram per liter decrease photosynthesis and growth
of marine algae and some freshwater phytoplankton.
Mercury has been observed in significant quantities in the
wastewater in operations associated with sulfide mineralization,
including mercury ores, lead and zinc ores, and copper ores,
as well as precious-metal operations of gold and silver.
It may be liberated in mine waters as well as in effluents
of flotation concentration and acid-leaching extraction.
Molybdenum
Molybdenum and its salts are not normally considered serious
pollutants, but the metal is biologically active. Although
the element occurs in some minerals, it is not widely distri-
buted in nature. The mean level of molybdenum in the U.S.
has been reported to be 68 micrograms per liter.
The 96-hour minimum toxic level of fathead minnows for molybdic
anhydride (MoOjl) was found to be 70 mg/1 in soft water and
370 mg/1 in hard water. The threshold concentration for dele-
terious effects upon the alga Scenedesmus occurs at 54 mg/1.
I-i coli and Daphnia tolerate concentrations of 1000 mg/1 without
perceptible injury. Molybdenum can be concentrated from 8 to
60 times by a variety of marine organisms, including benthic
algae, zooplankton, mollusks, crustaceans, and teleosts.
Concentrations of a maximum of 0.05 of the 96-hour minimum
toxic level are recommended for protection of the most sen-
sitive species in sea water, while the 24-hour average should
not exceed 0.02 of the 96-hour minimum toxic level.
Molybdenum is found in significant quantities in molybdenum
mining and in milling of uranium ores, where molybdenum is
sometimes recovered as a byproduct.
Nickel
Elemental nickel seldom occurs in nature, but nickel compounds
are found in many ores and minerals. As a pure metal, it is
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not a problem in water pollution because it is not affected
by, or soluble in, water. Many nickel salts, however, are
highly soluble in water.
Nickel is extremely toxic to citrus plants. It is found in
many soils in California, generally in insoluble form, but
excessive acidification of such soil may render it soluble,
causing severe injury to or the death of plants. Many experi-
ments with plants in solution cultures have shown that nickel
at 0.5 to 1.0 mg/1 is inhibitory to growth.
Nickel salts can kill fish at very low concentrations. Data
for the fathead minnow show death occurring in the range of 5
to 43 mg, depending on the alkalinity of the water.
Nickel is present in coastal and open ocean concentrations in
the range of 0.1 to 6.0 micrograms per liter, although the
most common values are 2 to 33 micrograms per liter. Marine
animals contain up to 400 micrograms per liter, and marine
plants contain up to 3,000 micrograms per liter. The lethal
limit of nickel to some marine fish has been reported to be as
low as 0.8 ppm (mg/1) (800 micrograms per liter). Concen-
trations of 13.1 mg/1 have been reported to cause a 50-percent
reduction of photosynthetic activity in the giant kelp
(Macrocystis pyrifers) in 96 hours, and a low concentration
has been found to kill oyster eggs.
Nickel is found in significant quantities as a constituent of
raw wastewater in the titanium, rare-earth, mercury, uranium,
and, occasionally, in the iron-ore mining and milling industries.
Vanadium
Metallic vanadium does not occur free in nature, but minerals
containing vanadium are widespread. Vanadium is found in
many soils and occurs in vegetation grown in such soils.
Vanadium adversely affects some plants in concentrations as
low as 10 mg/1. Vanadium as calcium vanadate can inhibit the
growth of chicks and, in combination with selenium, increases
mortality in rats. Vanadium appears to inhibit the synthesis
of cholesterol and to accelerate its catabolism in rabbits.
Vanadium causes death to occur in fish at low concentrations.
The amount needed for lethality depends on the alkalinity of
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the water and the specific vanadium compound present. The
common bluegill can be killed by about 6 mg/1 in soft water
and 55 mg/1 in hard water when the vanadium is expressed as
vanadyl sulfate. Other fish are similarly affected.
Limitation and control of vanadium levels appear to be necessary
in the effluents from operations employing leaching methods to
extract vanadium as a primary product or byproduct. As treated
here, it can be expected to be contributed by the ferroalloy
industry, where high vanadium levels were observed both in
barren solutions from a solvent extraction circuit and in
scrubber waters from ore roasting units. High vanadium values
are also found associated with uranium operations, where
vanadium is also obtained as a byproduct.
Zinc
Occurring abundantly in rocks and ores, zinc is readily refined
into a stable pure metal and is used extensively for galvanizing,
in alloys, for electrical purposes, in printing plates, for
dye manufacture and for dyeing processes, and for many other
industrial purposes. Zinc salts are used in paint pigments,
cosmetics, pharamaceuticals, dyes, insecticides, and other
products too numerous to list herein. Many of these salts
(e.g., zinc chloride and zinc sulfate) are highly soluble in
water; hence,- it is to be expected that zinc might occur in
many industrial wastes. On the other hand, some zinc salts
(zinc carbonate, zinc oxide, and zinc sulfide) are insoluble
in water; consequently, it is to be expected that some zinc
will precipitate in and be removed readily from most natural
waters.
In zinc-mining areas, zinc has been found in waters in concen-
trations as high as 50 mg/1; in effluents from metal-plating
works and small-arms ammunition plants, it may occur in
significant concentrations. In most surface and ground waters,
it is present only in trace amounts. There is some evidence
that zinc ions are adsorbed strongly and permanently on silt,
resulting in inactivation of the zinc.
Concentrations of zinc in excess of 5 mg/1 in raw water used
for drinking water supplies cause an undesirable taste which
persists through conventional treatment. Zinc can have an
adverse effect on man and animals at high concentrations.
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In soft water, concentrations of zinc ranging from 0.1 to
0.1 mg/1 have been reported to be lethal to fish. Zinc is
thought to exert its toxic action by forming insoluble com-
pounds with the mucous that covers the gills, by damage to
the gill epithelium, or possibly by acting as an internal
poison. The sensitivity of fish to zinc varies with species,
age, and conditions, as well as with the physical and chemi-
cal characteristics of the water. Some acclimatization to
the presence of zinc is possible. It has also been observed
that the effects of zinc poisoning may not become apparent
immediately, so fish relocated from zinc-contaminated water
to zinc-free water, after 4 to 6 hours of exposure to zinc, may
die 48 hours later. The presence of copper in water may
increase the toxicity of zinc to aquatic organisms, but
the presence of calcium (hardness) may decrease the relative
toxicity.
Observed values for the distribution of zinc in ocean waters
very widely. The major concern with zinc compounds in marine
water is not one of acute toxicity, but rather of the long-
term sublethal effects of the metallic compounds and complexes.
From an acute-toxicity point of view, invertebrate marine
animals seem to be the most sensitive organisms tested.
The growth of the sea urchin, for example, has been retarded
by as little as 30 micrograms per liter of zinc.
Zinc sulfate has also been found to be lethal to many plants,
and it could impair agricultural uses.
Elevated zinc levels were found at operations for the mining
and milling of lead and zinc ores; at copper mines and flo-
tation mills; at gold, silver, titanium, and beryllium opera-
tions; and at most ferroalloy-ore mining and milling sites.
Radioactivity
Ionizing radiation, when absorbed in living tissue in quan-
tities substantially above that of natural background levels,
is recognized as injurious. It is necessary, therefore, to
prevent excessive levels of radiation from reaching any liv-
ing organism: humans, fishes, or invertebrates. Beyond the
obvious fact that radioactive wastes emit ionizing radiation,
such wastes are also similar in many respects to other chemi-
cal wastes. Man's senses cannot detect radiation unless it
is present in massive amounts.
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Plants and animals, to be of any significance in the cycling
of radionuclides in the aquatic environment, must accumulate
the radionuclide, retain it, be eaten by another organism,
and be digestible. However, even if an organism accumulates
and retains a radionuclide and is not eaten before it dies,
the radionuclide will enter the "biological cycle" through
organisms that decompose the dead organic material into its
elemental components. Plants and animals that become radio-
active in this biological cycle can thus pose a health hazard
when eaten by man.
Aquatic life may receive radiation from radionuclides present
in the water and substrata and also from radionuclides that
may accumulate within their tissues. Humans can acquire
radionuclides through many different pathways. Among the
most important are drinking contaminated water and eating
fish and shellfish that have concentrated nuclides from the
water. Where fish or other fresh or marine products that
have accumulated radioactive materials are used as food by
humans, the concentrations of the nuclides in the water must
be further restricted, to provide assurance that the total
intake of radionuclides from all sources will not exceed
the recommended levels.
To prevent unacceptable doses of radiation from reaching
humans, fish, and other Important organisms, the concentra-
tions of radionuclides in water, both fresh and marine, must
be restricted.
Radium
Radium is a natural product of the disintegration of uranium.
It undergoes spontaneous disintegration with the formation of
radon (Rn 222), one gram of radium producing about 10 exp (-14)
ml of radon per day. The mean radium concentration in
raw waters has been measured at 0.049 x 10 exp(-9) mg/1,
based on 42 sources serving approximately one-fifth of the
population of the U.S.
Radium is the most hazardous radioelement of the 14 radio-
active isotopes that occur in nature with uranium as their
parent. The bulk of the radium discarded as waste is gener-
ally retained undlssolved in tailing ponds; however, the
remaining dissolved portion can constitute a significant
stream-pollution problem, affecting domestic and industrial
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water supplies, irrigation, stock watering, and aquatic life.
In addition, dissolved radium can also escape to streams and
build up in bottom muds. When first introduced into natural
waters, a large part of the materials present in radioactive
wastes, including radium, becomes associated with suspended
organic and inorganic particulates that settle to the bottom,
and many radloisotopes are eventually bound chemically to the
sediments.
Radium, like the other radioisotopes, is passed through the
various trophic levels of the food chain and are either bio-
concentrated or released. Possible effects to an individual
organism exposed to radium may include death, inhibition
or stimulation of growth, physiological change, behavioral
changes, developmental abnormalities, or shortening of life
span.
Radium is present in mine and process waters of the uranium
industry because of the radioactive decay of uranium isotopes.
Thorium
Thorium is a grayish-white, lustrous metal which occurs
naturally in several mineral forms, such as the silicate
(thorite), the oxide (thortanite) , and the phosphate (mona-
zite) . The toxicity of thorium, based on radiation effects,
is estimated to be three times that of uranium.
Besides radioactive toxicity, thorium minerals also exhibit
chemical toxicity. The median threshold effect of thorium
nitrate on Scenedesmus . E^. coll. and a protozoan (Mlcroregna)
was found to be 0.4 to 0.8 mg/1 of thorium. The lethal con-
centration of thorium chloride for three mature, small fresh-
water fish exposed for 24 hours was reported to be about 18
Thorium is present in mine and process waters of the uranium
mining and processing industry.
Uranium
Uranium is present in wastes from uranium mines and mills
and nuclear fuel processing plants, and the uranyl ion may
VI-24
DRAFT
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naturally occur in drainage waters from uranium-bearing ore
deposits.
Many of the salts of uranium are present in sea water,
which has an average concentration of the metal of about 3
micrograms per liter. Uranium is stabilized by hydrolysis,
which tends to prevent its physical and chemical interaction
and, thus, prevents its removal from sea water. The uranium
salts are considered to be four times as germicidal as phenol
to aquatic organisms.
Natural uranium (U 238) is concentrated by the algae Ochro-
monas by a factor of 330 in 48 hours. The threshold of
uranyl nitrate, expressed as uranium, was found to be 28 mg/1
for a protozoan (Microregma), 1.7 to 2.2 mg/1 for Escherichia
coli, 22 mg/1 for the alga Scenedesmus, and 13 mg/1 for Daphnia.
The nitrate, sulfate, and acetate salts of uranium have been
found to be more toxic to fathead minnows, Pimephales promelas,
in soft water than in hard water, with the 96-hour minimum
toxic level for uranyl sulfate being 2.8 mg/1 in soft water
and 135 mg/1 in hard water.
A factor of 0.01 has been recommended for application to
marine 96-hour minimum toxic-level data for sensitive organisms.
The radiation hazard from uranium is considered to be much
less than that of radium; in addition, uranium has a much
shorter "biological" half-life and affects less-sensitive
tissues (gastrointestinal tract). The chemical toxicity is
believed to exceed the radiation hazard from ingestion.
Uranium concentrations in raw mine/mill wastes are often
below 40 mg/1, and discharge levels below 1 mg/1 are routinely
achieved.
Flotation Reagents
The toxicity of organic floation agents—particularly, collec-
tors and their decomposition products—is an area of consider-
able uncertainty, particularly in the complex chemical environ-
ment present in a typical flotation-mill discharge. Standard
analytical tests for individual organic reagents have not
evolved to date. The tests for COD and TOC are the most
reliable tests currently available which give indications
of the presence of some of the flotation reagents.
VI-25
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Data available on the fates and potential toxicities of many
of the reagents indicate that only a broad range of tolerance
values is known. Table VI-1 is a list of some of the more
common flotation reagents and their known toxicities as judged
from organism tolerance information.
Asbestos
"Asbestos" is a generic term for a number of fire-resistant
hydrated silicates that, when crushed or processed, separate
into flexible fibers made up of fibrils noted for their great
tensile strength. The asbestos minerals differ in their
metallic elemental content, range of fiber diameters, flexi-
bility, hardness, tensile strength, surface properties, and
other attributes which may affect their respirability, deposi-
tion, retention, translocation, and biologic reactivity.
Asbestos is toxic by inhalation of dust particles, with the
tolerance being 5 million particles per cubic foot of air.
Prolonged inhalation can cause cancer of the lungs, pleura,
and peritoneum. Little is known about the movement of asbes-
tos fibers within the human body, including their potential
entry through the gastrointestinal tract. There is evidence
that bundles of fibrils may be broken down within the body
to individual fibrils. Asbestos has the possibility of being
a hazard when waterborne in large concentrations; however, it
is insoluble in water.
To date, there is little data on the concentrations of asbestos
in ore mining and milling water discharges. Knowledge of the
concentrations in water that pose health problems is poorly
defined. Currently, this area is being investigated by many
researchers concerning themselves with health, movement, and
analytical techniques.
Because of public reports concerning the presence of asbestos
in wastewater from an iron-ore beneficiation operation, a
reconaissance analysis for asbestos was performed on samples
collected as part of site visits to four discharging iron-ore
beneficiation operations. The raw wastewater and effluent
of tailing ponds at each facility were examined for the pre-
sence or absence of asbestos or asbestos-like fibers. The
method of analysis used for detection was one based upon pub-
lished literature and employed scanning electron microscopy.
VI-26
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TABLE Vl-l. KNOWN TOXIC ITY OF SOME COMMON FLOTATION REAGENTS
USED IN ORE MINING AND MILLING INDUSTRY
TRADE NAME
Aerofloat 26
Aerofloat 31
Aarof loot 238
Aeroftaat242
AerofrothW
AerofrothTI
AflVO ^TOHIOlVf
404
3477
AROSURF
MO-MA
Dowf roth 260
DowZ-6
OowZ-11
DowZ-200
Jaguar
M.I.B.C.
-
Suparfloc 16
CHEMICAL COMPOSITION
FIT Mitt laHv onrf ffltlilnirfimntwirir arlri
Etnntlally «ryl dithiophoaphorlc acid
Sodium dl-taeondary butyl
dithlophosphate
essentially aryl dithlophosphorlc MM
PMydJyeol type compound
Mixture of tulf hydryl type compounds
Unknown
Unknown
Chromium tills (ammonium, potassium.
•nd todlum chromota and ammonium,
potassium, and codlum dkhromata)
Copper ulfata
Cresylic acid
Polypropylene dlycol methyl •than
Potassium amyl xanthata
Sodium bopropyl xanthata
liopropyl ethylthionocarfaamate
Basad on guar gum
Uma (calcium oxide)
Mathvllwbutylcarblnol
Pinaoll
Potaolum farrleyanida
Sodium farrocyanlda
Sodium hydroxide
Sodium olaata
Sodium illicata
Sodium aulflda
Sulfuric acid
Polyaerylamida
FUNCTION
Collertor/Promoter
Collector/Promoter
Collector/Promoter
Frothor
Frothar
Collect or/Promoter
Collect or/Promotar
11 m
Oepreolng agent
Activetlng agent
Frothar
Frother
Collector/Promoter
Collector/Promoter
Collector/Promoter
Floeculant
pH modlfiar and
floeculant
Frother
Frother
Dapiming agent
DepraBlng agent
pH (nooificr
Frother
Deprenlng agent
Activating agent
pH modlfiar and
floeculant
Flocculant
KNOWN TOXIC
RANGE imafi)
1000 to 10^00
10 to 1000
1000 to \OJOOO
>1000
1 to 100
100 to 1000
10 to 1000
0.01 to 1.0
0.1 to 1.0
>1000
0.1 to 200
Oi to 2.0
10 to 100
10 to 1000
>1000
1 to 100
0.25 to 2.6
1 to 1600
1 to 1000
1 to 1000
100 to 1000
1 to 100
1 to 100
>1000
TOXICITY"
Low
Moderate
Low
Low
Moderate
Moderate
Moderate
High
High
Low
Moderate to High
High
Modorata
Moderate
Low
Moderate
Moderate to High
Moderate
Moderate
Moderate
Moderate
Moderate
Moderate
Low
Texterty Tolerance Level
High <1.0 mg/l
Moderate I.Oto lOOOmg/Jt
Low
>1000mg/£
NOTE: Toxic range » a function of orginlim tailed and water quality. Including hardneai
and pH. Therefore, toxiclty data presented in this table aro only generally indica-
tive of reagent toxiclty. Although the toxiclty ration presented here are based on
many different organism*, much of the data are presented in relation to almon,
fathead minnows, sticklebacks, and Daphnia.
VI-27
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Fibers were not detected in any of the samples with the excep-
tion of the influent to the tailing pond from Mill 1107.
Energy-dispersive x-ray analysis indicated, however, that
the fiber was not of an asbestos type. Both raw and treated
wastewaters from mills 1107, 1108, 1109, and 1110 were examined,
and no asbestos or asbestos-like minerals were found.
While the results of the survey indicate the absence of asbes-
tos fibers at each of the sites investigated, the presence
or absence of asbestos at other locations in the iron-ore
mining and beneficiation industry cannot be confirmed. It
does not appear possible to recommend effluent levels or
treatment technology at this time. It is recommended, how-
ever, that a reconalssance evaluation for asbestos be performed
at each iron-ore mining and beneficiation operation to deter-
mine whether possible asbestos levels of concern are present.
VI-28
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SIGNIFICANCE AND RATIONALE FOR REJECTION OF POLLUTION
PARAMETERS
A number of pollution, parameters besides those selected and
just discussed were considered In each category but were
rejected for one or more of these reasons:
(1) Simultaneous reduction Is achieved with another
parameter which is limited.
(2) Treatment does not "practically" or economically
reduce the parameter.
(3) The parameter was not usually observed In quantities
sufficient to cause water-quality degradation.
(4) There are insufficient data on water-quality degra-
dation or treatment methods which might be employed.
Because of the great diversity of the ores mined and the
processes employed in the ore mining and dressing Industry,
selections for subcategories of the parameters to be monitored
and controlled—as well as those rejected—vary considerably.
Parameters listed in this section are parameters which have
been rejected for the ore mining and dressing industry as a
whole.
Barium and Boron
Barium and boron are not present in quantities sufficient
to justify consideration as harmful pollutants.
Calcium, Magnesium, Potassium, Strontium, and Sodium
Although these metals commonly occur in effluents associated
with ore mining and dressing activities, they are not present
in quantities sufficient to cause water-quality degradation,
or there are no practical treatment methods which can be
employed on a large scale to control these elements.
Carbonate
There are insufficient data for dissolved carbonate to justify
consideration of this ion as a harmful pollutant.
VI-29
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Nitrate and Nitrite
There are Insufficient data for dissolved nitrates and nitrites
to justify their consideration as harmful pollutants, although
nitrogen and nitrate contributions are known to stimulate
plant and algal growth. There is no treatment available to
practically reduce these ions.
Selenium
The levels of selenium observed in the wastewaters from mines
and mills are not sufficiently high for selenium to be con-
sidered as a harmful pollutant.
Silicates
Silicates may be present in the wastewaters from the ore min-
ing and dressing industry, but the levels encountered are not
sufficiently high to warrant classification as a harmful
pollutant.
Tin
Tin does not exist in sufficient quantities from mines or
mills to be considered a harmful pollutant.
Zirconium
There is no information available which indicates that signi-
ficant levels of zirconium are present in the industry to be
classed as harmful.
Total Dissolved Solids
High dissolved-solid concentrations are often caused by acid
conditions or by the presence of easily dissolved minerals
in the ore. Since economic methods of dissolved-solid reduction
do not exist, effluent limitations have not been proposed
for this parameter.
VI-30
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SUMMARY OF POLLUTION PARAMETERS SELECTED BY CATEGORY
Because of the wide variations observed with respect to both
waste components discharged and loading factors in the differ-
ent segments of the ore mining and dressing industry, a single,
unified list of all parameters selected for the industry as
a whole would not be useful. Therefore, Table VI-2 summarizes
the parameters chosen for effluent limitation guidelines for
each industry metal category.
VI-31
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TABLE VI-2. SUMMARY OF PARAMETERS SELECTED FOR EFFLUENT
LIMITATION BY METAL CATEGORY
PARAMETERS
PARAMETERS SELECTED FOR EFFLUENT LIMITATIONS
Ore
Ores
Uranium, Radium,
and Vanadium Ores
Metal Ores, Not Elsewhere Classified
Anti
Ores
Plati
Ores
pH (Acidity/ Alkalinity)
Total Suspended Solids (TSS)
Oil and Grease
*
Chemical Oxygen Demand (COO)
Lit I
n
I
i
i
Total Organic Carbon (TOC)
Cyanide
Ammonia
Aluminum
Antimony
Arsenic
Beryllium
Cadmium
Chromium
Copper
Fluoride
Lead
Manganese
Mercury
Molybdenum
Nickel
Vanadium
Zinc
Radium
Uranium
i
i
i
IE' H 1C 11 If 11 !
i
i
i
t
i
i
i
a
i
i
^.
o
1
i
i
VI-32
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SECTION VII
CONTROL AND TREATMENT TECHNOLOGY
INTRODUCTION
Waterborne wastes from the mining of metal-ore minerals consist
primarily of suspended solids and metals in solution. The
mineralogy of the ore and associated overburden and the chemical
character of percolating mine waters influence the metal con-
tent of mine wastewater, while solids suspended in the wastewater
are influenced by the methods of mining as well as the physical
nature and general geologic characteristics of the ore. Sep-
arate treatment of mine water by methods other than pH control,
Impoundment for solids removal, or combined treatment with mill
effluents is not common.
The wastewaters from ore milling and beneflciation operations
are characterized by high suspended-solid loads, heavy metals
in solution, dissolved solids, and process reagents added
during the concentration process. Impoundment and settling-
pond facilities, primarily for suspended-solid removal, are in
widespread use in the treatment of mill effluents, and this
treatment technology is effective in removal of other waste-
water components as well. Space requirements and location
often affect the utilization of this widespread treatment
technology and dictate the economics of the operations.
Other treatment technologies for removal of dissolved com-
ponents are, for the most part, well-known but are often
limited in usage by the large volumes of wastewater to be
treated and the costs of such large-scale operations.
The control and treatment of the waterborne wastes found in
the mining and beneficiation of metal-ore minerals are influ-
enced by several factors:
(1) Large volumes of mine water and wastewater from
ore-concentrating operations to be controlled and
treated.
(2) Seasonal, as well as daily, variations in the
amount and chemical characteristics of mine water
influenced by precipitation, runoff, and
underground-water contributions.
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(3) Differences in wastewater composition caused by
ore mineralogy and processing techniques and
reagents.
(4) Geographic location and climatic conditions.
(Treatment and control technology selection and
economics are influenced by the amount of water
to be handled.)
CONTROL PRACTICES AND TECHNOLOGY
Control technology, as discussed in this report, includes
techniques and practices employed before, during, and after
the actual mining or milling operation to reduce or eliminate
adverse environmental effects resulting from the discharge
of mine or mill wastewater. Effective pollution-control
planning can reduce pollutant contributions from active
mining and milling sites and can also minimize post-operational
pollution potential. Because pollution potential may not cease
with closure of a mine or mill, control measures also refer to
methods practiced after an operation has terminated production
of ore or concentrated product. The presence of pits, storage
areas for spoil (non-ore material, or waste), tailing ponds,
disturbed areas, and other results or effects of mining or
milling operations necessitates integrated plans for reclamation,
stabilization, and control to return the affected areas to a
condition at least fully capable of supporting the uses which it
was capable of supporting prior to any mining and to achieve a
stability not posing any threat of water diminution, or
pollution and to minimize potential hazards associated with
closed operations.
Mining Techniques
Mining techniques can effectively reduce amounts of pollutants
coming from a mine area by containment within the mine area or
by reducing their formation. These techniques can be combined
with careful reclamation planning and Implementation to provide
maximum at-source pollution control.
Several techniques have been implemented to reduce environmental
degradation during strip-mining operations. Utilization of the
box-cut technique in moderate- and shallow-slope contour mining
VII-2
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has increased recently because more stringent environmental
controls are being Implemented.
A box cut is simply a contour strip mine in which a low-wall
barrier is maintained. Spoil may be piled on the low wall side.
This technique significantly reduces the amount of water dis-
charged from a pit area, since that water is prevented from
seeping through spoil banks. The problems of preventing slides,
spoil erosion, and resulting stream sedimentation are still
present, however.
Block-cut mining was developed to facilitate regrading, mini-
mize overburden handling, and contain spoil within mining
areas. In block-cut mining, contour stripping is typically
accomplished by throwing spoil from the bench onto downslope
areas. This downslope material can slump or rapidly erode
and must be moved upslope to the mine site if contour regrading
is desired. The land area affected by contour strip mining is
substantially larger than the area from which the ores are
extracted. When using block-cut mining, only material from
the first cut Is deposited in adjacent low areas. Remaining
spoil is then placed in mined portions of the bench. Spoil
handling is restricted to the actual pit area for all areas
but the first cut, which significantly reduces the area dis-
turbed .
Pollution-control technology in underground mining is largely
restricted to at-source methods of reducing water influx into
mine workings. Infiltration from strata surrounding the
workings is the primary source of water, and this water reacts
with air and sulfide minerals within the mines to create acid-
pH conditions and, thus, to Increase the potential for solubili-
zatlon of metals. Underground mines are, therefore, faced
with problems of water handling and mine-drainage treatment.
Open-pit mines, on the other hand, receive both direct rainfall
and runoff contributions, as well as infiltrated water from
intercepted strata.
Infiltration in underground mines generally results from rainfall
recharge of a ground-water reservoir. Rock fracture zones,
joints, and faults have a strong influence on ground-water
flow patterns since they can collect and convey large volumes
of water. These zones and faults can intersect any portion
of an underground mine and permit easy access of ground water.
In some mines, infiltration can result in huge volumes of
VI I-3
DRAFT
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DRAFT
water that must be handled and treated. Pumping can be a
major part of the mining operation In terms of equipment and
expense—particularly, In mines which do not discharge by
gravity.
Water-Infiltration control techniques, designed to reduce
the amount of water entering the workings, are extremely
Important In underground mines located In or adjacent to
water-bearing strata. These techniques are often employed
in such mines to decrease the volume of water requiring handling
and treatment, to make the mine workable, and to control
energy costs associated with dewatering. The techniques
Include pressure grouting of fissures which are- entry points
for water Into the mine. New polymer-based grouting materials
have been developed which should improve the effectiveness
of such grouting procedures. In severe cases, pilot holes
can be drilled ahead of actual mining areas to determine
if excessive water is likely to be encountered. When water
is encountered, a small pilot hole can be easily filled by
pressure grouting, and mining activity may be directed toward
non-water-contributing areas in the formation. The feasi-
bility of such control is a function of the structure of the
ore body, the type of surrounding rock, and the characteristics
of ground water in the area.
Decreased water volume, however, does not necessarily mean
that wastewater pollutant loading will also decrease. In
underground mines, oxygen, In the presence of humidity,
interacts with minerals on the mine walls and floor to permit
pollutant formation e.g., acid mine water, while water flowing
through the mine transports pollutants to the outside. If the
volume of this water is decreased but the volume of pollutants
remains unchanged, the resultant smaller discharge will contain
increased pollutant concentrations, but approximately the same
pollutant load. Rapid pumpout of the mine can, however, reduce
the contact time and significantly reduce the formation of
pollutants.
Reduction of mine discharge volume can reduce water handling
costs. In cases of acid mine drainage, for example, the same
amounts of neutralizing agents will be required because pollu-
tant loads will remain unchanged. The volume of mine water
to be treated, however, will be reduced significantly, together
with the size of the necessary treatment and settling facilities.
This cost reduction, along with cost savings which can be
VII-4
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attributed to decreased pumping volumes (hence, smaller pumps,
lower energy requirements, and smaller treatment facilities),
makes use of water infiltration-control techniques highly
desirable.
Water entering underground mines may pass vertically through
the mine roof from rock formation above. These rock units
may have well-developed joint systems (fractures along which
no movement occurs), which tend to facilitate vertical flow.
Roof collapses can also cause widespread fracturing in over-
lying rocks, as well as joint separation far above the mine
roof. Opened joints may channel flow from overlying aquifers
(water-bearing rocks), a flooded mine above, or even from
the surface.
Fracturing of overlying strata is reduced by employing any
or all of several methods: (1) Increasing pillar size; (2)
Increasing support of the roof; (3) Limiting the number of
mine entries and reducing mine entry widths; (4) Backfilling
of the mined areas with waste material.
Surface mines are often responsible for collecting and
conveying large quantities of surface water to adjacent or
underlying underground mines. Ungraded surface mines often
collect water in open pits when no surface discharge point
is available. That water may subsequently enter the ground-
water system and then percolate into an underground mine.
The influx of water to underground mines from either active
or abandoned surface mines can be significantly reduced
through implementation of a well-designed reclamation plan.
The only actual underground mining technique developed
specifically for pollution control is preplanned flooding.
This technique is primarily one of mine design, in which a
mine is planned from its inception for post-operation
flooding or zero discharge. In drift mines and shallow slope
or shaft mines, this is generally achieved by working the
mine with the dip of the rock (inclination of the rock to
the horizontal) and pumping out the water which collects
in the shafts. Upon completion of mining activities, the
mine is allowed to flood naturally, eliminating the possibility
of acid formation caused by the contact between sulfide
minerals and oxygen. Discharges, if any, from a flooded
mine should contain a much lower pollutant concentration.
A flooded mine may also be sealed.
VII-5
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Surface-Water Control
Pollution-control technology related to mining areas, ore-
beneficiation facilites, and waste-disposal sites is generally
designed for prevention of pollution of surface waters (i.e.,
streams, impoundments, and surface runoff). Prior planning
for waste disposal is a prime control method. Disposal sites
should be Isolated from surface flows and impoundments to
prevent or minimize pollution potential. In addition, several
techniques are practiced to prevent water pollution:
(1) Construction of a clay or other type of liner
beneath the planned waste disposal area to prevent
infiltration of surface water (precipitation) or
water contained in the waste into the ground-water
system.
(2) Compaction of waste material to reduce infiltration.
(3) Maintenance of uniformly sized refuse to enhance
good compaction (which may require additional
crushing).
(4) Construction of a clay liner over the material
to minimize infiltration. This is usually succeeded
by placement of topsoll and seeding to establish
a vegetative cover for erosion protection and
runoff control.
(5) Excavation of diversion ditches surrounding the
refuse disposal site to exclude surface runoff
from the area. These ditches can also be used
to collect seepage from refuse piles, with subse-
quent treatment, if necessary.
Surface runoff in the immediate area of beneficiation facili-
ties presents another potential pollution problem. Runoff
from haul roads, areas near conveyors, and ore storage piles
is a potential source of pollutant loading to nearby surface
waters. Several current industry practices to control this
pollution are:
VII-6
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(1) Construction of ditches surrounding storage areas
to divert surface runoff and collect seepage that
does occur.
(2) Establishment of a vegetative cover of grasses in
areas of potential sheet wash and erosion to
stabilize the material, to control erosion
and sedimentation, and to Improve the aesthetic
aspects of the area.
(3) Installation of hard surfaces on haul roads,
beneath conveyors, etc., with proper slopes to
direct drainage to a sump. Collected waters
may be pumped to an existing treatment facility
for treatment.
Another potential problem associated with construction of
tailing-pond treatment systems is the use of existing valleys
and natural drainage areas for impoundment of mine water or
mill process wastewater. The capacity of these Impoundment
systems frequently is not large enough to prevent high
discharge flow rates—particularly, during the late winter
and early spring months. The use of ditches, flumes, pipes,
trench drains, and dikes will assist in preventing runoff
caused by snowmelt, rainfall, or streams from entering
Impoundments. Very often, this runoff flow is the only
factor preventing attainment of zero discharge. Diversion of
natural runoff from impoundment treatment systems, or con-
struction of these facilities in locations which do not
obstruct natural drainage, is therefore, desirable.
Ditches may be constructed upslope from the impoundment to
prevent water from entering it. These ditches also convey
water away and reduce the total volume of water which must be
treated. This may result in decreased treatment costs, which
could offset the costs of diversion.
Segregation or Combination £f_ Mine and Mill Wastewaters
A widely adopted control practice in the ore mining and dressing
industry is the use of mine water as a source of process water.
In many areas, this is a highly desirable practice, because
it serves as a water-conservation measure. Waste constituents
may thus be concentrated into one waste stream for treatment.
VII-7
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In other cases, however, this practice results In the necessity
for discharge from a mill-water impoundment system because,
even with recycle of part of the process water, a net positive
water balance results.
At several sites visited as part of this study, degradation
of the mine water quality is caused by combining the waste-
water streams for treatment at one location. A negative
efxect results because water with low pollutant loading serves
to dilute water of higher pollutant loading. This often
results in decreased water-treatment efficiency because con-
centrated waste streams can often be treated more effectively
than dilute waste streams. The mine water in these cases may
be treated by relatively simple methods; while the volume of
wastewater treated in the mill impoundment system will be
reduced, this water will be treated with increased efficiency.
There are also locations where the use of mine water as process
water has resulted in an improvement in the ultimate effluent.
Choice of the options to segregate or combine wastewater
treatment for mines and mills must be made on an individual
basis, taking into account the character of the wastewater
to be treated (at both the mine and the mill), the water
balance in the mine/mill system, local climate, and topography.
The ability of a particular operation to meet zero or reduced
effluent levels may be dependent upon this decision at each
location.
Regradlnfl
Surface mining may often require removal of large amounts
of overburden to expose the ores to be exploited. Regrading
Involves mass movement of material following ore extraction
to achieve a more desirable land configuration. Reasons for
regrading strip mined land are:
(1) aesthetic Improvement of land surface
(2) returning usefulness to land
(3) providing a suitable base for revegetation
(4) burying pollution-forming materials, e.g. heavy metals
(5) reducing erosion and subsequent sedimentation
(6) eliminating landslidlng
(7) encouraging natural drainage
(8) eliminating ponding
(9) eliminating hazards such as high cliffs and deep pits
(10) controlling water pollution
VII-8
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Contour regrading is currently the required reclamation
technique for many of the nations's active contour and area
surface mines. This technique involves regrading a mine to
approximate original land contour. It is generally one of
the most favored and aesthetically pleasing regrading tech-
niques because the land is returned to its approximate pre-
mined state. This technique is also favored because nearly
all spoil is placed back in the pit, eliminating oversteepened
downslope spoil banks and reducing the size of erodable re-
claimed area. Contour regrading facilitates deep burial of
pollution-forming materials and minimizes contact time
between regraded spoil and surface runoff, thereby reducing
erosion and pollution formation.
However, there are also several disadvantages to contour
regrading that must be considered. In area and contour
stripping, there may be other forms of reclamation that provide
land configurations and slopes better suited to the intended
uses of the land. This can be particularly true with steep-
slope contour strips, where large, high walls and steep final
spoil slopes limit application of contour regrading. Mining
is, therefore, frequently prohibited in such areas, although
there may be other regrading techniques that could be
effectively utilized. In addition, where extremely thick
ore bodies are mined beneath shallow overburden, there may
not be sufficient spoil material remaining to return the land
to the original contour.
There are several other reclamation techniques of varying
effectiveness which have been utilized in both active and
abandoned mines. These techniques include terrace, swale,
swallow-tail, and Georgia V-ditch, several of which are quite
similar in nature. In employing these techniques, the upper
high-wall portion is frequently left exposed or backfilled
at a steep angle, with the spoil outslope remaining somewhat
steeper than the original contour. In all cases, a terrace of
some form remains where the original bench was located, and
there are provisions for rapidly channeling runoff from the
spoil area. Such terraces may permit more effective utiliza-
tion of surface-mined land in many cases.
Disposal of excess spoil material is frequently a problem
where contour backfilling is not practiced. However, the
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same problem can also occur, although less commonly, where
contour regrading is in use. Some types of overburden rock—
particularly, tightly packed sandstones—substantially expand
in volume when they are blasted and moved. As a result,
there may be a large volume of spoil material that cannot be
returned to the pit area, even when contour backfilling is
employed. To solve this problem, head-of-hollow fill has been
used for overburden storage. The extra overburden is placed
in narrow, steep-sided hollows in compacted layers 1.2 to 2.4
meters (4 to 8 feet) thick and graded to control surface drainage.
In this regrading and spoil storage technique, natural ground
is cleared of woody vegetation, and rock drains are constructed
where natural drains exist, except in areas where Inundation
has occurred, This permits ground water and natural percola-
tion to leave fill areas without saturating the fill, thereby
reducing potential landslide and erosion problems. Normally,
the face of the fill is terrace graded to minimize erosion of
the steep outslope area.
This technique of fill or spoil material deposition has been
limited to relatively narrow, steep-sided ravines that can be
adequately filled and graded. Design considerations include
the total number of acres in the watershed above a proposed
head-of-hollow fill, as well as the drainage, slope stability,
and prospective land use. Revegetation usually proceeds as
soon as erosion and siltation protection have been completed.
This technique is avoided in areas where under-drainage materials
contain high concentrations of pollutants, since the resultant
drainage would require treatment to meet pollution-control
requirements.
Erosion Control
Although regrading is the most essential part of surface-mine
reclamation, it cannot be considered a total reclamation tech-
nique. There are many other facets of surface-mine reclamation
that are equally important in achieving successful reclamation.
The effectivenesses of regrading and other control techniques
are Interdependent. Failure of any phase could severly reduce
the effectiveness of an entire reclamation project.
The most Important auxiliary reclamation procedures employed
at regraded surface mines or refuse areas are water diversion
and erosion and runoff control. Water diversion involves
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collection of water before It enters a mine area and conveyance
of that water around the mine site, as discussed previously.
This procedure decreases erosion and pollution formation.
Ditches are usually excavated upslope from a mine site to
collect and convey water. Flumes and pipes are used to carry
water down steep slopes or across regraded areas. Riprap and
dumped rock are sometimes used to reduce water velocity in the
conveyance system.
Diversion and conveyance systems are designed to accommodate
predicted water volumes and velocities. If the capacity of
a ditch is exceeded, water erodes the sides and renders the
ditch ineffective.
Water diversion is also employed as an actual part of the
mining procedure. Drainways at the bases of high walls intercept
and divert discharging ground water prior to its contact with
pollution-forming materials. In some instances, ground water
above the mine site is pumped out before it enters the mine
area, where it would become polluted and require treatment.
Soil erosion is significantly reduced on regraded areas by
controlling the course of surface-water runoff, using inter-
ception channels constructed on the regraded surface.
Water that reaches a mine site, such as direct rainfall, can
cause serious erosion, sedimentation, and pollution problems.
Runoff-control techniques are available to effectively deal
with this water, but these techniques may conflict with
pollution-control measures. Control of chemical pollutants
forming at a mine frequently involves reduction of water
infiltration, while runoff controls to prevent erosion usually
increase infiltration, which can subsequently increase
pollutant formation.
There are a large number of techniques in use for controlling
runoff, with highly variable costs and degrees of effectiveness.
Mulching is sometimes used as a temporary measure which protects
the runoff surface from raindrop impacts and reduces the velo-
city of surface runoff.
Velocity reduction is a critical facet of runoff control.
This is accomplished through slope reduction by terracing or
grading; revegetatlon; or use of flow impediments such as
dikes, contour plowing, and dumped rock. Surface stabilizers
have been utilized on the surface to temporarily reduce eroda-
bility of the material itself, but expense has restricted use
of such materials in the past.
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Revegetation
Establishment of good vegetative cover on a mine area Is
probably the most effective method of controlling runoff and
erosion. A critical factor In mine revegetatlon Is the quality
of the soil or spoil material on the surface of a regraded mine.
There are several methods by which the nature of this material
has been controlled. Topsoll segregation during stripping Is
mandatory In many states. This permits topsoll to be replaced
on a regraded surface prior to revegetatlon. However, In
many forested, steep-sloped areas, there Is little or no top-
soil on the undisturbed land surface. In such areas, overburden
material Is segregated In a manner that will allow the most
toxic materials to be placed at the base of the regraded mine,
and the best spoil material Is placed on the mine surface.
Vegetative cover provides effective erosion control; contri-
butes significantly to chemical pollution control; results
in aesthetic improvement; and can return land to agricultural,
recreational, or silvicultural usefulness. A dense ground
cover stabilizes the surface (with its root system), reduces
velocity of surface runoff, helps build humus on the surface,
and can virtually eliminate erosion. A soil profile begins
to form, followed by a complete soil ecosystem. This soil
profile acts as an oxygen barrier, reducing the amount of
oxygen reaching underlying materials. This, in turn, reduces
oxidation, which is a major contributing factor to pollutant
formation.
The soil profile also tends to act as a sponge that retains
water near the surface, as opposed to the original loose spoil
(which allowed rapid infiltration). This water evaporates
from the mine surface, cooling it and enhancing vegetative
growth. Evaporated water also bypasses toxic materials under-
lying the soil, decreasing pollution production.. The vegetation
itself also utilizes large quantities of water in its life
processes and transpires it back to the atmosphere, again
reducing the amount of water reaching underlying materials.
Establishment of an adequate vegetative cover at a mine site
is dependent on a number of related factors. The regraded
surface of many spoils cannot support a good vegetative cover
without supplemental treatment. The surface texture is often
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too irregular, requiring the use of raking to remove as much
rock as possible and to decrease the average grain size of
the remaining material. Materials toxic to plant life, usually
buried during regrading, generally do not appear on or near
the final graded surface. If the surface is compacted, it is
usually loosened by discing, plowing, or roto-tilling prior
to seeding in order to enhance plant growth.
Soil supplements are often required to establish a good vege-
tative cover on surface-mined lands and refuse piles, which are
generally deficient in nutrients. Mine spoils are often acidic,
and lime must be added to adjust the pH to the tolerance range
of the species to be planted. It may be necessary to apply
additional neutralizing material to revegetated areas for some
time to offset continued pollutant generation.
Several potentially effective soil supplements are currently
undergoing research and experimentation. Flyash is a waste
product of coal-fired boilers and resembles soil with respect
to certain physical and chemical properties. Flyash is often
alkaline, contains some plant nutrients, and possesses moisture-
retaining and soil-conditioning capabilities. Its main function
is that of an alkalinity source and a soil conditioner, although
it must usually be augmented with lime and fertilizers. How-
ever, flyash can vary drastically in quality—particularly,
with respect to pH—and'may contain leachable materials capable
of producing water pollution. Future research, demonstration,
and monitoring of flyash supplements will probably develop the
potential use of such materials.
Limestone screenings are also an effective long-term neutra-
lizing agent for acidic spoils. Such spoils generally continue
to produce acidity as oxidation continues. Use of lime for
direct planting upon these surfaces is effective, but it
provides only short-term alkalinity. The lime is usually
consumed after several years, and the spoil may return to its
acidic condition. Limestone screenings are of larger particle
size and should continue to produce alkalinity on a decreasing
scale for many years, after which a vegetative cover should be
well-established. Use of large quantities of limestone should
also add alkalinity to receiving streams. These screenings
are often cheaper than lime, providing larger quantities of
alkalinity for the same cost. Such applications of limestone
are currently being demonstrated in several areas.
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Use of digested sewage sludge as a soil supplement also has
good possibilities for replacing fertilizer and simultaneously
alleviating the problem of sludge disposal. Sewage sludge
is currently being utilized for revegetation in strip-mined
areas of Ohio. Besides supplying various nutrients, sewage
sludge can reduce acidity or alkalinity and effectively increase
soil absorption and moisture-retention capabilities. Digested
se./age sludge can be applied in liquid or dry form and must
be incorporated into the spoil surface. Liquid sludge applica-
tions require large holding ponds or tank trucks, from which
sludge is pumped and sprayed over the ground, allowed to dry,
and disced into the underlying material. Dry sludge application
requires dryspreadlng machinery and must be followed by discing.
Limestone, digested sewage sludge, and flyash are all limited
by their availabilities and chemical compositions. Unlike
commercial fertilizers, the chemical compositions of these
materials may vary greatly, depending on how and where they
are produced. Therefore, a nearby supply of these supplements
may be useless if it does not contain the nutrients or pH
adjusters that are deficient in the area of intended application.
Flyash, digested sewage sludge, and limestone screenings are
all waste products of other processes and are, therefore, usually
inexpensive. The major expense related to utilization of any
of these wastes is the cost of transporting and applying the
material to the mine area. Application may be quite costly
and must be uniform to effect complete and even revegetation.
When such large amounts of certain chemical nutrients are
utilized, it may also be necessary to institute controls to
prevent chemical pollution of adjacent waterways. Nutrient
controls may consist of preselection of vegetation to absorb
certain chemicals, or of construction of berms and retention
basins in which runoff can be collected and sampled, after
which it can be discharged or pumped back to the spoil.
The specific soil supplements and application rates employed
are selected to provide the best possible conditions for the
vegetative species that are to be planted.
Careful consideration should be given to species selection
in surface-mine reclamation. Species are selected according
to some land-use plan, based upon the degree of pollution con-
trol to be achieved and the site environment. A dense ground
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cover of grasses and legumes is generally planted, in addition
to tree seedlings, to rapidly check erosion and siltation.
Trees are frequently planted in areas of poor slope stability
to help control landsliding. Intended future use of the land
is an important consideration with respect to species selection.
Reclaimed surface-mined lands are occasionally returned to
high-use categories, such as agriculture, If the land has
potential for growing crops. However, when toxic spoils are
encountered, agricultural potential is greatly reduced, and
only a few species will grow.
Environmental conditions—particularly, climate—are important
in species selection. Usually, species are planted that are
native to an area—particularly, species that have been success-
fully established on nearby mine areas with similar climate
and spoil conditions.
Revegetation of arid and semi-arid areas involves special
consideration because of the extreme difficulty of establishing
vegetation. Lack of rainfall and effects of surface distur-
bance create hostile growth conditions. Because mining in
arid regions has only recently been initiated on a large scale,
there is no standard revegetation technology. Experimentation
and demonstration projects exploring two general revegetation
techniques—moisture retention and irrigation—are currently
being conducted to solve this problem.
Moisture retention utilizes entrapment, concentration, and
preservation of water within a soil structure to support vege-
tation. This may be obtained utilizing snow fences, mulches,
pits, and other methods.
Irrigation can be achieved by pumping or by gravity, through
either pipes or ditches. This technique can be extremely
expensive, and acquisition of water rights may present a major
problem. Use of these arid-climate revegetation techniques in
conjunction with careful overburden segregation and regrading
should permit return of arid mined areas to their natural states.
Exploration, Development, and Pilot-Scale Operations
Exploration activities commonly employ drilling, blasting,
excavation, tunneling, and other techniques to discover,
locate, or define the extent of an ore body. These activities
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vary from small-scale (such as a single drill hole) to large-
scale (such as excavation of an open pit or outcrop face).
Such activities frequently contribute to the pollutant loading
in wastewater emanating from the site. Since available facili-
ties (such as power sources) and ready accessibility of special
equipment and supplies often are limited, sophisticated treat-
ment is often not possible. In cases where exploration activity
is being carried out, the scale of such operations is such that
primary water-quality problems Involve the presence of increased
suspended-solid loads and potentially severe pH changes. Ponds
should be provided for settling and retention of wastewater,
drilling fluids, or runoff from the site. Simple, accurate
field tests for pH can be made, with subsequent pH adjustment
by addition of lime (or other neutralizing agents).
Protection of receiving waters will thus be accomplished, with
the possible additional benefits of removal of metals from
solution—either in connection with solids removal or by precipi-
tation from solution.
Development operations frequently are large-scale, compared to
exploration activities, because they are intended to extend
already known or currently exploited resources. Because these
operations are associated with facilities and equipment already
in existence, it is necessary to plan development activities to
minimize pollution potential, and to use existing mine or mill
treatment and control methods and facilities. These operations
should, therefore, be subject to limitations equivalent to
existing operations with respect to effluent treatment and
control.
Pilot-scale operations often involve small to relatively large
mining and beneficiation facilities even though they may not
be currently operating at full capacity or are in the process
of development to full-scale. Planning of such operations
should be undertaken with treatment and control of wastewater
in mind to ensure that effluent limitation guidelines and
standards of performance for the category or subcategory will
be met. Although total loadings from such operations and
facilites are not at the levels expected from normal operating
conditions, the compositions of wastes and the concentrations
of wastewater parameters are likely to be similar. Therefore,
Implementation of recommended treatment and control technologies
must be accomplished.
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Mine and Mill Closure
Mine Closure (Underground). Unless well-planned and well-
designed abatement techniques are implemented, an underground
mine can be a permanent source of water pollution.
Responsibility for the prevention of any adverse environmental
impacts from the temporary or permanent closure of a deep mine
should rest solely and permanently with the mine operator.
This constitutes a substantial burden; therefore, it behooves
the operator to make use of the best technology available for
dealing with pollution problems associated with mine closure.
The two techniques most frequently utilized in deep-mine pollu-
tion abatement are treatment and mine sealing. Treatment tech-
nology is well defined and is generally capable of producing
acceptable mine effluent quality. If the mine operator chooses
this course, he is faced with the prospect of costly permanent
treatment of each mine discharge.
Mine sealing is an attractive alternative to the prospects of
perpetual treatment. Mine sealing requires the mine operator
to consider barrier and ceiling-support design from the perspec-
tives of strength, mine safety, their ability to withstand
high water pressure, and their utility for retarding ground-
water seepage. In the case of new mines, these considerations
should be included in the mine design to cover the eventual
mine closure. In the case of existing mines, these considera-
tions should be evaluated for existing mine barriers and ceiling
supports, and the future mine plan should be adjusted to include
these considerations if mine sealing is to be employed at mine
closure.
Sealing eliminates the mine discharge and inundates the mine
workings, thereby reducing or terminating the production of
pollutants. However, the possibility of the failure of mine
seals or outcrop barriers increases with time as the sealed
mine workings gradually became inundated by ground water and
the hydraulic head increases. Depending upon the rate of
ground-water influx and the size of the mined area, complete
inundation of a sealed mine may require several decades.
Consequently, the maximum anticipated hydraulic head on the
mine seals may not be realized for that length of time. In
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addition, seepage through, or failure of, the barrier or mine
seal could occur at any time. Therefore, the mine operator
should be required to permanently maintain the seals, or to
provide treatment in the event of seepage or failure.
Mine Closure (Surface). The objectives of proper reclamation
management of closed surface mines and associated workings are
tc (1) restore the affected lands to a condition at least fully
capable of supporting the uses which they were capable of
supporting prior to any mining, and (2) achieve a stability
which does not pose any threat to public health, safety, or
water pollution. With proper planning and management during
mining activities, it is often possible to minimize the amount
of land disturbed or excavated at any one time. In preparation
for the day the operation may cease, a reclamation schedule
for restoration of existing affected areas, as well as those
which will be affected, should be specified. The use of a planned
methodology such as this will return the workings to their
premined condition at a faster rate, as well as possibly reduce
the ultimate costs to the operator.
To accomplish the objectives of the desired reclamation goals,
it Is mandatory that the surface-mine operator regrade and
revegetate the disturbed area during, or upon completion of,
mining. The final regraded surface configuration is dependent
upon the ultimate land use of the specific site, and control
practices described in this report can be incorporated into
the regradlng plan to minimize erosion and sedimentation.
The operator should establish a diverse and permanent vegeta-
tive cover and a plant succession at least equal in extent of
cover to the natural vegetation of the area. To assure compli-
ance with these requirements and permanence of vegetative cover,
the operator should be held responsible for successful revege-
tatlon and effluent water quality for a period of five full
years after the last year of augmented seeding. In areas of
the country where the annual average precipitation is 64 cm
(26 in.) or less, the operator's assumption of responsibility
and liability should extend for a period of ten full years
after the last year of augmented seeding, fertilization, irri-
gation, or effluent treatment.
Mill Closure. As with closed mines, a beneficiation faci-
lity's potential contributions to water pollution do not cease
upon shutdown of the facility. Tailing ponds, waste or refuse
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piles, haulage areas, workings, dumps, storage areas, and
processing and shipping areas often present serious problems
with respect to contributions to water pollution. Among the
most important are tailing ponds, waste piles, and dump areas.
Failure of tailing ponds can have catastrophic consequences,
with respect to both immediate safety and water quality.
To protect against catastrophic occurrences, tailing ponds
should be designed to accommodate, without overflow, an abnormal
storm which is observed every 25 years. Since no wastewater
is contributed from the processing of ores (the facility being
closed), the ponds will gradually become dewatered by evapora-
tion or by percolation into the subsurface. The structural
integrity of .the tailing-pond walls should be periodically
examined and, if necessary, repairs made. Seeding and vegetation
can assist in stabilizing the walls, prevent erosion and sedi-
mentation, lessen the probability of structural failure, and
improve the aesthetics of the area.
Refuse, waste, and tailing piles should be recontoured and
revegetated to return the topography as near as possible to
the condition it was in before the activity. Techniques
employed in surface-mine regrading and revegetatlon should be
utilized. Where mills are located adjacent to mine workings,
the mines can be refilled with tailings. Care should be taken
to minimize disruption of local drainage and to ensure that
erosion and sedimentation will not result. Maintenance of
such refuse or waste piles and tailing-disposal areas should
be performed for at least five years after the last year of
regrading and augmented seeding. In areas of the country
where the annual average precipitation is 64 cm (26 in.)
or less, the operator's assumption of responsibility should
extend for a period of ten full years after the last year of
augmented seeding, fertilization, irrigation, or effluent
treatment.
TREATMENT TECHNOLOGY
Each of the techniques currently employed in the ore mining
and dressing industry, as well as advanced waste treatment
technology which might be employed in present or future opera-
tions, is discussed in this section.
The treatment technologies currently practiced in the ore
mining and dressing industry encompass a wide variety of tech-
niques ranging from the very simple to the highly sophisticated.
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While a limited number of basic treatment practices are standard
(settling or tailing ponds, pH control, etc.) and employed
at almost all operations, individual operations have approached
specific pollution problems in many different ways.
Impoundment Systems
This group of systems utilizes treatment technology which is
primarily designed to deal with suspended solids, but which is
frequently used with such other techniques as pH control, to
accomplish removal of dissolved constituents as well.
Tailing Ponds. This type of treatment is the most common
treatment technique used in the ore mining and dressing industry
today. The design of a tailing pond is primarily for suspended-
solid removal and retention. Such a pond must be large enough
to provide sufficient retention time and quiescent conditions
conducive to settling. If properly designed, and if retention
time and surface area are sufficient, a tailing pond may also
effect to some degree the stabilization of oxidizable consti-
tuents as well as the balancing of influent quality and quantity
fluctuations and the storage of storm water.
Tailing ponds are often situated to capitalize upon natural
terrain factors in order to minimize the requirements for dam
construction. The containment dam is often constructed of
available earth and rock materials, as well as tailings. In
other cases, concrete basins may be constructed. Because of
natural terrain conditions, they may be constructed using one,
two, three, or even four walls. The containment dam must be
raised periodically to accommodate the rising level of contained
tailings and water. In most cases, the basin provides perpetual
storage for any materials settled out of the water treated.
Typically, a concrete basin is periodically dredged and the
solids stored in a waste-disposal area. Retention time in
ponds has been reported to vary from as little as four hours
to as much as several months at average flow conditions (for
discharging systems).
Water leaves a tailing pond by decantatlon, evaporation,
seepage through- the dam or to underlying materials, or by discharge.
Decanted water may be recycled for use in the mill, discharged,
or treated further. In some operations, in arid or semi-arid
areas, evaporation from the tailing-pond surface may equal the
rate of input, allowing zero-discharge operation of the pond
without recycle of water.
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Seepage losses from tailing ponds may flow into permeable
underlying strata and enter ground water, or may flow through
the containment dam and result in surface flows of water.
Where dam construction has made use of tailings, some seepage
will almost always be observed. Resulting seepage waters are
often collected in ditches and pumped back into the tailing
pond. Seepage may also be limited by the use of pond liners
of various materials (clay, asphalt, plastic, etc.).
Low-cost, relatively simple construction and the ability to
perform multiple functions simultaneously have led to the wide
acceptance of tailing ponds as a prime treatment and tailing-
disposal method utilized by the ore mining and dressing industry.
There are a number of problems associated with the utilization
of tailing ponds as treatment facilities, however. Improper
design of inlet and discharge locations, Insufficient size
and number, and insufficient retention time are the most common
problems. Algal growths in tailing ponds are quite common
during warm months, a factor which may influence such effluent
water-quality parameters as TOG, COD, and BOD. A minimum
retention time of 30 days and the added capability of retaining
runoff associated with a storm likely to occur once in 20 years
are recommended by one source (Reference 29).
The relative advantages and disadvantages of a tailing pond as
a treatment system are listed below.
Advantages
Disadvantages
Performs large number of
treatment processes—parti-
cularly, suspended-solid
removal.
Can achieve high treatment
efficiency and often pro-
duce acceptable effluent
quality.
Often, only practical means
of long-term solids disposal
Lacks responsive means of control;
difficult to optimize large number
of processes performed.
Covers large surface area—may
contribute high net precipitation
to overall water balance; land
availability and topography Influ-
ence location.
Creates potentially severe rehabili-
tation problem if tailings contain
sulfide minerals.
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Disadvantages
Often difficult to isolate from
contributing drainage areas—
storm water influences retention.
Subject to climatic variations—
particularly, thermal skimming and
seasonal variation in bio-oxidation
efficiency.
Often difficult to ensure good
flow distribution.
Requires careful control of
seepage through dams.
Installation expensive in some
situations, due to high cost of
retaining structures.
Advantages
Large retention has a balan-
cing effect on effluent
quality.
Large surface area aids
oxidation and evaporation.
Can often be constructed
using mining equipment
and materials.
Little operating expertise
normally required.
Commonly used treatment
method, familiar to
industry.
Clear supernatant water may
serve as a reservoir for
reuse.
Tailing ponds in the ore mining and dressing industry range
from pits to large, engineered structures of 1000 acres with
massive retaining dams. For large tailing dams, wall heights
of 200 feet or more have been reached by building up the dams
over a period of time.
Routinely reached levels of suspended-solid concentrations in
treated effluent range from 10 to 30 mg/1 at mines and mills
visited or surveyed as part of this study. In tailing ponds
with decant structures for recycle of water, levels in excess
of 50 mg/1 of suspended solids were rarely observed.
Settling Ponds. Settling ponds differ from tailing ponds
primarily in size and in the concentrations of influent solids
treated. In general, relatively low initial solid loads are
removed, necessitating only occasional dredging to maintain
adequate settling volume behind the dam. Suspended-solid
removal to very low levels is often possible when initial
concentrations of suspended solids are low. Settling ponds
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find their greatest usefulness in association with mines having
low wastewater solids loads. Effluent levels of A.3 mg/1,
6.2 mg/1, and 17 mg/1, for example, were observed from three
different settling ponds.
Such ponds may serve a variety of purposes in addition to
removal of suspended solids, including COD reduction and cooling.
As basins for a variety of chemical treatments, they can
provide sufficient retention time for completion of reactions,
for pH control, for chemical precipitation, and for the removal
of solids produced.
Secondary Settling Ponds. Settling ponds or tailing ponds are
frequently used in a multiple arrangement. The purpose of
this scheme is to further reduce suspended-solid loading in
the sequential ponds and to allow the subsequent use of precipi-
tation or pH control before discharge or recycle. The ponds
enable further reduction in suspended solids and in dissolved
parameters. An excellent example is the use of secondary
settling ponds (sometimes called polishing ponds) in the copre-
cipitation of radium with barium. Removal of radium could be car-
ried out in the tailing pond, but the pollutant radium would be
distributed over a large area, presenting a hazard to workers
and a potential future hazard from the use of abandoned tailings
in construction (a practical use in some areas). By using a
small secondary settling pond to coprecipitate radium, the
pollutant is confined to an area to which access can be res-
tricted, even after the operation is closed. Similar considera-
tions would suggest the use of such ponds for the precipitation
of heavy metals by lime or sulflde treatments.
Clarifiers and Thickeners
A method of removing large amounts of suspended solids from
wastewater is the use of clarifiers or thickeners, which are
essentially large tanks with directing and segregating systems.
The design of these devices provides for concentration and
removal of suspended and settleable solids in one effluent
stream and a clarified liquid in the other. Clarified waters
may be produced which have extremely low solids content through
proper design and application.
Clarifiers are not generally capable of handling tailing-solid
levels above about 50 percent, due to the necessity for rake
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operation and hydraulic transport of suspended solids from
the device. The concentration from a mine-water clarifier
at one site, for example, is 3 mg/1 suspended solids.
Clarifiers may range in design from simple units to more complex
systems involving sludge blanket pulsing or sludge recycle to
improve settling and increase the density of the sludge.
Settled solids from clarifiers are removed periodically or
continuously for either disposal or recovery of contained
values. Thickeners are used when the main purpose is to produce
a clarified overflow with a concentrated tailing effluent in
the underflow.
Thickeners have a number of distinct advantages over settling
or tailing ponds:
(1) Less land space Is required. Area-for-area, these
devices are much more efficient in settling capacity
than ponds.
(2) Influences of rainfall are reduced compared to ponds.
If desired, the clarifiers and thickeners can be
covered.
(3) Since the external construction of clarifiers and
thickeners consists of concrete or steel (in the
form of tanks), ground-seepage and rain-water runoff
influences do not exist.
(4) Thickeners can generally be placed adjacent to a
mill, making reclaim water available nearby with
minimal pumping requirements.
The use of clarifiers and thickeners, together with tailing
or settling ponds, may improve treatment efficiency; reduce
the area needed for tailing ponds; and facilitate the reuse
or recycle of water in the milling operation. The use of
flocculants to enhance the performance of thickeners and
clarifiers is common practice.
Clarifiers and thickeners also suffer some distinct disadvan-
tages compared to ponds:
(1) They have mechanical parts and, thus, require
maintenance.
VII-24
DRAFT
-------
DRAFT
(2) They have limited storage capacity for either
clarified water or settled solids.
(3) The internal sweeps and agitators in thickeners
and clarifiers require more power and energy for
operation than ponds.
Flocculation
This treatment process consists basically of adding reagents
to the treated waste stream to promote settling of suspended
solids. The solids may be deposited in tailing ponds (where
high suspended solids are involved) or in clarifier tanks (in
cases of lower solids loads).
Flocculating agents Increase the efficiency of settling facili-
ties and are of several general types: ferric compounds, lime,
aluminum sulfate, and cationic or anionic polyelectrolytes.
Causticized wheat and corn starch have also been used. The ionic
types, such as alum, ferrous sulfate, lime, and ferric chloride,
function by destroying the repelling double-layer ionic charges
around the suspended particles and thereby allowing the particles
to attract each other and agglomerate. Polymeric types function
by forming physical bridges from one particle to another and
thereby agglomerating the particles. Recyclable magnesium
carbonate has also been proposed as a flocculant in domestic
water treatment.
Flocculating agents are added to the water to be treated under
controlled conditions of concentration, pH, mixing time, and
temperature. They act to upset the stability of the colloidal
suspension by charge neutralization and flocculation of sus-
pended solids, thus Increasing the effective diameter of these
solids and increasing their subsequent settling rate.
Flocculating agents are most commonly used after the larger,
more readily settled particles (and loads) have been removed
by a settling pond, hydrocyclone, or other treatment. Agglomer-
ation, or flocculation, can then be achieved with less reagent,
and with less settling load on the polishing pond or clarifier.
Flocculation agents can be used with minor modifications and
additions to existing treatment systems, but the costs for
VII-25
DRAFT
-------
DRAFT
the flocculating chemicals are often significant. Ionic types
are used in concentrations of 10 to 100 mg/1 in the wastewater,
while the highest-priced polymeric types are effective in
concentrations of 2 to 20 mg/1.
The effectiveness and performance of individual flocculating
systems may vary over a substantial range with respect to
suspended-solid removal, accessory removal of soluble com-
ponents by adsorptive phenomena, and operating characteristics
and costs. Specific system performance must be analyzed and
optimized with respect to mixing time, flocculant addition
level, settling (detection) time, thermal and wind-induced
mixing, and other factors.
Centrifugation
Centrifugation, which may be considered as a form of forced
or assisted settling, may be feasible in specific control
applications. With the volume of gross wastewater flows
at most mine/mill complexes, it is probable that centrifuga-
tion may be more applicable to component in-process waste
streams. The presence of abrasive components or significant
amounts of solid material smaller than approximately 5
micrometers in diameter in the treated water would tend to
disqualify Centrifugation as a solid-removal option.
Hydrocyclones
While hydrocyclones are widely used in the separation, classifi-
cation, and recovery operations involved in mineral processing,
they are used only Infrequently for wastewater treatment.
Even the smallest-diameter units available (stream-velocity
and centrifugal-separation forces both increase as the diameter
decreases) are ineffective when particle size is less than
25 to 50 micrometers. Larger particle sizes are relatively
easy to settle by means of small ponds, thickeners or clarifiers,
or other gravity-principle settling devices. It is the smaller
suspended particles that are the most difficult to remove, and
it is these that cannot be removed by hydrocyclones but may
be handled by ponds or other settling technology. Also, hydro-
cyclones are of doubtful effectiveness when flocculating agents
are used to increase settling rates. This method is generally
most effective in the 25- to 200-micrometer size range for
particles.
VII-26
DRAFT
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DRAFT
Filtration
Filtration is accomplished by passing the wastewater stream
through solid-retaining screens or cloths or particulate
materials such as sand, gravel, coal, or diatomaceous earth
using gravity, pressure, or vacuum as the driving force.
Filtration is a versatile method in that it can be used to
remove a wide range of suspended particle sizes.
Filtration is not generally useful on mill waste streams, due
to the high volumes of material involved. The large volumes
to be treated would require large filters. The cost of these
units, and their relative complexity compared to settling ponds,
has restricted their use.
A variety of filtration techniques, including disc and drum
units, find process applications and may be applicable to some
waste streams—particularly, where segregated waste streams
require special treatment.
Likely applications of filtration include pretreatment of input streams
using reverse-osmosis and ion-exchange units (discussed later).
High values contained in suspended solids may, in some cases,
offset the capital and operating expenses of filtering systems.
The use of filtration as a normal unit process in treating
uranium-mill tailings for value recovery through countercurrent
washing is indicative of the possible use of filtration in
tailing treatment. In this instance, the final washed tail
filter cake is reslurried for transport to the tailing pond.
In situations where biological treatment of component or
combined waste streams is required to reduce BOD, COD, or
bacterial loads, trickling filters may be required, but their
application as primary treatment for the bulk mine or mill
effluent is considered unlikely.
The specific applicability and size specifications for filter
modules must be evaluated on a case-by-case basis, taking into
account the process stream characteristics, solids filter-
ability, desired dryness of filter cake, and other parameters.
Ultimate clarification of filtered water will be a function of
particle size, filter-media porosity, filtration rate, and
other variables. In general, for the majority of mine or
mill waste waters subjected to this treatment, post-treatment
VII-27
DRAFT
-------
DRAFT
suspended-solid levels of less Chan 20 percent of influent
loadings are anticipated. Thus, If used after primary floccu-
lation and settling, suspended solids levels of 20 mg/1 should
be obtainable.
Neutralization
Adjustment of pH is the simplest and most common treatment
practiced in the mining and milling Industry today. The addi-
tion of either acidic or basic constituents to a wastewater
stream to achieve neutralization generally Influences the
behavior of both suspended and dissolved components. In most
instances of interest in mining and milling activities, waste-
waters are treated by base addition to achieve pH conditions
in the range of 6 to 9.
Acid waste streams (considerably more common than highly basic
effluents) may be neutralized by addition of a variety of
basic reagents, including lime (calcium oxide), limestone,
dolomite (CaMg(C03)2J, magnesite (MgC03), sodium hydroxide,
soda ash (sodium carbonate), ammonium hydroxide, and others
to raise the pH of treated waste streams to the desired level.
Lime is most often used because it is inexpensive and easy to
apply. Soda ash and caustic soda are commonly used to supply
alkalinity in leaching and hydrometallurgical processes, where
the formation of calcium precipitates would be objectionable,
but the cost advantages of using lime generally preclude the
use of soda ash and caustic soda in large-scale waste treatment.
Ammonia neutralization is most frequently a processing techni-
que, where ammonia affords a strong advantage in being volatile
in the final product, allowing the recovery of nearly pure
oxides. In waste treatment, its volatility is a disadvantage
because of the COD it presents, its toxicity, and the produc-
tion of undesirable nitrites and nitrates as oxidation products.
Its use is not widespread, although ammonia neutralization of
a wastewater stream is practiced at one site in the ferroalloy-
ore mining and milling category.
Excessively basic waste streams are not common but may be
neutralized by addition of an acid—most commonly, sulfuric.
Since many heavy metals form insoluble hydroxides in highly
basic solutions, sedimentation prior to neutralization may
prevent the resolubillzation of these materials and may
simplify subsequent waste-treatment requirements. Carbon
dioxide has also been used to adjust the pH of effluent
waters to acceptable levels prior to discharge (recarbona-
tion).
VII-28
DRAFT
-------
DRAFT
Essentially any wastewater stream may be treated to a final pH
within the range of 6 to 9. Generally, the stream will be
sufficiently uniform to allow adequate pH control based only
on the volume of flow and predetermined dosage rates, with
periodic adjustments based on effluent pH. Automated systems
which monitor and continously adjust the concentration
of reagents added to the wastewater are also currently available.
As discussed previously, pH control is often used to control
solubility (also discussed under Chemical Precipitation Pro-
cesses) . Examples of pH control being used for precipitating
undesired pollutants are:
(1) Fe(+3) + 30H(-)-*Fe(OH)3_
(2) Mn(+2) + 20H(-)->Mn(OH)2_+ 2H(+) + 4e(-)
(3) Zn(+2) + OH(-)-»Zn(OH)2_
(4) Pb(+2) + 20H--»Pb(OH)2_
(5) Cu + 20H(-) -»Cu(OH)2_
Reaction (1) is used for removal of iron contaminants. Reaction
(2) is used for removal of manganese from manganese-containing
wastewater. Reactions (3), (4), and (5) are used on wastewater
containing copper, lead, and zinc salts. The use of lime to
attain a pH of 7 will theoretically reduce heavy metals to these
levels (Reference 30):
Metal Concentration (mg/1 at pH 7)
Cu(+2) 0.2 to 0.3
Zn(+2) 1.0 to 2.5
Cd(+2) 1.0
Ni(+2) 1.0
Cr(+2) 0.4
The careful control of pH, therefore, has other ancillary bene-
fits, as illustrated above. The use of pH and solubility rela-
tionships to Improve removal of wastewater contaminants is
further developed below.
VII-29
DRAFT
-------
DRAFT
Chemical Precipitation Processes
The removal of materials from solution by the addition of
chemicals which form insoluble (or sparingly soluble) compounds
with them is a common practice in hydrometallurgical ore bene-
ficiation and in waste treatment in the ore mining and dressing
industry. It is especially useful for the removal of heavy
metals from mine effluents and process wastes.
To be successful, direct precipitation depends primarily upon
two factors:
(1) Achievement of a sufficient excess of the added
ion to drive the precipitation reaction to completion.
(2) Removal of the resulting solids from the waste stream.
If the first requirement is not met, a portion of the pollutant(s)
will be removed from solution, and desired effluent levels may
not be achieved. Failure to remove the precipitates formed
prior to discharge is likely to lead to redissolutlon, since
ionic equilibria in the receiving stream will not, in general,
be those created in treatment. Effective sedimentation or
filtration is, thus, a vital component of a precipitation
treatment system and frequently limits the overall removal
efficiency. Sedimentation may be effected in the tailing basin
Itself, in secondary or auxllliary settling ponds, or in
clarifiers. Industry experience has shown the value of treat-
ment of wastes prior to delivery to the tailing impoundment.
Benefits derived include: improved settling of precipitates
due to interaction with tailings; simplified disposal of sludges;
and, generally, suppressed solubility of materials in tailing
solids.
The use of precipitation for wastewater treatment varies from
lime treatments (to precipitate sulfates, fluorides, hydroxides,
and carbonates) to sodium sulfide precipitation of copper, lead,
and other toxic heavy metals. The following equations are
examples of precipitation reactions used for wastewater treat-
ment:
(1) Fe(+3) + Ca(OH)2_ » Ca(+2) + Fe(OH)3_
(2) Mn(+2) + Ca(OH)2_ —> Ca(+2) + Mn(OH)2_
(3) Zn(+2) + Na2C03_ —> Na(+) + ZnC03_
.VII-30
DRAFT
-------
DRAFT
(4) S04/-2) + Ca(OH)£ > CaSOA. + 20H(-)
(5) 2F(-) + Ca(OH)2 > CaF2_ + 20H(-)
One drawback of the precipitation reactions is that the varying
solubility of metal species and the possibility of widely diver-
gent formation and precipitation rates limit the ability of
this treatment to deal with all waste constituents.
Lime Precipitation. The use of lime to cause chemical preci-
pitation has gained widespread use in the ore mining and dressing
Industry because of its ease of handling, because of its economy,
and because of its effectiveness in treatment of a great variety
of dissolved materials. The use of other bases is, of course,
possible, as previously discussed. However, the use of lime
as a treatment reagent Is probably the best-known and best-
studied method.
A typical lime neutralization/precipitation system is illustrated
in Figure VII-1, Generally, water is pumped or discharged to
a holding or settling pond, where suspended-solid levels are
reduced. Either in conjunction with the primary pond itself
or in a mixing basin or tank, a slurry of lime and water is
delivered for mixing with the wastewater stream. Secondary
settling ponds are then used to collect the usually high volumes
of sludges which may be recovered. These impoundments may
be dredged periodically to remove sludges, or the sides of the
basin may be built up. Discharge of the water then usually
takes place.
The treatment conditions, dosages, and final pH must be optimized
for any given waste stream, but, In general, attainment of a
pH of at least 9 is necessary to ensure removal of heavy metals.
To attain desired levels of control for many heavy metals, it
is necessary to attain a pH of 10 to 12 in many instances.
The levels of concentration attainable In an actual operating
system may vary from the limits predicted on the basis of purely
theoretical considerations, but extremely low levels of metals
discharged have been reached by the use of this treatment method.
Figure VII-2 Illustrates the theoretical solubilities of several
metal ions as a function of pH. The minimum pH value for complete
precipitation of metal ions as hydroxides is shown in Figure VII-3.
VI1-31
DRAFT
-------
DRAFT
Figure VIM. LIME NEUTRALIZATION AND PRECIPITATION PROCESS FOR
TREATMENT OF MINE WATER PRIOR TO DISCHARGE
FROM MINE OR MILL
SLUDGE
REQUIRING
DISPOSAL
TO
DISCHARGE
SOURCE: Reference 31
VII-32
DRAFT
-------
DRAFT
Figure VII-2. THEORETICAL SOLUBILITIES OF METAL IONS AS A FUNCTION OF pH
0.0001
6 7
SOURCE: Reference 32
VII-33
DRAFT
-------
DRAFT
Figure VI1-3. MINIMUM pH VALUE FOR COMPLETE PRECIPITATION OF METAL IONS AS
HYDROXIDES
PH
in n
9.0
o n
7.0
6.0
5.0
4.0
3.0
2.0
1.0
0.0
•
i
I
7.2
^^m
(
5.2
4.2 '
1.3
••
12
IM
i.4
1.3
«
1.5
>.7
1
i
D.i
MM
1
Sn+2 Fe+3 AI+3 Rb+2 Cu+2 Zn+2 Nj+2 Fe+2 Cd+2 Mn+2
LIME
NEUTRALIZATION
LIME PRECIPITATION
SOURCE: Reference 31
VII-34
DRAFT
-------
DRAFT
An example of the performance of lime precipitation at elevated
pH is given for Fe, Fb, Zn, Cd, Hg, and F in Figure VII-4.
These data are taken from a representative mill, where removal
efficiency is plotted against pH. The curves are not always
complete for lack of data; it is not advisable to extrapolate
them without further measurements, because chemical changes
may occur that reverse an apparent consistent trend.
Purely theoretical considerations of metal-hydroxide solubility
relationships suggest that the metal levels tabulated below
are attainable (Reference 29).
Final Concentration
(microgram per liter) p_H
1 to 8 9.5
10 to 60 10
Pb 1 8
Fe(total) 1 8 (if totally Ferric)
Many factors, such as the effects of widely differing solu-
bility products, mixed-metal hydroxide complexing, and metal
chelation, render these levels of only limited value when
assessing attainable concentrations in a treatment system.
Among the metals effectively removed at basic pH are: As, Cd,
Cu, Cr(-f3), Fe, Mn, Hi, Pb, and Zn. Based upon published
sources, industry data, and analysis of samples, it appears
that the concentrations given in the tabulation below may be
routinely and reliably attained by hydroxide precipitation
in the ferroalloy-ore mining and milling industry.
Metal Concentration Metal Concentration
(mg/1) (mg/1)
As 0.50 Mn 1.0
Cd 0.50 Ni 0.05
Cu 0.05 Pb 0.10
Cr(+3) 0.05 Zn 0.15
Fe 1.0
VII-35
DRAFT
-------
DRAFT
Figure VI1-4. HEAVY-METAL PRECIPITATION vs pH FOR TAILING-POND
EFFLUENT pH ADJUSTMENTS BY LIME ADDITION
50
60
#
H
o
iu
S 70
an
w
90
100
11
13
pH
SOURCE: Reference 33
VII-36
DRAFT
-------
DRAFT
Some metallic pollutants of interest in the uranium-ore mining
and milling industry, together with results produced by lime
precipitation in conjunction with a rise in pH from 6.7 to
12.7, are shown below:
Metal Concentration (mg/1)
pH°6.7 pH=12.7
Cd 1.3 less than 0.02
Fe 6.0 less than 0.1
Ni 0.13 less than 0.05
Cu 5.3 0.05
Zn 31.25 0.11
Mn 26.5 0.04
The use of lime or limestone will also provide some removal of
fluoride ion, by precipitation of calcium fluoride. The
theoretical limit of fluoride-ion solubility in a solution
containing 10 mg/1 of calcium, for example, is less than 4 mg/1.
Other examples of the efficiency of lime precipitation as a
treatment method are discussed by ore category later in this
section. An Important point is Illustrated in the data pre-
viously presented here, however. All metals do not remain In
solution at elevated pH. Examples of that phenomenon are
the variations in solubilities of lead and zinc, which are
precipitated at approximately pH 9. Above pH 9, these metals
rapidly resolubllize. Lime precipitation has not been demon-
strated as being effective for removal of tellurium, molybdenum,
arsenic, and mercury at high pH, because these metals become
more soluble with increasing alkalinity.
Sulfide Precipitation. The use of sulfide ion as a precipi-
tant for removal of heavy metals accomplishes more complete
removal than the use of hydroxide for precipitation. Sulfide
precipitation is currently being used In wastewater treatment
to reduce mercury levels to extremely low levels (Reference
34). Highly effective removal of Cd, Cu, Co, Fe, Hg, Mn,
Ni, Pb, Zn, and other metals from mine and mill wastes can be
VII-37
DRAFT
-------
DRAFT
accomplished by treatment with either sodium sulfide or hydrogen
sulfide. The use of this method depends somewhat on the avail-
ability of methods for effectively removing precipitated solids
from the waste stream, and on removal of the solids to an
environment where reoxidation is unlikely.
Several steps enter into the process of sulfide precipitation:
(1) Preparation of sodium sulfide. Although this product
is often in oversupply from byproduct sources, it
can also be made by the reduction of sodium sulfate,
a waste product of acid-leach milling. The process
Involves an energy loss in the partial oxidation of
carbon (such as that contained in coal).
Na2S04. + 4C - > Na2S + 4CO )gas)
(2) Precipitation of the pollutant metal (M) in the
waste stream by an excess of sodium sulfide:
Na2S + MS04_ — > MS (precipitate) + Na2_S04_
(3) Physical separation of the metal sulfide in thickeners
or clarlfiers, with reducing conditions maintained
by excess sulfide ion.
(4) Oxidation of excess sulfide by aeration:
Na2S + 202 — >
This process usually involves iron as an intermediary
and is seen to regenerate unused sodium sulfate.
On the whole, sulfide precipitation removes both heavy metals
and some sulfur from waste streams but requires some energy
expenditure.
In practice, sulfide precipitation can be applied only when
the pH is sufficiently high (greater than about 8) to assure
generation of sulfide ion rather than bisulfide or hydrogen
sulfide gas. It is then possible to add just enough sulfide,
in the form of sodium sulfide, to precipitate the heavy metals
present as cations; alternatively, the process can be continued
until dissolved oxygen in the effluent is reduced to sulfate
and aerobic conditions are obtained. Under these conditions,
VII-38
DRAFT
-------
DRAFT
some reduction and precipitation of molybdates, uranates,
chromates, and vanadates may occur, but ion exchange seems more
appropriate for the removal of these anions.
Due to the toxiclty of sulfide ion, and of hydrogen sulfide
gas, the use of sulfide precipitation may require both pre-
and post-treatment and close control of reagent additions.
Pretreatment involves raising the pH of the waste stream to
minimize evolution of H2S, which would pose a safety hazard
to personnel. If desirable, this may be accomplished at
essentially the same point as the sulfide treatment, or by
addition of a solution containing both sodium sulfide and a
strong base (such as caustic soda). The sulfides of many
heavy metals, such as copper and mercury, are sufficiently
Insoluble to allow essentially complete removal with extremely
low residual sulfide levels. Treatment for these metals with
close control on sulfide concentrations could be accomplished
without the need for additional treatment. Adequate aeration
should be provided to yield an effluent saturated with oxygen.
Coprecipitation. In coprecipitation, materials which cannot
be removed from solution effectively by direct precipitation
are removed by incorporating them into particles of another
precipitate, which is separated by settling, filtration, or
another technique such as flotation. Current practice is
exemplified by the use of barium chloride addition for radium
control in the uranium industry.
Radium sulfate, one of the least soluble substances, is soluble
to 20 micrograms per liter, while allowable concentrations in
drinking water are about 6 million times less. The process of
coprecipitation for radium separation was perfected by M.S.
Curie and has been used extensively in radiochemistry. The
carrier for radium is barium, usually added as barium chloride
(BaC12) in a concentration of about 10 mg/1 and In the presence
of more sulfate ion than is necessary to precipitate barium
sulfate (BaSOA). Almost all RaSOA^ that is present is copreci-
pitated, and removal to a level of about 1 picocurie (1 pc/1)
or 1 plcogram per liter, is current practice. The results of
tests on the addition of BaC12^ BaS04^, and BaC03_ to neutral
and acidic effluents are shown in Table VII-1.
The importance of coprecipitation in the ferroalloy industry
has been demonstrated by extensive experiments (References
35 and 36). In that work, molybdenum, which appears
VII-39
DRAFT
-------
DRAFT
TABLE VIM. RESULTS OF COPRECIPITATION REMOVAL OF
RADIUM FROM WASTEWATER
EFFLUENT pH
Neutral
Acidic
REAGENT
BaS04
BaCOj
BaCI2
BaC03
BaC12
REAGENT
ADDITION
(rngfi.}
300
1000
100
200
30
60
100
200
100
200
300
100
PRE- AND POST-PRECIPITATION
RADIUM CONCENTRATIONS
(pc/£)
BEFORE
100
300
470
490
800
440
400
430
160
150
150
150
AFTER
30
70
30
40
20
6
2
2
18
20
30
6 to 15
X RADIUM
REMOVED
70
77
94
92
97
99
99
99
88
87
80
90 to 97
VII-40
DRAFT
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DRAFT
in effluents from many mines and mills as the molybdate (Mo04_-)
anion (which is not removed effectively by hydroxide or sulfide
precipitation), is removed by incorporation into ferric hydrox-
ide precipitates formed at acid pH (4.5 optimum) by the addition
of ferric sulfate or ferric chloride (at levels of about 100
mg/1). Removal of resulting precipitates by filtration and
flotation has been reported to yield effluents containing 0.2
mg/1 for mill waters initially containing 4.9 mg/1 of molybdenum
(Reference 37). In a pilot-plant study using ferric sulfate
and flotation recovery of precipitates, removal of more than
95 percent of influent molybdenum, to levels of 0.02 to 0.1
mg/1, has been obtained.
Since the process used for molybdenum removal is performed at
acid pH, it is necessary to acidify the (typically, alkaline)
mill waste stream after separation of solids in the tailing
pond effects the molybdenum removal. A base is then added to
neutralize the effluent prior to discharge. For large waste
stream flow, reagent costs may be an important consideration.
Although molybdenum values are concentrated to about 5 percent
in the precipitates removed, they do not appear to represent
a marketable product at this time.
Other Precipitation Systems. Other types of precipitation
systems have been employed, such as those used for the preci-
pitation of sulfate (Reference 38), fluoride (as calcium
fluoride), or others (Reference 39). Starch-xanthate com-
plexes have recently been reported to be effective in aiding
precipitation of a variety of metals, Including Cd, Cr, Cu,
Pb, Hg, Ni, Ag, and Zn (Reference 40). Scavenging or
coprecipitation studies have been conducted on municpal waste-
waters (Reference 41). In specialized cases, precipitation
may be induced by oxidation, which produces a less soluble
heavy-metal product. The chlorine oxidation of Co(+2) to Co
(+3) at a pH of approximately 5 produces the insoluble Co^3_
(xH20). Oxidation of Fe(+2) to Fe(+3) results in the precipi-
tation of hydrous ferric oxide, even at relatively low pH.
Oxidation of As(+3) to As(+4) improves precipitation removal
(Reference 40). The use of oxidation is further discussed
later in this section.
VII-41
DRAFT
-------
DRAFT
Reduction
Reduction techniques have particular applicability to the removal
of hexavalent chromium and copper from waste streams in the
ferroalloy-ore mining and milling industry. Copper is often
recovered in current practice by reduction of the metal and
subsequent deposition on scrap iron in the waste stream (cementa-
tion) . Since the effluent levels resulting from cementation
are still high, generally 10 mg/1 or more, it is necessary to
follow use of this process with another removal step, such as
hydroxide precipitation.
Reduction of chromates to trivalent chromium, with subsequent
precipitation of the chromium as the hydroxide, is a standard
waste-treatment practice in a number of industries and may find
application in the ore mining and dressing industry, where
leaching practices give rise to wastewater contaminated with
chromates. Commonly used reducing agents include sulfur
dioxide and ferrous salts of iron. With sulfur dioxide and a
pH of 2.5, chrornate may be reduced rapidly and completely.
Removal of the Cr(OH)3^ precipitate formed in treatment of the
relatively dilute wastes to be expected in mill effluents may
prove difficult, necessitating careful management of the
treatment system and the use of flocculants such as Fe(OH)3^
to aid in settling. Effluent levels of 0.5 mg/1 of total chromium
and 0.05 mg/1 of hexavalent chromium may be reliably attained
by the treatment (Reference 42).
Sodium borohydrlde reduction has been applied to reducing
soluble mercury levels in chlor-alkali and mercury processing
plants and to reducing lead levels in wastes arising in the
tetra-alkyllead manufacturing process (U.S. Patents 3,736,253,
3,764,528, and 3,770,423). Stannous (tin) compounds have been
used for the reductive deposition of palladium during electro-
plating processes. Electroreduction of metals is widely practiced
in electrowlnning and electrorefining systems for copper, nickel,
cobalt, and other metals.
Treatment in the ore mining and dressing industry differs from
the above techniques, chiefly because of the lower concentra-
tions of soluble, reducible species and because of the presence
of numerous other reducible species in the wastewater. Unless
preconditioning of treated waters is employed, excessive reducing-
agent consumption may occur. Secondary recovery systems (settling,
filters, etc.) may be necessary to permit removal of reduced
VII-42
DRAFT
-------
DRAFT
components. The recovery of values from waste residues is a
potential option with this treatment method. In some instances,
application of this process option to internal streams prior
to discharge and/or combination with other waste streams may
offer substantial enhancement of value recovery from treatment
products.
Oxidation, Aeration, and Air Stripping
A number of the waste components resulting from mining and
milling may be removed or rendered less harmful by oxidation
or removal to the atmosphere. Among these are cyanide, sulfide,
ammonia, and a variety of materials presenting high COD levels.
The simplest approach to effecting these processes is aeration
of the waste stream, which occurs naturally in pumping it and
in distributing it at the tailing pond. More elaborate imple-
mentation achieves more complete and rapid results in air strippers,
and by controlled introduction of stronger oxidants, such as
chlorine or ozone.
Cyanide (CN-) is removed by oxidation to cyanate (CNO-) and,
ultimately, to C02^ and N2^. This is accomplished in standard
practice by rapid chlorination at alkaline pH (about 10.5)
using caustic soda. The probable reaction with excess chlorine
has been expressed as:
2NaCN + 5C1^ + 12NaOH > N£ + 2Na2C03, + lONaCl + 6H20
A pH of 10 to 11 is recommended for operating conditions.
This process may be performed on either a batch or continuous
process. Approximately 2.72 kg (6 Ib) each of caustic soda
and chlorine are normally required to oxidize 0.45 kg (1 Ib) of
cyanide. If metal-cyanide complexes are present, extended
chlorination for several hours may be necessary.
In treatment of mill effluent in the gold milling industry,
some cyanide is lost in the process and is present in the mill
tailings. Some of the cyanide decomposes in the tailing pond,
and it appears that a high level of removal is generally effected
by naturally occurring oxidation in tailing ponds. Except
where cyanide is used as a leaching reagent, high concentrations
of cyanide are not normally encountered. The use of cyanide as
a depressant in the flotation process is an additional source of
cyanide in wastewater. Effluent levels characteristically
encountered are less than 0.05 mg/1 total cyanide.
VII-43
DRAFT
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DRAFT
Effective and proper use of chlorlnation or ozonation should
result in complete destruction of cyanide in mill treatment
systems. At locations where very low levels are encountered
in wastewater streams, aeration devices, auxiliary ponds, or
long retention times may provide removal to below acceptable
levels.
Ammonia used in a solvent extraction and precipitation opera-
tion at one milling site is removed from the mill waste stream
by air stripping. The countercurrent-flow air stripper used
at this plant operates with a pH of 11 to 11.7 and an air /liquid
flow ratio of 0.83 cubic meter of air per liter water (110
cubic feet of air per gallon of water). Seventy-five percent
removal of ammonia is achieved by reducing total nitrogen
levels for the mill effluent to less 5 mg/1, 2 mg/1 of which
is in the form of nitrates. Ammonia may also be removed from
waste streams through oxidation to nitrate by aeration—or, more
rapidly, by ozonation—or use of chemical oxidants, although
these procedures are less desirable due to the impact of
nitrates on the receiving water.
The removal of a variety of COD-producing pollutants from effluent
streams by oxidation in the tailing ponds and/or delivery lines
is evident in data from visited sites. Where high reagent
dosages or other process factors lead to elevated effluent
COD levels, aeration or the use of stronger oxidants may be of
value. In general, the use of strong oxidants in the tailing
pond will be highly undesirable, since the oxidation of sulfide
minerals in the tails can lead to increased acid production
and greater solubility of ore constituents, including heavy
metals. Aeration will be best practiced in other impoundments
also.
Adsorption
Activated carbon is a sorptive material characterized by high
surface area within its internal pore system. Pores generally
range from 10 to 100 Angstrom units (0.001 to 0.01 micrometer),
and surface areas of up to 1000 square meters/gram are considered
normal for carbons of this type. Due to the dimensions of the
pores, to the highly convoluted internal surface (and, thus, very
high surface area), and to the residual organic contents of
carboxyic, carbonyl, and hydroxyl compounds, activated carbon
VII-44
DRAFT
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DRAFT
exhibits adsorptive, absorptive, and slight residual ion-
exchange capabilities. In contrast to alumina, silica gel,
and other adsorbents, however, activated carbon exhibits a
relatively low affinity for water. Compounds which are readily
removed by activated carbon include aromatica, phenolics,
chlorinated hydrocarbons, surfactants, organic dyes, organic
acids, higher-molecular-weight alcohols, and amines. Current
applications of this material also center around the control
and removal of color, taste, and odor components in water.
Activated carbon has been shown to significantly reduce concen-
trations of a variety of inorganic salts, including most heavy
metals. Lead concentrations have been reduced from 100 mg/1
to 0.5 mg/1 (Reference 43). Reports of Hg, V, Cr, Pb, Ni,
Cd, Zn, Fe, Mn, Ca, Al, Bi, Ge, As, Ba, Se, and Cu removal have
appeared in the literature—most often, as results of laboratory-
scale treatment (References 44 and 40).
In addition to use in tertiary sewage treatment, activated carbon
has found a variety of industrial-waste applications. At one
facility, phenols are removed from 600 cubic meters (150,000
gallons) per day of chemical plant wastewater containing 62,000
mg/1 of total dissolved solids (Reference 45). Influent
and effluent levels for this treatment facility are 100 mg/1
and less than 1 mg/1 of phenol, respectively. As in this operation,
carbon may be regenerated in a furnace with approximately
95-percent carbon recovery to reduce materials cost for the
operation.
In addition to the economics of operation dictating regenerative
processes, recovery of metal values using the principles of this
treatment is possible. Some indication of the economic success
of this approach may be gained from the reported viability of
the "resin-in-pulp" or "carbon-in-pulp" process employed at mill
4105 in the gold-recovery circuit. In this case, cyano-complexes
of gold (and, probably, other metals) are reversibly adsorbed
from alkaline solution by activated carbon. Activated-carbon
treatment of acid mine water has been used for iron (+2) removal
(Reference 46).
The application of carbon adsorption, or adsorption by other
materials (such as peat), to mining and milling wastewater is
more likely to be limited by cost than by technical feasibility.
Removal of flotation or solvent-extraction reagents from waste
streams may be practical in some operations, If waste streams
VII-45
DRAFT
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DRAFT
are segregated. Carbon adsorption could be an important factor
in achieving a high degree of water recycle in flotation mills
where reagents or decomposition products in the feed water
would interfere with processing.
Other Adsorption Methods. While activated carbon is one specific
adsorbent used for wastewater treatment, there are many addi-
tional materials which show varying adsorptive capacities for
wastawater constituents. Many of these candidate sorbing media
have been evaluated only in a preliminary fashion under full-
scale conditions, and few of these have been evaluated with
reference to behavior in actual mine/mill effluents.
Reported adsorbing species include tailing materials (Reference
47), waste wool (Reference 48), silica gel, alumina,
hydrous zirconium oxide (Reference 49), peat moss (Refer-
ence 50)f, hydrous manganese oxides (Reference 51), and
others. The sorptive capacity of various soils is currently
under study in conjunction with increased utilization of spray
irrigation as a method of wastewater disposal (Reference 52).
To date, little experience in large-scale wastewater disposal
involving waters similar to mine/mill effluents has been reported
for land disposal by spray irrigation. Capital costs, operating
costs, and performance experience with municipal, food-industry,
and paper-industry waste disposal, however, suggest the potential
desirability of this procedure (Reference 53). Any spray-
irrigation disposal of mine/mill wastes must be preceded by
settling systems or other treatments to reduce the suspended-
solid load.
Ion Exchange
Ion exchange is basically a process for removal of various
ionic species in or on fixed surfaces. During the fixing
process, ions in the matrix are exchanged for soluble ionic
species. Cationic, anionic, and chelating ion exchangers
are available and may be either solid or liquid. Solid ion
exchangers are generally available in granular, membrane, and
bead forms (ion-exchange resins) and may be employed in upflow
or downflow beds or columns, in agitated baskets, or in co-
current- or countercurrent-flow modes. Liquid ion exchangers
are usually employed in equipment similar to that employed
VII-46
DRAFT
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DRAFT
in solvent-extraction operations (pulsed columns), mixed
settlers, rotating-disc columns, etc.)- In practice, solid
resins are probably more likely candidates for end-of-pipe
wastewater treatment, while either liquid or solid ion exchang-
ers may be utilized in internal process streams.
Individual ion-exchange systems do not generally exhibit equal
affinity or capacity for all ionic species (cationic or
anlonic) and, so, may not be suited for broad-spectrum removal
schemes in wastewater treatment. Their behavior and perfor-
mance are usually dependent upon pH, temperature, and concen-
tration, and the highest removal efficiencies are generally
observed for polyvalent ions. In wastewater treatment, some
pretreatment or preconditioning of wastes to adjust suspended-
solid concentrations and other parameters is likely to be
necessary.
Progress in the development of specific ion-exchange resins and
techniques for their application has made the process attrac-
tive for a wide variety of industrial applications in addition
to water softening and deionization. It has been used exten-
sively in hydrometallurgy—particularly, in the uranium indus-
try—and in wastewater treatment (where it often has the
advantage of allowing recovery of marketable products). This
is facilitated by the requirement for periodic stripping or
regeneration of ionic exchangers. If regeneration produces
a solution waste, its subsequent treatment must be considered.
Table VII-2 shows different types of ion-exchange resins and
the range of conditions and variety of purposes for which
they are employed.
Disadvantages of using ion exchange in treatment of mining
and milling wastewater are relatively high costs, somewhat
limited resin capacity, and insufficient specificity—
especially, in cationic exchange resins for some applica-
tions.
Although it is suitable for complete deionization of water,
ion exchange Is generally limited in this application, by
economics and resin capacity, to the treatment of water con-
taining 500 mg/1 or less of total dissolved solids. Since
IDS levels in mining and milling effluents are often higher
than this level, application of ion exchange to the economic
reduction of total dissolved solids at high flow rates must
be evaluated.
VII-47
DRAFT
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DRAFT
TABLE VI1-2. PROPERTIES OF ION EXCHANGERS FOR
METALLURGICAL APPLICATIONS
DESIRED
CHARACTERISTIC
CHEMICAL
STABILITY TO:
PHYSICAL STABILITY FOR:
Acidi
Alkalies
Oxidation
Temperature
Organic Solvents
Removal of weak
acids
Removal of strong
acids
High regeneration
efficiency
High capacity
High porosity
Hydrogen exchange
at low pH
Salt splitting
pH range (operating)
GENERALLY RECOMMENDED APPLICATION
CATION EXCHANGERS
Inorganic
Zeo-Dur
•
6.2 to
8.7
SI
I
•
•
•
6.9 to
7.9
Organic
Sulfonated
Coal
Zeo-Karb
•
•
•
•
0 to
11
Resins
Permutit Q
•
•
•
•
•
•
•
0 to
13
Car-
boxylic
Resin
Permutit H-70
•
•
•
•
•
•
3.5 to
12
ANION EXCHANGERS
Weakly
Basic
Gran-
ular
De-Acidite
•
•
•
•
•
•
0 to
12
Bead
Permutit W
•
•
•
•
•
•
0 to
13.9
Strongly
Basic
Gran-
ular
Permutit A
•
•
•
•
•
•
0 to
13.9
Bead
Permutit S
•
•
•
•
•
•
•
•
0 to
13.9
SOURCE: Reference 54
VII-48
DRAFT
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DRAFT
For recovery of specific ions or groups of ions (e.g., dival-
ent heavy-metal cations, or metal anions such as molybdate,
vanadate, and chromate), ion exchange is applicable to a much
broader range of solutions. This use is typified by the
recovery of uranium from ore leaching solutions using strongly
basic anion-exchange resin. As additional examples, one may
consider the commercial reclamation of chromate plating and
anodizing solutions, and the recovery of copper and zinc
from rayon-production wastewaters (Reference 54).
Chromate plating and anodizing wastes have been purified and
reclaimed by ion exchange on a commercial scale for some time,
yielding economic as well as environmental benefits. In
tests, chromate solutions containing levels in excess of
10 mg/1 chromate, treated by ion exchange at practical resin-
loading values over a large number of loading elution cycles,
consistently produced an effluent containing no more than
0.03 mg/1 of chromate.
High concentrations of ions other than those to be recovered
may interfere with practical removal. Calcium ions, for
example, are generally collected along with the divalent
heavy-metal cations of copper, zinc, lead, etc. High calcium-
ion concentrations, therefore, may make ion-exchange removal
of divalent heavy-metal ions impractical by causing rapid
loading of resins and necessitating unmanageably large resin
inventories and/or very frequent elution steps. Less diffi-
culty of this type is experienced with anion exchange. Avail-
able resins have fairly high selectivity against the common
anions, such as Cl(-) and SOji(-2). Anions adsorbed along
with uranium include vanadate, molybdate, ferric sulfate
anionlc complexes, chlorate, cobalticyanide, and polythionate
anions. Some solutions containing molybdate prove difficult
to elute and have caused problems.
Ion-exchange resin beds may be fouled by particulates, pre-
cipitation within the beds, oils and greases, and biological
growth. Pretreatment of water, as discussed earlier, is
therefore, commonly required for successful operation. Gen-
erally, feed water is required to be treated by coagulation
and filtration for removal of iron and manganese, C02_, H^S,
bacteria and algae, and hardness. Since there is some lati-
tude in selection of the ions that are exchanged for the con-
taminants that are removed, post-treatment may or may not be
required.
VII-49
DRAFT
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DRAFT
Since, in many cases, calcium is present in ore mining and
milling wastewater in appreciably greater concentrations than
are the heavy-metal cations whose removal to low levels is
sought, use of ion exchange in that mode may have several
disadvantages, since costs would be high, and little advan-
tage would be offered over lime or sulfide precipitation.
For the removal of anions, however, the relatively high
costs of ion-exchange equipment and resins may be offset par-
tially or totally by the recovery of a marketable product.
This has been demonstrated in the removal of uranium from
mine water, and the removal of molybdate anions is now under
investigation in pilot-plant studies at two operations,
although results are not yet available. The application of
this technique will depend upon a complex set of factors,
including resin loading achieved, pretreatment required, and
the complexity of processing needed to produce a marketable
product from eluent streams.
The practicality of the ion-exchange process will be enhanced
by practices such as waste segregation, recycle, etc., which
allow the treatment of smaller volumes of more concentrated
solutions. Similar factors apply to the treatment of mining
and milling waste streams bearing vanadate and chromate anions,
although prior experience in ion-exchange recovery of these
materials should aid the development of treatment schemes for
such wastes.
Modified Desal Process. A demonstration plant for generating
potable water from acid coal-mine drainage, in operation since
early 1973, treats 3,028 cubic meters (800,000 gallons) per day
of water which contains pollutant loadings similar to
those of acid mine drainage (Reference 55). The plant
was originally designed for a capacity of 1,893 cubic meters
(500,000 gallons) per day, but it is expected that the plant's
capacity can be further increased to 3,785 cubic meters
(1,000,000 gallons) per day through use of improved operating
techniques.
The Modified Desal Process portrayed in Figure VII-5 is a
variation of a system originally developed to produce potable
water from brackish supplies by means of cation- and anion-
exchange resins. The primary purpose of ion exchange in treat-
ing acid mine water, however, is to remove sulfate, so only
VII-50
DRAFT
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DRAFT
Figure VI1-5. DIAGRAM OF MODIFIED DESAL PROCESS
FROM
MINE
I
I
SETTLING
BASINS
GRAVITY
FILTERS
PRODUCT
WATER
LEGEND
MAIN PROCESS
• ADDITIONS OR LOSSES
• REGENERATION PROCESS
SOURCE: Reference 66
VII-51
DRAFT
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DRAFT
an anion-exchange resin is necessary. The process uses a weak-
base anion resin in the bicarbonate form to replace sulfate or
other anions. The solution of metal bicarbonates is aerated
to oxidize ferrous iron to the ferric form and to purge the
carbon dioxide gas. The increase in pH causes iron, aluminum,
and manganese to precipitate as insoluble hydrous oxides.
Some calcium and magnesium carbonates also precipitate. To
produce improved quality water, well within potable limits,
lime treatment precipitates more calcium and magnesium by
converting the bicarbonates into less soluble carbonates.
The exhausted resin is regenerated with ammonium hydroxide,
which converts the resin to the free-base form. Introduction
of carbon dioxide converts the resin back to the bicarbonate
form, and the regenerated solution of ammonium sulfate is pro-
cessed to recover the ammonia through lime addition. The
resultant calcium sulfate is transported to mine pits for dis-
posal. Regeneration occurs after about 18 hours of operation,
and the plant currently utilizes the original ion-exchange
resin.
Operating data for the plant are shown in Table VII-3. It
is felt that this system, or a modification thereof, might
provide effective removal of sulfate and dissolved solids in
the ore mining and dressing industry.
Present operating costs for water produced at the Phillipsburg,
Pennsylvania, plant are $0.40 to 0.50 per 3.79 cubic meters
(1,000 gallons) of water. However, a considerable reduction
in cost might be achieved for the mining industry for two
reasons. The first is that the demonstration plant contains
much instrumentation and many features that would be unnec-
essary in a facility designed merely for production. Secondly,
integration of the ion-exchange system with presently existing
lime-neutralization plants could eliminate the necessity for
many features of the Modified Desal Process system.
Although the cost for treating 3.79 cubic meters (1,000 gallons)
of raw mine drainage appears favorable, volumes in excess of
57,000 cubic meters (15,000,000 gallons) of drainage generated
daily at many facilities require a substantial total invest-
ment in time, material resources, and energy. Also,
individual treatment plants with design capacities of
up to 34,065 cubic meters (9,000,000 gallons) per day would
necessitate the installation of multiple ion-exchange units
VII-52
DRAFT
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DRAFT
TABLE VII-3. ANALYTICAL DATA FOR MODIFIED DESAL PROCESS
PARAMETER
pH
Total hardness (CaCO3)
TDS
Calcium (CaCO3)
Magnesium (CaCOg)
Iron
Sulfate
CONCENTRATION (mg/£ )
RAW WASTEWATER
3.7»
395
1,084
295
100
101
648
EFFLUENT WATER
9.5"
184
284
85
99
0.2
192
•Value in pH units
VII-53
DRAFT
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DRAFT
at most discharge outfalls. This configuration would greatly
decrease cost effectiveness for a treatment aimed specifically
at removing sulfate and dissolved solids.
Ultrafiltration and Reverse Osmosis
Ultrafiltration and reverse osmosis are similar processes in
which pressure is used to force water through membranes which
do not allow passage of contaminants. They differ in the
scale of contaminants passed and in the pressures required.
Ultrafiltration generally retains particulates and materials
with a molecular weight greater than 500, while reverse-
osmosis membranes generally pass only materials with
a molecular weight below 100 (Sodium chloride, although below
a molecular weight of 100, is retained, allowing application
to desalinization) . Pressures used in Ultrafiltration gen-
erally range from 259 to 517 cm of Hg (50 to 100 psi), while
reverse osmosis is run at pressures ranging from 2,068 to
9,306 cm of Hg (400 to 1,800 psi).
Ultrafiltration has been applied on a significant commercial
scale to the removal of oil from oil emulsion, yielding a
highly purified water effluent and an oil residue sufficiently
concentrated to allow reuse, reclamation, or combustion.
Equipment is readily available, and present-day membranes are
tolerant of a broad pH range. Application of Ultrafiltration
to mining and milling waste streams, where high dosages of
oils are used in flotation—as at a formerly operated man-
ganese mill—may provide a practical technique for removing
these waste components, possibly allowing reuse as well.
Reverse osmosis (RO) is conceptually similar to ultrafiltra-
tion. It also involves the application of an external pressure
to a solution in contact with a semipermeable membrane to
force water through the membrane while excluding both soluble
and insoluble solution constituents. In its rejection of
soluble constituents, reverse osmosis performs a water-treat-
ment function not fulfilled by Ultrafiltration systems under
simple operating conditions.
Reverse osmosis is considerably less tolerant of input-stream
variations in conditions and requires, in general, considerable
pretreatment. Concentration of wastes is generally limited
by saturation of solutions and the formation of precipitates,
VI I-54
DRAFT
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DRAFT
which can decrease the effectiveness of the apparatus. As a
result, residual volumes of waste in the mining and milling
industry would, in many cases, be unmanageably large. A
pilot-plant operation has been run on mine drainage streams,
and production of a high-quality water effluent has been
shown to be technically feasible. Pretreatment requirements,
costs, and the problems of disposal of residual wastes make
the practicality and economic achievability important con-
siderations.
Reverse osmosis has been demonstrated capable of rejecting
heavy-metal species from purified water streams with a high
degree of efficiency (Table VII-4). Reverse-osmosis systems
have been evaluated for acid mine water treatment (References
57 and 58). Related studies have been conducted with metal-
finishing effluents (Reference 59). In most instances,
pretreatment of water, and conditioning with respect to pH,
temperature, and suspended-solid levels, is necessary for
reverse-osmosis module use. Membrane lifetime and constancy
of efficiency are both adversely affected by inadequate treat-
ment of waters prior to membrane contact. In general, labora-
tory performance of reverse-osmosis systems has shown some-
what higher purification efficiencies than have been observed
in pilot-plant operations (Reference 40). The present
state-of-the-art with regard to RO technology indicates that
details of extrapolation of laboratory and current pilot-plant
data to full-scale operation need to be worked out. Data on
membrane lifetime, operating efficiency, rejection specificity,
and other factors remain to be more fully quantified.
High-Density-Sludge Acid Neutralization
The conventional lime neutralization of acid or mine wastes
usually leads to the formation of low-density sludges which
are difficult to dewater (floes). The use of ground lime-
stone avoids this problem but does not allow for the attain-
ment of pH levels necessary to effectively remove such metals
as zinc and cadmium. A process which utilizes extensive
recycle of the previously precipitated sludge allows the
attainment of sludges of much higher density, thus allowing
more rapid sedimentation of the sludges ultimately produced
and easing solid-disposal problems.
VII-55
DRAFT
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TABLE VI1-4. REJECTION OF METAL SALTS BY REVERSE-
OSMOSIS MEMBRANES
PARAMETER
Iron
Magnesium
Copper
Nickel
Chromium (hexavalent)
Strontium
Cadmium
Silver
Aluminum
TYPICAL REJECTION PERCENT
99
98
99
99.2
97.8
99
98
96
99
SOURCE: Reference 40
VII-56
DRAFT
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Solvent Extraction
Solvent extraction is a widely utilized technique for the
separation and/or concentration of metallic and nonmetallic
species in the mineral processing industry. It has been
applied to commercial processing of uranium, vanadium, tung-
sten, thorium, rhenium, rare earths, beryllium, columbium,
copper, zirconium, molybdenum, nickel, boron, phosphoric acid,
and others (References 60 and 61). Reagent-processing
equipment for this technique is highly developed and generally
available (Reference 62). It is anticipated that such
equipment would require modification to be applicable to treat-
ing the low levels of soluble metals in most waste streams.
Pretreatment and post-treatment of waters treated by this
technique would probably be required to control influent pH,
suspended solids, and other parameters, as well as effluent
organic levels. It is likely that this treatment strategy may
be most applicable in internal process streams or as an add-on
for the recovery of values from waste-concentration streams
such as distillate or freeze residues, reverse-osmosis brines,
etc.
Because of the speculative nature of solvent extraction as
applied to wastewater treatment, the unknown costs of rea-
gents, and possible pretreatment/post-treatment demands,
accurate treatment or capital costs for this option do not
appear readily derivable at this time.
Evaporation and Distillation
Evaporation may be employed as a wastewater-treatment tech-
nique in a variety of ways:
(1) Total evaporation of wastewater may produce solid
residues and eliminate effluent water discharge.
(2) Concentration of wastewater by evaporation may
balance dilution by makeup and infiltration water
and allow for an approach to total recycle, thus
minimizing discharge volume. The buildup of detri-
mental species upon evaporation will normally
require a bleed stream from the evaporation system,
thus precluding total water recycle. A bleed stream,
of course, might be handled by total evaporation,
rather than by discharge to a waterway.
VII-57
DRAFT
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(3) Concentration by evaporation may allow subsequent
removal of concentrated wastewater components to
acceptable levels for smaller-volume discharge or
reuse.
(4) Ultimately, complete distillation of wastewater
may allow the almost total reuse or recycle of
contained water, while rendering discharge unnec-
essary and allowing potential recovery of values
from nonvolatile residues. In the absence of
recoverable values, disposal of sludge resulting
from distillation might become a problem of sub-
stantial magnitude. The presence of volatile
wastes in the effluent may require additional
treatment of distillate to achieve adequate quality
for some uses.
Energy sources for evaporation may be artificial (steam, hot
gases, and electricity) or natural (solar, geothermal, etc.).
In present practice, many of the mining and milling operations
in the Western and Southwestern United States employ solar
evaporation as a principal means of water treatment. Evapora-
tive losses of water at some installations may exceed 7,572
cubic meters (2,000,000 gallons) per year for each 0.4 hectare
(1 acre) of evaporative surface; with adequate surface acreage,
this loss may allow for zero-effluent-discharge operation.
At present, this evaporated water is not collected for reuse
at these operations.
A multistage flash-distillation process has been applied to
treat acid mine drainage (from a coal mine) in a pilot plant
(Reference 63 ) • The process is mechanically complex but
results in a solid residue and essentially pure water, suit-
able for human consumption. This approach to pollution con-
trol involves the use of considerable energy associated with
vaporizing vast volumes of water. Its technical applicability
to treating mine water has been demonstrated, but it is not
clear that organic wastes potentially present in mill effluents
would be successfully controlled by such a process.
Techniques for Reduction of_ Wastewater Volume
Pollutant discharges from mining and milling sites may be
reduced by limiting the total volume of_ discharge, as well
VII-58
DRAFT
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as by reducing pollutant concentrations In the waste stream.
Volumes of mine discharges are not, In general, amenable to
control, except Insofar as the mine water may be used as Input
to the milling process In place of water from other sources.
Techniques for reducing discharges of mill wastewater include
limiting water use, excluding incidental water from the waste
stream, recycle of process water, and impoundment (with water
lost to evaporation or seepage).
In most of the industry, water use should be reduced to the
extent practical, because of the existing incentives for doing
so (i.e., the high costs of pumping the high volumes of water
required, limited water availability, and the cost of water-
treatment facilities). Incidental water enters the waste
stream primarily through precipitation directly and through
the resulting runoff influents to tailing and settling ponds.
By their very nature, the water-treatment facilities are sub-
ject to precipitation inputs which, due to large areas, may
amount to substantial volumes of water. Runoff influxes are
often many times larger, however, and may be controlled to a
great extent by diversion ditches and (where appropriate)
conduits. Runoff diversion exists at many sites and is under
development at others.
Recycle of process water is currently practiced primarily where
it is necessary due to water shortage, or where it is economi-
cally advantageous because of high water costs. Recycle to
some degree is accomplished at many ore mills, either by
reclamation of water at the mill or by the return of decant
water to the mill from the tailing pond or secondary impound-
ments. Recycle is becoming, and will continue to become, a
more frequent practice. The benefits of recycle in pollution
abatement are manifold and frequently are economic as well as
environmental. By reducing the volume of discharge, recycle
not only reduces the gross pollutant load, but also allows
the employment of abatement practices which would be uneconomic
on the full waste stream. Further, by allowing concentrations
to increase, the chances for recovery of waste components to
offset treatment cost—or, even, achieve profitability—are
substantially improved. In addition, costs of pretreatment of
process water—and, in some instances, reagent use—may be
reduced.
Recycle of mill water almost always requires some treatment of
water prior to its reuse. In many instances, however, this may
VII-59
DRAFT
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DRAFT
entail only the removal of solids In a thickener or tailing
basin. This is the case for physical processing mills, where
chemical water quality is of minor importance, and the practice
of recycle is always technically feasible for such operations.
In flotation mills, chemical interactions play an important
part in recovery, and recycled water can, in some instances,
pose problems. The cause of these problems, manifested as
decreased recoveries or decreased product purity, varies and
is not, in general, well-known, being attributed at various
sices and times to circulating-reagent buildup, inorganic salts
in recycled water, or reagent decomposition products. Exper-
ience in arid locations, however, has shown that such problems
are rarely insurmountable. In general, plants practicing bulk
flotation on sulfide ores can achieve a high degree of recycle
of process waters with minimal difficulty or process modifica-
tion. Complex selective flotation schemes can pose more
difficulty, and a fair amount of work may be necessary to
achieve high recovery with extensive recycle in such a circuit.
Numerous examples where this has been achieved may be cited
(Reference 64 ). Problems of achieving successful recycle
operation in such a mill may be substantially alleviated by
the recycle of specific process streams within the mill, thus
minimizing reagent crossover and degradation. The flotation
of non-sulfide ores (such as scheelite) and various oxide
ores using fatty acids, etc., has been found to be quite
sensitive to input water quality. Attempts at water recycle
in such operations have posed severe problems, and successful
operation may require a high degree of treatment of recycle
water. In many cases, economic advantage may still exist over
treatment to levels which are acceptable for discharge, and
examples exist in current practice where little or no treat-
ment of recycle water has been required.
Technical limitations on recycle in ore leaching operations
center on inorganic salts. The deliberate solubilization of
ore components, most of which are not to be recovered, under
recycle operations can lead to rapid buildup of salt loads
incompatible with subsequent recovery steps (such as solvent
extraction or ion exchange). In addition, problems of corro-
sion or sealing and fouling may become unmanageable at some
points in the process. The use of scrubbers for air-pollution
control on roasting ovens provides another substantial source
of water where recycle is limited. At leaching mills, roasting
will be practiced to increase solubility of the product material.
Dusts and fumes from the roasting ovens may be expected to
VII-60
DRAFT
-------
DRAFT
contain appreciable quantities of soluble salts. The buildup
of salts in recycled scrubber water may lead to plugging of
spray nozzles, corrosion of equipment, and decreased removal
effectiveness as salts crystallizing out of evaporating scrubber
water add to particulate emissions.
Impoundment is a technique practiced at many mining and milling
operations in arid regions to reduce point discharges to, or
nearly to, zero. Its successful employment depends on favorable
climatic conditions (generally, less precipitation than evapo-
ration, although a slight excess may be balanced by process
losses and retention in tailings) and on availability of land
consistent with process-water requirements and seasonal or storm
precipitation influxes. In some instances where impoundment
is not practical on the full process stream, impoundment and
treatment of smaller, highly contaminated streams from specific
processes may afford significant advantages. Many operations
currently practicing Impoundment achieve a sufficiently favor-
able water balance to allow zero surface discharge only through
major seepage losses to ground water. The development of
restrictions on seepage may have substantial impact on the
practice of impoundment.
Electrodialysis
Electrodialysis is fundamentally similar to both reverse osmosis
and ultrafiltration to the extent that it employs semipermeable
membranes to allow separation of soluble cationic and anionic
impurities from water. An imposed electrical field is used
to provide a driving force for ion migration, in analogy to
either osmotic or external pressure in reverse-osmosis, dialytic,
or ultrafiltration systems.
Electrodialysis is generally employed in the treatment of waters
containing less than 5,000 to 10,000 mg/1 of dissolved solids
to achieve final levels of less than 500 mg/1 (Reference 39 ).
Applications have been reported in desalinization of seawater
involving feed water containing 38,000 mg/1 chloride and producing
a product water containing 500 mg/1 chloride (Reference 49 ).
To date, electrodialysls has not been employed in large-scale
operations within the mining/milling industry segments reviewed
and studied In this program. The potential for isolation and
recovery of byproduct or waste values exists but has not been
confirmed.
VII-61
DRAFT
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DRAFT
Freezing
This process depends on the formation of pure ice crystals
from the contaminated solution being treated. Results of
freezing experiments on acid mine-drainage samples (from a coal
mine) indicates that suspended solids act as condensation
nuclei and, if present, are entrained with the "pure" ice
obtained. Once solids have been removed, of course, the mine
drainage may still contain other contaminants.
Experimentally, agitation and slow freezing rates have allowed
reductions in dissolved materials in the range of 35 to 90
percent (Reference 40 ).
This process results in a concentrated stream, which still
requires treatment. It has a theoretical advantage over
distillation because only about one-sixth of the energy should
be required. Laboratory-scale experiments indicate it may be
a feasible treatment technique for mine and mill water treat-
ment, but it has not been fully tested.
Biological Treatment
The ability of various biota—both flora and fauna—to assimilate
soluble constituents from contacting waters is being documented
with increasing frequency. In general, these studies have
considered the undesirability of such assimilations, rather
than viewing them from the standpoint of potential water-
treatment options or systems. If trace or toxic constituents
can be metabolized, detoxified, or fixed by various organisms,
the periodic removal of organisms containing concentrates of
these materials may be a viable removal mechanism.
The use of this technique at one facility visited involves a
combination of sedimentation ponds and biological treatment
in the form of meanders. The meander system is an artificial
system designed to contain—and, thereby, control—excessive
algal growth and the associated heavy metals which are trapped
and assimilated by the algae (Reference 65 ). The algal
growth occurs naturally and was a problem associated with
the discharge prior to installation of the present system.
The system was designed as a series of broad, shallow, rapidly
flowing meanders, which increase the length of the treatment
section and encourage the growth of algae before discharge,
VII-62
DRAFT
-------
DRAFT
while simultaneously trapping any suspended heavy metals.
To prevent the algae and the associated heavy metals from
escaping the system, an additional final sedimentation pond
is placed at the end of the system.
The system can be effective if sufficient land is available
to allow the construction of an adequate meander system, and
if the climate is such that algae growth is not precluded during
parts of the year. These conditions effectively prevent wide-
spread application of this treatment technique.
VII-63
DRAFT
-------
DRAFT
EXEMPLARY TREATMENT OPERATIONS BY ORE CATEGORY
The manner in which ore mine and mill operators have approached
the design and construction of treatment and control facilities
varies from quite simple to somewhat sophisticated (utilizing
recycling, zero-discharge operations). To attain extensive
recycling or zero discharge, extensive process changes and/or
redesign have often been necessary. Performance of the many
vdired operations used In each ore category varies with the
operating characteristics of the facility, the ore mineralogy,
and other factors. Descriptions, by ore category, of the treat-
ment and control processes used in the ore mining and dressing
industry and the consequent treatment levels attained are
Included here to provide a more complete explanation and examina-
tion of the control and treatment technology currently in use.
Iron Ore
This discussion includes examples of mines that have discharges
(Subcategory I), mills which employ physical or chemical
benefIciation (Subcategory II), and mills using magnetic- and
physical-separation methods to extract iron from Mesabi Range
ifon formations.
Mining Operations. Mine 1105 is an open-pit operation that
accumulates water. Water is pumped directly from the pit to a
settling pond of sufficient volume to remove suspended solids
prior to discharge. No chemical coagulants are used, because
the suspended-solid concentration generally is less than 10
mg/1. Because this operation produces low levels of dissolved
components, dissolved-solid treatment is unnecessary. Suspended-
solid concentrations after treatment have been observed to
remain low, but historical data obtained during periods of high
rainfall and high pumping rates are lacking.
Table VII-5 is a compilation of data measured in this study and
by the operators. It can be observed that many of the parameters
measured appear to increase in the effluent stream after treatment,
Measurements made during this study were confirmed by duplicate
Industry sample analysis. Conditions existing at the mine
settling pond should be noted, however. At the mine discharge,
an extremely low flow was encountered, and only intermittent
pumping of the mine was being employed. At the settling-pond
discharge, however, flow conditions were adequate for sampling.
VII-64
DRAFT
-------
DRAFT
TABLE VI1-5. CHEMICAL CHARACTERISTICS OF SETTLING-POND DISCHARGE AT
MINE 1105
PARAMETER
PH
TSS
TDS
COD
Oil and Grease
Total Fe
Dissolved Fe
Mn
Sulfate
AVERAGE
MINE-DISCHARGE
CONCENTRATION (mQ/Jt)
This Study
7.4*
10
225
9.7
<1
<0.02
< 0.02
0.04
24
Industry
7.9*
6
243
4.5
< 5
—
^0.1
^ 0.1
—
AVERAGE
SETTLING-POND
DISCHARGE
CONCENTRATION (mg/£)
This Study
7.4*
25
283
13.7
<1
0.1
<0.02
<0.02
35
Industry
8.0*
as
291
15
<5
-
^0.1
^ 0.1
—
AVERAGE
SETTLING-POND
DISCHARGE
CONCENTRATION
(mg/£)
8.0*
3.4
-
-
(<10)
—
-
-
—
Value in pH units
VII-65
DRAFT
-------
DRAFT
Historical data obtained at this location for nine months during
1974 show that a range of 1 to 9 (average of 3.4) mg/1 of
TSS was encountered after settling.
Mills Employing Physical and/or Chemical Separations. Iron-
beneflciation plant 1109 uses magnetic separation, coupled
with a froth-flotation sequence that removes undesired silica
in the iron concentrate. The processing circuit uses 587
cubic meters (155,000 gallons) of water per minute, with a
recycle rate of 568 cubic meters (150,000 gallons) per minute.
Thickeners, located adjacent to the concentrator, are used to
reclaim water close to the site of reuse so as to minimize
pumping requirements. Superfloe 16, an anionic polyacrylamide,
is added to the thickeners at a rate of 2.5 grams per metric
ton (0.0049 pound per short ton) of mill feed to aid in clarifi-
cation of the water in the thickeners. The thickener underflow
is pumped to a 850-hectare (2,100-acre) tailing basin for the
sedimentation of the solids. Mine water is also pumped to the
basin. The effluent leaves the basin after sufficient retention
and flows into a creek at an average rate of 22.3 cubic meters
(5,900,000 gallons) per day. Chemical analysis of the waste-
water to the tailing pond (mine and mill water) in comparison
to the effluent water quality and waste loading is given in
Table VII-6.
Mills Employing Magnetic and Physical Separations. Mill 1105
is located in the Mesabl Range of Minnesota and is processing
ore of the Biwabik formation. Crude magnetic taconite is
milled to produce a fine magnetite. The mill's water system
is a closed loop having no point-source discharges to the
environment. The plant processes use 20.4 cubic meters (54,000
gallons) per minute, with 189 cubic meters (50,000 gallons)
per minute returned from the tailing-thickener overflow and
15.1 cubic meters (4,000 gallons) per minute returned from the
tailing pond or basin. The tailing thickener accumulates all
the milling-process wastewater containing the tailings. A
nontoxic polyacrylamide flocculant (SuperFloc 16) is added to
the thickener to assist the settling out of solids. Tailing-
thickener underflow is pumped to a tailing basin of 470 hectares
(1,160 acres), where the solids are settled and the clear water
is recycled back into the plant water-use system. A simplified
water-use sequence is shown in Figure VII-6.
VII-66
DRAFT
-------
DRAFT
TABLE VII-6. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTELOADING AT MINE/MILL 1109
PARAMETER
PH
TSS
TDS
COD
Total Ft
Oinolnd Ft
Mn
Sultal*
Alkalinity
MINE EFFLUENT
CONCENTRATION
(main
8.3"
12
308
27.8
0.30
0.02
0.66
37
181
MILL EFFLUENT
CONCENTRATION
(mo/£>
8.6"
(66%)
360
138
0.04
0.04
-
20.7
238
WASTE 1
PER UNIT F
kg/rmtrle ton
1.346
0.88
0.033
0.0001
0.0001
-
0.06
0.68
-OAD
RODUCT
Ib/ihort ton
2.890
178
0.066
0.0002
0.0002
-
010
1 16
FINAL DISCHARGE
CONCENTRATION
Img/ 1)
8J"
10
222
18X1
0.76
OM
<0.02
3£
120
WASTE I
PER UNIT P
kg/mttrle ton
0.02
0.48
0.039
0.0018
00010
< 0.00004
00078
0.26
.OAD
RODUCT
Ib/thort ton
0.04
0.86
0.078
0.0032
0.0020
0.00008
00162
0.52
HISTORICAL
CONCENTRATION*
(mg/t>
7.7"
34
_
_
_
0.60
0.06
_
-
Anrige of nlno raluii (August through October 1974)
•
Vilu* In pH unln.
VII-67
DRAFT
-------
DRAFT
Figure VI1-6. MILL 1105 WATER-USE SYSTEM (ZERO DISCHARGE)
WATER
1 '
^._.., ..... p
PROCESS
PRODUCT
\ >
THICKENING
1
OVERFLOW UNDERFLOW
t
FILTRATION
1 1
CAKE FILTRATE
1 1
\
TO FINAL
PROCESSING
ROCESS PLANT i
1
PROCESS
TAILING
1
THICKENING
1
UNDERFLOW OVERFLOW
/""SEDIMENTATIONS
\_^ BASIN^./
SETTLED CLARIFIED
SOLIDS EFFLUENT
|
TO WASTE
VII-68
DRAFT
-------
DRAFT
Copper Ores
The discussion that follows describes treatment and control
technology in current use in the five subcategories of the
copper-ore mining and dressing Industry.
Mining Operations. Mine water generated from natural drain-
age is reused in mining, leaching, and milling operations wherever
possible in the copper mining industry. Because of an excess
of precipitation in certain areas of the country, a location
which is not proximate to a milling facility, or an inability
to reuse the entire amount of mine wastewater at a particular
mill, a discharge may result. The amounts of precipitation
and evaporation thus have an Important influence on the
presence or absence of mine-water discharge.
To avoid discharge, mine effluent .may be reused in dump, heap,
or in-situ leaching as makeup water. As a leach solution, it
is acidified (if necessary), percolated through the waste dump,
sent through an iron-precipitation facility, and recycled to
the dump (Figure VII-7).
Large quantities of water are usually needed in the copper
flotation process. Mine-water effluent is used at many
facilities as mill process makeup water. The mine water may
pass through the process first, or it may be conveyed to the
tailing pond, from which it is used for mill flotation with
recycled process water (Figure VII-8). The practice of com-
bining mine water with mill water can create water-balance
difficulties unless the mill circuit is capable of handling
the water volumes generated without a discharge resulting.
The discharge of mine water into a mill process system which
creates an excess water balance and subsequent discharge may
have a detrimental effect on the mine water because of contam-
ination by mill flotation reagents and residual wastes.
Acid mine water is encountered in the copper mining
industry, and methods of neutralization usually employed
Include the addition of lime and limestone.
Lime precipitation is also often used to enable the removal of
heavy metals from wastewater by precipitation as hydroxides.
Tables VII-7 and VII-8 show examples of the use of lime
precipitation for treatment of mine water at two locations of
mine 2120. The use of this treatment technology yields reduc-
VII-69
DRAFT
-------
DRAFT
Figure VII-7. CONTROL OF EFFLUENT BY REUSE OF MINE WATER IN LEACHING
(MINE 2122)
EVAPORATION
AND SEEPAGE
MINE
— EFFLUENT
3270 m3/day
(864,000 gpd)
STORAGE
RESERVOIR
EVAPORATION
AND SEEPAGE
DUMP LEACH
BED
1
PREGNANT
SOLUTION
RECYCLED
BARREN
SOLUTION
IRON
PRECIPITATION
PLANT
I
CEMENT
COPPER
TO
STOCKPILE
VII-70
DRAFT
-------
DRAFT
Figure VI1-8. CONTROL OF MINE-WATER EFFLUENT BY REUSE IN THE
CONCENTRATOR (MINE/MILL 2119)
MINE
MILL/
CONCENTRATOR
37,100 m3/day
(9,792,000
gpd)
TAILING
THICKENERS
RECYCLED
'OVERFLOW
RECYCLED
POND
WATER
(NO DISCHARGE)
VII-71
DRAFT
-------
DRAFT
TABLE VI1-7. CONCENTRATION OF PARAMETERS PRESENT IN RAW WASTEWATER
AND EFFLUENT FOLLOWING LIME PRECIPITATION AT MINE 2120B
PARAMETER
PH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
CONCENTRATION (mg/£)
RAW WASTEWATER
6.1 •
2.200
40
<1
3.2
<10
0.04
5.3
0.1
< 0.07
31.25
<0.5
6.0
26.5
0.175
0.13
< 0.5
1.55
0.0005
< 0.1
TREATED WASTEWATER
12.7*
3.000
34
<1
1.2
<10
< 0.01
0.05
<0.04
<0.07
0.11
<0.5
<0.1
0.04
< 0.005
<0.05
<0.5
0.85
0.0002
<0.1
EFFICIENCY OF TREATMENT
IN REMOVAL OF POLLUTANTS
(% REMOVAL)
INCREASED
INCREASED
15%
-
63%
-
>75%
99%
>60%
-
99.7%
-
>98%
99.9%
>71%
62%
-
45%
60%
^™
•Value in pH units
VII-72
DRAFT
-------
DRAFT
TABLE VI1-8. CONCENTRATION OF PARAMETERS PRESENT IN RAW WASTEWATER
AND EFFLUENT FOLLOWING LIME PRECIPITATION AT MINE 2120C
PARAMETER
PH
TDS
TSS
Oil & Grease
TOC
COD
S04
Cu
Co
Ai
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
CONCENTRATION (mg/«)
RAW WASTEWATER*
4.7"
450
35
17
2.3
<10
300
6.2
0.06
<0.07
6.2
<0.5
8.6
1.42
0.03
<0.05
<0.05
0.09
0.0005
<0.1
TREATED WASTEWATER*
7.8"
—
3
—
—
.-
220
0.25
—
0.004
0.45
—
0.5
—
0.01
—
—
—
0.0005
0.01
EFFICIENCY OF TREATMENT
IN REMOVAL OF POLLUTANTS
(% REMOVAL)
INCREASES
—
91%
—
—
-
27%
96%
-
—
93%
—
94%
—
67%
—
—
—
-
—
Data obtained from sampling and analysis.
Data obtained from plant monitoring records.
Value in pH units.
VII-73
DRAFT
-------
DRAFT
tions approaching 100 percent for several heavy metals of
interest.
Various techniques are employed to augment the use of lime
neutralization. Among these are secondary settling ponds,
clarifier tanks, or the addition of flocculating agents (such
as polyelectrolytes) to enhance removal of solids and sludge
before discharge. Often, readjustment of the pH is necessary
after lime treatment. This can be accomplished by addition
of sulfuric acid or by recarbonation. The use of sulfide
precipitation may be necessary in some instances for further
removal of metals such as cadmium and mercury.
Mine Employing Hydrometallurgical Process. Acid solutions
employed in dump, heap, and in-situ leaching are recycled in
this subcategory of the copper industry, allowing the recovery
of copper in the iron precipitation plant. Water is added to
replace losses due to evaporation and seepage. Acid is added
to control pH. Table VII-9 lists the operations surveyed and
their control of acid solutions. Only one operation surveyed
discharges a small amount of "bleed water" to surface waters.
Control of seepage and collection of acid-leach solution are
sometimes aided by the construction of specially prepared
surfaces, upon which heaped ores are placed for leaching.
These surfaces may be constructed of asphalt, concrete, or
plastic.
One facility currently bleeds the acid-leach solution and
treats the bleed by neutralization and precipitation with
alkaline (limed) tailings from the mill. The treated water
flows into the tailing pond for settling and is subsequently
recycled with the decant water to the mill.
Treatment of the leach solutions used in this subcategory is
sometimes necessary for control of dissolved solids, which
build up during recycling. Holding ponds are constructed to
retain leach solutions for a sufficient time to allow the
iron salts to precipitate from solution and settle, before
the solution is recycled to leach beds. In conjunction
with, or in place of holding ponds, pH control aids in pre-
venting iron salts from precipitating in pipes or In the
leach dump.
VII-74
DRAFT
-------
DRAFT
TABLE VI1-9. DUMP, HEAP, AND IN-SITU LEACH-SOLUTION CONTROL
AND TREATMENT PRACTICE (1973)
PLANT
2101
2102
2103
2110
2116
2118
2123
2107
2108
2122
2124
2125
2104
2120
CONTROL
Zero discharge
Zero discharge
Zero discharge
99.4% recycle
98.7% recycle
TREATMENT
Recycle without treatment
20% to evaporation ponds
•
All effluent circulated through
holding ponds or reservoirs
None
Bleed is limed and settled in
tailing pond
DISCHARGE
None
None
None
654m3/day(avg)*
2551 m3/day (avg)"
to tailing pond (not
discharged)
•Inadequate pumps. Operation required to attain zero discharge by State Regulations in 1977.
••The treated bleed is recycled to the mill with the decant.
VII-75
DRAFT
-------
DRAFT
Evaporation ponds are also employed to accomplish zero dis-
charge of acid-leach bleed solutions.
Mill Employing Vat Leaching for Extraction. Zero discharge
has been reached by all facilities studied (Table VII-10).
Makeup water is required to replace evaporative losses and
the moisture which remains in the discarded, leached ores.
Complete recycling of barren leach and wash solutions is
usually practiced. However, one facility presently reuses
its spent vat-leach solution in a smelter process to achieve
zero discharge.
Mill Employing Concentration by_ Froth Flotation. Mills
employing froth flotation constitute two subcategories of the
copper-ore mining and dressing Industry. The two subcategories
are divided on the basis of climatic conditions as: (1) mills
located in areas where net evaporation is less than 76.2 cm
(30 in.); and (2) mills located in areas where net evaporation
equals or exceeds 76.2 cm (30 in.). All facilities currently in
operation in subcategory (2) discharge no wastewater effluent.
Process water from froth flotation contains large amounts of
suspended solids, which are normally directed to a large lagoon
to effect settling of these solids. Surface runoff, such as
that resulting from snow melt, heavy-rainfall events, streams,
and drainage, should be conveyed around the tailing pond, thus
preventing runoff water from contacting process effluents.
In this manner, the volume of water which must be treated or
impounded is reduced.
Mill tailing-pond water may be decanted after sufficient
retention time. One alternative to discharge, and an aid to
reducing the amount of effluent, is to reuse the water in
other facilities as either makeup water or full process water.
Usually, some treatment is required for reuse of this decanted
water. Figure VII-9 illustrates the control of effluent by
reuse, as practiced at mill 2124.
The volume of water to be treated in flotation mills can be
effectively reduced, and the quality of the discharge often
VII-76
DRAFT
-------
DRAFT
TABLE VI1-10. SOLUTION-CONTROL PRACTICE IN VAT LEACHING OF COPPER ORE
MILL
2102
2116
2124
2126
CONTROL
100% recycle
100% recycle
100% recycle
Zero discharge
RECYCLE TREATMENT
None
None
None
Spent acid sent to acid plant for
• reuse
VII-77
DRAFT
-------
DRAFT
Figure VI1-9. CONTROL OF EFFLUENT THROUGH REUSE OF MILL FLOTATION-
PROCESS WATER IN OTHER FACILITIES (MINE/MILL 2124)
VII-78
DRAFT
-------
DRAFT
substantially improved, by the separation of mine water, sewage,
smelter drainage, refinery wastes, and leach bleed solution
from the tailing-pond circuit. It has been observed that
separation of mine water, with subsequent treatment and dis-
charge of the mine water only, can allow mill tailing decant
water to be recycled fully. As a result, lower total pollu-
tant loads may be discharged to the environment. Using mine/mill
2121 as an example, Figure VII-10 was constructed to illustrate
current practice, as well as alternative future practice which
would result in a reduction of the waste loads discharged.
Separation of mine water and other wastes from contact with
mill process water is suggested in all cases where pollutant
load and water volume are factors. Not only do these waste
waters contribute to the pollutants present in the tailing-
pond water, but they may dilute the water to be treated or cause
excess water-volume conditions to result which cannot be
handled by recycling.
If sewage plant overflow contributes to the tailing-pond water
volume to the extent that it cannot be accommodated in recycling,
this water should be properly treated and handled separately.
Smelter and refinery wastes often contribute a heavy load of
dissolved metals to tailing ponds. These wastes can affect
the quality of the decant water, as well as effluent volumes.
It may be necessary to handle wastes from these sources separately,
and/or as recommended under the appropriate conditions for the
Effluent Limitation Guidelines for the Copper Smelting and
Refining Industry.
The most efficient control of the volume and pollutant dis-
charge of mill flotation-process water is to recycle the
excess water which would overflow from the tailing-pond decant
area. Of the 27 major copper mills surveyed, 24 are known to
be recycling all or a portion of their process water. The
impetus for recycling has often been the lack of an adequate
water supply. However, the feasibility of recycling process
water appears to have been considered at all facilities.
Through the use of diversion ditching, evaporation (when
available), reservoirs, and separation of other process water,
the volume of water to be recycled can be adjusted to allow
reuse. Treatment of the recycled water is usually required
and may include secondary settling, phosphate or lime addition
(for softening), pH adjustment, or aeration.
VII-79
DRAFT
-------
DRAFT
Figure VII-10. REDUCTION IN WASTE POLLUTANT LOAD IN DISCHARGE BY SEPARATION
OF MINEWATER FROM TAILING POND FOR SEPARATE TREATMENT
(MILL 2121)
CURRENT
OTHER
WASTES
MILL
PROCESS
WATER
(LIMED)
1 MINE 1
1 '
EFFLUENT
1
- 1 1
^ X^~TAILIr\H5X
^v.^ rotto^J
IT
1 EFFLUENT \(t
1 1^-
TOTAL WASTE LOAD DISCHARGED
Par 24 noun in kg/day (Ib/day)
Flaw
pH
TSS
Oil
Cu
A.
Zn
Fa
Cd
Ni
Hg
Pb
and Greata
LIME I
x
)
AT®
102.000 m3/day (27.000.000 gpd)
8.4'
020 (1,384)
41S (913)
27 (59.4)
<8 (<17.6)
6.2 (11.41
103 (227)
< 2 « 4 4)
<5.2 K11.4)
<001 (<0022)
C103 (<227)
ALTERNATIVE
1 1
1 »H MILL
-Tj^j
PROCESS
WATER
X POND_ /
RECYCLE T
1
1 MINE 1
1 '
T
LIME
TREATMENT
t
SETTLING |
{
DISCHARGE
K§)
ESTIMATED TOTAL WASTE LOAD DISCHARGED, USING LIME
PRECIPITATION/ AT ®
Par 24 noun In kg/day (Ib/day)
Flaw
pH
TSS
Oil and Gresu
Cu
At
Zn
Fa
Cd
Ni
Hg
Pb
Raw (No Trutmant)
3,800 m3/day
(1.000,000 gpd)
7.4'
267 (587)
<4 K8.8)
4(8.8)
<03 «066)
108 (23.8)
<04 «0.88)
<007 1< 0.1 541
<0.2 «044)
< 0.0005 « 0.001 10)
<0.4 «088)
After Treatment
3.800 m3/day
(1,000,000 gpd)
12.7'
129 (284)
<4 «8.8)
0.2 (0.441
<0.3 «066)
0.4 (038)
<04 (<0.88)
<002
« 0.044)
<0.2 I < 0.44)
0.0004 (040088)
<0.4 (
<0£8)
Value in pH units.
VII-80
DRAFT
-------
DRAFT
The majority of copper mills currently operating recycle their
mill process water. Of the remaining facilities that currently
discharge, half are recycling at least 35 percent of their
process water. Treatment of discharged water consists of
settling alkaline wastewater in a tailing pond. A variety of
treatment approaches are currently used in this subcategory,
including:
(1) Settling Only
(2) Lime Precipitation and Settling
(3) Lime Precipitation, Settling, Use of Polyelectrolytes,
and Secondary Settling
One operation is currently building a treatment facility which
will include lime precipitation, settling, and aeration.
Table VII-11 shows the reduction of pollutant concentrations
attained in six mills under different conditions of recycling,
lime addition, and settling. A wide variation in practice
is used to obtain varying degrees of concentration for waste
constituents present in treated wastewater.
Additional treatment of wastewater by polyelectrolyte addition,
to reduce suspended solids in tailing-pond discharge, is also
practiced at one mill. Secondary settling ponds are used
to settle the treated solids prior to discharge.
The effectiveness of the use of coagulants (polymers) is
demonstrated in Table VII-12. These data, obtained from a
pilot operation, indicate effective reductions of copper,
iron, and cobalt, with substantial reductions of aluminum
and manganese.
Recycling of process water from the tailing pond has not been
difficult for most copper mills surveyed employing this tech-
nique. However, treatment of the pond water has been necessary
for selected problems encountered. Potential problem areas
present at these operations include buildup of scale deposits,
pH changes in the tailing pond or in makeup water, and presence
of flotation reagents in the recycled water. Effective
methods of treatment to alleviate these conditions are phos-
phate treatment (softening) for scale control, adjustment of
pH by liming, and the use of aeration or secondary settling
ponds to assist in degradation of flotation reagents.
VII-81
DRAFT
-------
DRAFT
TABLE VII-11. REDUCTION OF POLLUTANTS IN CONCENTRATOR TAILS
BY SETTLING AT VARIOUS pH LEVELS
PARAMETER
PH
TSS
Al
Ai
Cd
Cr
Cu
Fa
Pb
Mn
Ho
Ni
Se
Zn
Sb
Co
Mo
Comments:
PARAMETER
pH
TSS
Al
At
Cd
Cr
Cu
Fa
Pb
Mn
Hg
Ni
SB
Zn
Sb
Co
Mo
Conwitsfits*
CONCENTRATION (mg/ 1)
MILL 21 19
BEFORE
SETTLING
11.6-
705,000
< 1.0
< 0.07
< 0.05
< 0.05
0.1 B
0.8
< 0.5
< 0.05
0.0002
< 0.1
0.02
< 0.05
< 0.2
< 0.05
< 0.2
AFTER
SETTLING
7.7*
10
< IX)
< 0.07
< 0.05
< 0.05
0.05
0X18
< 0*5
0.3
< 0.0001
<0.1
0.08
< 0.05
< 0.2
< 0.05
<0.2
lima added after mill
water recycled
MILL 2120
BEFORE
SETTLING
11.1"
282.000
1.6
0.6
< 0.02
< 0X15
0.8
52
< 0.1
OX»7
0.0008
< 0.05
—
0.1
< 0.5
< 0.04
< 0.5
AFTER
SETTLING
9j6*
8
<0£
< OX>7
< OXW6
< 0X15
048
< 0.1
<0.1
OX>3
0X1011
< 0X15
0X14
< 0X15
< OS
< 0.04
< OS
lime added after mill
water recycled
MILL 2121
BEFORE
SETTLING
10J»
166X»0
lOXi
< OJ07
< 0X12
< 0.06
3.5
18.5
02
0.35
0X1098
< 0X15
0X12
0.9
< OX>
< OXM
< OS
AFTER
SETTLING
8.4*
6
<0.5
< 0.07
< 0.02
< OXS
0.3
<0.1
<0.1
OXM
< 0X1001
<0.04
< 1.0
< 0.07
< 0.05
< 0.05
0.08
< 0.1
2.8
O.OB
0.0002
< 0.1
0.02
< 0.05
<1.0
0X18
< 0.2
AFTER
SETTLING
8.4»
16
< 1.0
< 0.07
< 0.006
< 0.05
0.12
0.93
2X1
0.06
< 0.0001
<0.1
0X13
< 0.05
< 1.0
0.12
< 0.2
no lime addition after mill
water partially recycled
MILL 2123
BEFORE
SETTLING
13«
335,000
1.0
< 0.07
< 0X13
< 0X15
0X1
0.2
< 0.1
< OXtt
OXW2
< 0.05
0.07
< 0.05
< 0.5
< 0X16
< 0.5
AFTER
SETTLING
9£*
17
0.4
< 0X17
< 0X13
< 0X15
1.7
< 0.1
< 0.1
< 0X16
OX»2
< 0.05
0X108
< 0X15
< 05
< 0.06
-
no lime addition after mill
water recycled
MILL 2124
BEFORE
SETTLING
10*
640,000
<0.5
< OX»7
0X15
3.6
91 2.5
1.882
0.4
31
OXM06
2X1
< 0X103
5.6
< OS
1.7
29.3
AFTER
SETTLING
8.4*
14
< 0X17
< 0X8
OJOS
< ox»
903
< 0.1
< OX>6
OXW9
< 0.05
0.02
< 0.05
<0.06
< 0.5
no lima addition after mill
water recycled
•Value in pH unto
VII-82
DRAFT
-------
DRAFT
TABLE VII-12. EFFICIENCY OF COAGULATION TREATMENT TO REDUCE
POLLUTANT LOADS IN COMBINED WASTE (INCLUDING
MILL WASTE) PRIOR TO DISCHARGE (PILOT PLANT)
POLLUTANT
PARAMETER
Flow
pH
TDS
TSS
Al
Ai
Cd
Cu
Fo
Pb
Mn
Hg
Ni
Co
Zn
WASTE LOAD IN INFLUENT TO PROCESS
kg/1000 metric toni
76.134 m3/day
7.6«
3.600
10
2.3
0.2
<0.05
B.8
120
3.3
0.4
a 0001
< ai
as
<0.05
Ib/IOOOgal
ie.860,400gpd
7.6»
6
0.02
0.004
0.0003
< 000009
0.02
0.21
0.008
0.0007
0.0000001
< 0.0002
0.02
< 0.00009
WASTE LOAD IN EFFLUENT TO DISCHARGE
kg/1000 metric tons
76.198 m3/dev
9.0*
3.900
14
< 1
0.9
< 0.05
0.9
0.7
2.8
0.1
0.0003
<0.1
0.9
< 0.002
0.002
< 0.00009
0.002
0.001
0.005
0.0002
0.0000005
< 0.0002
0002
< 0.00009
K EFFICIENCY
IN REMOVAL
-
-
-
-
>57%
-
-
90%
>99%
15%
71%
-
-
90%
-
•Value in pH units
VII-83
DRAFT
-------
DRAFT
Lead and Zinc Ores
A discussion of the treatment and control technologies currently
employed in the lead and zinc ore mining and dressing industry
is included in this section. Three subcategories are repre-
sented: Mines with alkaline drainage not exhibiting solubili-
zatlon of waste constituents, mines with acid or alkaline drainage
exhibiting extensive solubilization of metals, and lead and/or
zinc mills.
Mines With Alkaline Drainage Not Exhibiting Solubilization of
Metals. The operations represented by this subcategory gen-
erally employ treatment by impoundment in tailing or sedimen-
tation ponds. Mine 3105 (producing lead/zinc/copper concentrates)
is located in Missouri. The mine recovers galena
(FbS) , sphalerite (ZnS), and chalcopyrite (CuFeS). Production
began in 1973, and the operation was expected to produce 997,700
metric tons (1,100,000 short tons) of ore in 1974.
The water pumped from this mine is treated by sedimentation
in an 11.7-hectare (29-acre) pond. The average mine drainage
flow rate is 8,300 cubic meters (2,190,000 gallons) per day.
The effluent from this basin flows to a nearby surface stream.
The chemical characteristics of the wastewater before and
after treatment are presented in Table VII-13, together with
data for nine months of 1974. The treatment sequence is as
follows: mine pumping, followed by clarification basin, followed
by discharge (8,300 cubic meters (2,190,000 gallons) per day).
Relatively simple treatment employed for mine waters exhibiting
chemical characteristics similar to mine 3105 can result in
attainment of low discharge levels for most constituents.
Reduction of parameters such as total dissolved solids, oil
and grease, chloride, sulfate, lead, and zinc—as well as excel-
lent reduction of total suspended-solid concentrations—is obtained
by this treatment method.
Mine Drainage (Acid or Alkaline) Exhibiting Extensive Solubili-
zation £f_ Metals. The characteristics of wastewater from
mines in this subcategory are such that treatment must be
applied to prevent the discharge of soluble metals, as well
as suspended solids. The treatment practice, as currently
VII-84
DRAFT
-------
DRAFT
TABLE VII-13. CHEMICAL COMPOSITIONS OF RAW AND TREATED MINEWATERS
FROM MINE 3105 (HISTORICAL DATA PRESENTED FOR COMPARISON)
PARAMETER
pH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
Ammonia
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£)
RAW MINE
DRAINAGE*
7.4»*
196.0
330.4
138
326
<10
< 1.0
29.0
0.030
<0.05
0.0001
0.3
0.03
<0.02
< 0.002
<0.02
<0.02
<0.02
63.5
57
1.2
DISCHARGE*
8.1 **
16^0
173.2
< 2
204
<10
3.0
17.0
0.032
<0.05
< 0.0001
0.1
<0.02
<0.02
0.005
<0.02
0.35
0.11
45.5
44.5
1.0
DISCHARGE (HISTORICAL)*
AVERAGE
7.8**
—
-
3.4
-
—
-
1.9
-
-
-
0.050
0.032
< 0.005
< 0.005
-
-
0.086
-
-
-
RANGE
7.4**to8.1**
—
-
<1 to 9
-
-
-
< 1 to 5
-
-
-
0.011 to 0.01 2
0.008 to 0.11
<0.050 to 0.070
(< 0.005)
—
-
0.033 to 0.21
-
-
-
•Analysis of single 4-hour composite sample
tMonthly analysis over January 1974 through September 1974
••Value in pH units
VII-85
DRAFT
-------
DRAFT
employed, involves chemical (often, lime) precipitation and
sedimentation.
Mine wastewaters in this subcategory are often treated by
discharge into a pond or basin in which the pH is controlled.
An approach often used is to discharge the mine wastewater
into a mill tailing pond, where wastewater is treated at a
pH range which causes the precipitation of the heavy metals
as insoluble hydroxides. The presence of residual solids from
the milling process is thought to provide nucleation sites
for the precipitation of the hydroxides. In cases where
ferrous iron is present, it is desirable to cause the oxidation
to the ferric form and, thus, to avoid the potential for acid
formation by processes similar to the reactions forming acid
mine drainage. Vigorous aeration of the wastewater can accom-
plish oxidation, usually after addition of the pH-adjusting
agent.
The treatment process described is similar to the type of pH
control, and subsequent physical treatment, usually associated
with froth-flotation recovery of sulfides of lead, zinc, and
copper (which is followed by sedimentation of the tailings).
The milling process itself is, therefore, an analog for a
process of treating mine wastes in this subcategory.
Mine 3107 is an underground lead/zinc mine located in Idaho.
Galena and sphalerite are mined, with approximately 544,200
metric tons (600,000 short tons) of ore mined per year. The
mine has been in operation most of this century.
Mine water pumped from lower levels of the mine, as well as
water from upper levels (which flows by gravity), exits the
mine tunnel and is piped to a central impoundment, 48.5 hectares
(120 acres) in area. The average mine flow is 16,500 cubic
meters (4,360,000 gallons) per day. Waste streams, including
the tailings from the concentrator, a smelter, and an electrolytic
zinc plant, also flow to the central impoundment area. The
overflow from this impoundment area, 29,000 cubic meters
(7,700,000 gallons) per day, is treated in a high-density,
sludge-type chemical-precipitation plant. The characteristics
of the raw mine waste, the overflow from the central impound-
ment area, and the final effluent from the treatment process
are presented in Table VII-14.
VII-86
DRAFT
-------
DRAFT
TABLE VII-14. CHEMICAL COMPOSITIONS OF RAW AND TREATED WASTEWATERS
FROM MINE 3107 (HISTORICAL DATA PRESENTED FOR COMPARISON)
PARAMETER
pH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
Ammonia
Mercury
Lead
Zinc
Copper
Cadmium
Chromium
Manganese
Iron
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£)
RAW
MINE WATER
3.2'
14.6
671
<2
1.722
47.6
2.3
3.0
<0.02
1.8
0.0001
0.3
38.0
0.04
0.055
0.17
57.2
2.5
750
<001
0.063
OVERFLOW FROM
CENTRAL POND
2.0t
0.0
2.356
<2
2.254
39.7
4.3
<1
0.08
1.6
0.0468
3.1
180.0
0.52
140
0.67
41.0
59.0
1.862
1.2
19
TREATED
EFFLUENT
8.5*
3.2
1.242
<2
2.030
43.6
4.0
17
<0.02
0.80
0.0007
<0.1
5.1
0.04
0.048
0.50
0.32
0.85
1.744
15
21
HISTORICAL DATA
AVERAGE
7.4*
-
-
-
-
-
-
-
-
-
0.002
0.093
1.43
0.020
0.044
_
-
-
-
-
-
RANGE
6.9* to 7.6*
-
-
-
-
-
-
-
-
-
< 0.001 to 0.005
0.057 to 0.153
0.79 to 2.08
0.010 to 0.043
0.032 to 0.058
_
-
_
-
-
-
Average foi months includes 10-24 hour composite samples.
Value in pH units.
VII-87
DRAFT
-------
DRAFT
The treatment process is shown schematically in Figure VII-11.
Provision has been made for pumping the recovered sludge back
to the mill, should recovery of metal values prove practical.
At present, the sludge is disposed of at a solid-waste disposal
site.
Mine 3101 is an underground mine, located in Maine. The
mine recovers sphalerite and the byproducts chalcopyrite,
galena, and pyrite which are present in the formation.
The mine began production 1972 and produced 208,610 metric
tons (230,000 short tons) of ore in 1973.
The water pumped from the mine, 950 cubic meters (250,000
gallons) per day, is treated by mixing it with mill tailing
discharge, plus additional lime as required for pH control,
in a reservoir with a capacity of 37.85 cubic meters (10,000
gallons). The combined waste is then pumped to a 25-hectare
(62-acre) tailing pond. The discharge from the tailing pond
is sent to an auxiliary pond. The combined retention time in
the two ponds is 35 days at maximum flow. Water is recycled
for the process from the auxiliary pond, and the excess is
discharged. The chemical characteristics of the mine water
and the final discharge, treated in the above manner, are given
in Table VII-15.
Lead and/or Zinc Mills. As discussed in Section V, the
wastewater from lead/zinc flotation mills differs from mine water
in that a number of reagents are added to effect the separation
of the desired mineral or minerals from the host rock. These
waste streams also contain finely ground rock, as well as
minerals, as a result of grinding to allow liberation of the
desired minerals during the froth-flotation process.
The most common treatment method in use in the lead/zinc-
milling industry is the tailing or sedimentation pond. Often
considered a simple method of treatment, properly designed
tailing ponds perform a number of important functions simultan-
eously. Some of these functions include removal of tailing
solids by sedimentation, formation of metal precipitates,
long-term retention of settled tailings and precipitates,
stabilization of oxidizable constituents, and balancing of
influent-water quality and quantity of flow.
VII-88
DRAFT
-------
DRAFT
Figure VII-11. SCHEMATIC DIAGRAM OF TREATMENT FACILITIES AT MINE 3107
22.7 m /mm
(6,000 gpm) max
49-hectare
(120-acre)
CENTRAL
POND
56.8-m3 (15,000-gal)
RECYCLE MIX TANK
(5-minute retention time)
RECYCLE
SLUDGE
1,442-m
(381,000-gal)
AERATION BASIN
(41-minute retention time)
333-m
FLOCCULATION
16,653-m
(4,400,000-gal)
THICKENER
(8-hour retention time)
SLUDGE
BLEED
MILL
RESERVOIR
VII-89
DRAFT
-------
DRAFT
TABLE VIMS. CHEMICAL COMPOSITIONS OF RAW AND
TREATED MINE WATERS FROM MINE 3101
PARAMETER
pH
TSS
TDS
COD
Pb
Zn
Cu
Cd
Cr
Mn
Fe
CONCENTRATION (mg/£>*
RAW MINE
WATER
6.9t
-
—
-
0.035
2.608
0.012
0.004
< 0.010
0.996
0.359
TREATED
DISCHARGE
8.7*
7.2
595
25
< 0.024
0.096
0.016
< 0.002
< 0.010
0.055
0.303
•Average for year of 1974 as reported for NPDES permit
*Value in pH units
VII-90
DRAFT
-------
DRAFT
In the lead/zinc-ore milling industry, a biological treatment
method, used in conjunction with stream meanders, was observed
at one location. This treatment method has been described in
the previous discussion in this section.
The ability to recycle the water in lead/zinc flotation mills
is affected by the buildup of complex chemical compounds
(which may hinder extraction metallurgy) and sulfates (which
may cause operating problems associated with gypsum deposits).
One solution to these problems is a cascade pond system.
There, the reclaimed water from thickeners, filters, and
tailing ponds may be matched with the requirements for each
point of the circuit (Reference 66).
In another study (Reference 67), the many operational
problems associated with the recycling of mill water are
described in detail. The researchers have observed that
recycling at the operations studied had not caused any
unsolvable metallurgical problems and, in fact, indicate that
there are some economic benefits to be gained through decreasing
the amounts of flotation reagents required.
Mill 3103 is located in Missouri and recovered galena,
sphalerite, and chalcopyrite from 846,000 metric tons (934,000
short tons) of ore in 1973.
The mill utilizes both mine water and water recycled from the
tailing pond as feed water. The concentrator discharges
9,500 cubic meters (215,000,000 gallons) per day of tailing
slurry to its treatment facility. The treatment facility
utilizes a 42.5-hectare (105-acre) tailing pond with esti-
mated retention of 72 days, a small stilling pond at the
base of the tailing-pond dam, and a shallow 6.1-hectare (15-
acre) polishing pond before discharge. A schematic diagram of
average daily water flows for the facility is given in Figure
VII-12. Effluent chemical composition and waste load discharged
at mill 3103 using the above treatment are given in Table VII-16.
Mill 3102 is located in Missouri. This mill processed
approximately 1,450,000 metric tons (1,600,000 short tons)
of ore in 1973. Galena and sphalerite are recovered as concen-
trates at this operation.
VII-91
DRAFT
-------
DRAFT
Figure VII-12. SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT
FACILITIES AT MILL 3103
7,570 m3/day
(2,000,000 gpd)
RECYCLE
WATER
3,785 m3/day
(1,000,000 gpd)
EVAPORATION
AND
SEEPAGE
I
15,150-m3 (4,000,000-gal)
RESERVOIR
•WATER-
est 1,160 m3/day
test 300,000 gpd)
TO
SMELTER
MILL
i
9,500 m3/day
(2,500.000 gpd)
i
CONCENTRATES
37.9 m3/day
(10,000 gpd)
TO
STOCKPILE
1,515 m3/day
(400,000 gpd)
( STILLING POND
DND)
10.100 m3/day
(2.600.000 gpd)
DISCHARGE
est 3,785 m3/day
(est 1,000.000 gpd)
VII-92
DRAFT
-------
TABLE VII-16. CHEMICAL COMPOSITIONS AND WASTE LOADS FOR RAW AND
TREATED MILL WASTE WATERS AT MILL 3103
PARAMETER
PH
TSS
COD
Oil and Gr«M
Cyanide
Ho
Pb
Zn
Cu
Cd
Cr
Mn
Total Fa
MILL RAWWASTEWATER
CONCENTRATION (mg/fj
THIS PROGRAM*
7.9"
464.000
til
3.0
<0.01
< 0.0001
03
0.12
0.36
0.011
<0.02
003
0.05
HISTORICAL*
7.9* •
1.7
-
-
-
-
0.107
0.288
0014
0.001
0.002
0.169
0.03
RAW WASTE LOAD par unit me millad
kg/1000 matnc tana
1490.000"
400
12
< 0.024
0.00024
0.480
0.288
0.865
0.026
0.048
0.072
0.12
lb/1000 chart lam
2.180.000*'
800
24
< 0.048
< 0.00048
0.860
0.576
1.730
0.052
< 04)96
0.144
0.24
FINAL TREATED DISCHARGE
CONCENTRATION (mg/l)
THIS PROGRAM*
13"
16
726
30
<0.01
< 00001
0.1
0.07
<0.02
< 0.002
<0.02
0.05
0.09
HISTORICAL*
J3"
1.4
-
-
N/A
-
0.028
04)45
0.006
< 0.001
0.001
0074
0.032
EFFLUENT WASTE LOAD par unit on mllad
kg/1000 matric torn
40
1.700
7
0024
000024
0.24
0.168
< 04)48
< 0.005
< 0.048
0.12
0.282
lb/1000 abort tons
80
3,400
14
0.048
< 040048
048
0336
< 0.010
< 04)96
04)24
0.564
i 3
5 8
•Data band on 4-hour compotita nmplai
Data average ovar period January through October 1974
• 'Value in pH unit!
Raw watta load computed on bain of average flow and avaraga ore milled par day
NA - Not Available
-------
DRAFT
The mill utilizes mine water exclusively as feed. It discharges
15,150 cubic meters (4,000,000 gallons) per day of tailing slurry
to a large tailing pond. This pond also receives about 3,785
cubic meters (1,000,000 gallons) per day of excess mine water
and another 3,785 cubic meters (1,000,000 gallons) per day of
surface-drainage water. This tailing pond presently occupies
32.4 hectares (80 acres) and will occupy 162 hectares (400 acres)
when completed to design. The tailing-pond decant water is
discharged to a small stilling pool and then enters a meander
system, where biological treatment occurs. An additional
sedimentation basin of approximately 6.1 hectares (15 acres),
for removal by sedimentation of any algae which breaks loose
from the meander system, has been constructed near the end of
the meander system for use Just before final discharge. A
schematic diagram of the mill operation and the treatment facility
is presented in Figure VII-13.
Water characteristics for the effluent from the mill, the overflow
from the tailing pond, and the final discharge treated utilizing
the above technology are presented in Table VII-17.
Mill 3105 is located in Missouri and recovered galena, spha-
lerite, and chalcopyrite from an estimated 997,000 metric
tons (1,100,000 short tons) of ore in 1974.
This mill utilizes water recycled from its tailing-pond system
and makeup water from its mine as feed water. The mill dis-
charges 7,910 cubic meters (2,090,000 gallons) per day of wastes
to a 11,8-hectare (29-acre) tailing pond. The decant from this
pond is pumped to an 7.3-hectare (18-acre) reservoir, which also
receives the required makeup water from the mine. The mill
draws all its feed water from this reservoir. No discharge
occurs from the mill.
A schematic diagram of the water flows and treatment facilities
is presented in Figure VII-14.
Mill 3101 is located in Maine and recovered sphalerite and
chalcopyrite from 208,000 metric tons (230,000 short tons)
of ore in 1973.
This mill utilizes only water recycled from its treatment
facilities as feed water. The mill discharges to a mixing tank,
VII-94
DRAFT
-------
DRAFT
Figure VII-13. SCHEMATIC DIAGRAM OF WATER FLOW AND TREATMENT
FACILITIES AT MILL 3102 (TAILING POND/STILLING POND/
BIOLOGICAL TREATMENT/POLISHING POND)
TO ATMOSPHERE
15.150 m3/dav
14,000,000 gpd)
22,300 m3/day
(5.900,000 gpd}
I
34.100 m
<9,000,00
r
DISCHARGE
7,560 mj/day
(2,000,000 gpd)
9,100ms/day
(2,400.000 gpd)
NATURAL
SPRING
VII-95
DRAFT
-------
DRAFT
TABLE VI1-17. CHEMICAL COMPOSITION AND WASTE LOADING FOR RAW AND
TREATED MILL WASTEWATER MILL 3102
PARAMETER
pH
TSS
COD
Oil ind Orana
Cyanide
MB
Pb
Zn
Cu
Cd
Cr
Mn
Total ft
PARAMETER
pH
TSS
COO
Oil and GIHM
Cyanida
H«
Pb
Zn
Cu
Cd
Ci
Mn
Total Fa
MILL RAW WASTEWATER
CONCENTRATION
(mg/ei*
8.8"
248.000
488.1
0
0.03
< 00001
1.9
046
<0.02
0.006
<0.02
0.08
053
RAW WASTE LOAD
par unit era mllM
kg/1000 mettle lorn
_
900.000
1.400
0
04187
< 0.0003
6.5
1.33
< 0.0068
0.014
< 0.0068
OJ32
164
lb/1000 riion torn
_
1,800.000
2,800
0
0174
< 00006
11
266
< 0X1116
0028
<00116
0464
308
TAILING-POND DECANT
CONCENTRATION
Img/ll'
7.8"
18
663.8
8X1
< 0X1001
036
OM
<0.02
0002
<0.02
028
016
FINAL DISCHARGE
CONCENTRATION
Img/il
THIS
PROGRAM*
78"
a
11*
3jO
<001
< 00001
<01
0.04
<0.02
0005
<001
-
0002
OJ006
O.O01
1
-
-
0.003
WASTE LOAD
par unit ora millad
kg/1000 matnc tana
-
86
88
26
0082
< 0.0003
0.26
0.1
< 04)13
< 04168
04
0.325
lb/1000 dion lorn
-
132
198
60
< 0.1 74
< OXM06
060
OJ
<01
< 0X126
<0116
0*
0.65
WASTE LOAD
par unit era millad
kg/1000 matrlc ton
_
464
1.600
174
<0029
< 0X003
1
0.84
< 0.058
00058
<0068
0.81
0464
lb/1000 ahort torn
-
828
3.200
348
< OXI5B
<0116
< 001 16
< 0.1 16
3162
0.828
•Data
*Data avaraga mar pariod January through Saptambar 1874
"Valua In pH unltj
VII-96
DRAFT
-------
DRAFT
Figure VM-14. SCHEMATIC DIAGRAM OF WATER FLOW AND TREATMENT
FACILITIES AT MILL 3105
MINE
10,900 m3/day
(2,880,000 gpd)
8.300 m3/day
(2,190,000 gpd)
2,615 m3/day
(690,000 gpd)
7.3-hectare
(18-acre)
RESERVOIR
MINE-WATER
TREATMENT
8,300 m3/day
(2,190,000 gpd)
DISCHARGE
7,900 m3/day
(2,090,000 gpd)
MILL
I
TAILINGS
(35% SOLIDS)
5.510 m3/day
(1.460,000 gpd)
7,900 m3/day
(2.090,000 gpd)
2,380 m3/day
(630,000 gpd)
RECYCLE
5.300 m3/day
(1,400,000 gpd)|
VII-97
DRAFT
-------
DRAFT
where nine water Is treated by chemical precipitation that
is achieved by combustion with the tailing slurry and liming
as required. This combined waste is introduced into a tailing
pond, which discharges to an auxiliary pond. The combined
retention time in the two ponds is 35 days at maximum flow.
A schematic diagram of the mill-water circuit is shown in
Figure VII-15. The separate treatment of mine water and surface
runoff would allow this operation to achieve total recycle.
Discharge data for this mine/mill complex were presented as
mine discharge for mine 3101 earlier in this section.
Gold Ores
The discussion that follows describes treatment and control
technology in current use in the gold-ore mining and dressing
industry. Aspects of treatment and control which are unique
to the gold-ore category are described, in addition.
Mining Operations. Wastewater treatment at mining operations
in the gold-ore mining industry consists of three options as
currently practiced in the U.S.: (1) Direct discharge without
treatment; (2) Incorporation of mine water into a mill process-
water circuit; and (3) Impoundment and discharge. Impoundment
of mine water without discharge may be currently practiced at
locations in arid regions, due to evaporation. Direct discharge
of mine waters with high suspended-solid content is one po-
tential hazard associated with direct discharge—particulary,
with respect to placer, dredging, or hydraulic mining opera-
tions. Current best practice in this segment of the industry is
use of the dredge pond or a sedimentation basin for settling,
and the use of tailing gravel and sands for filtration of the
discharge stream. Levels of suspended solids attained routinely
with this method can be approximately 30 mg/1, or less if an
adequate residence time for the wastewater in the impoundment
can be obtained. Few, if any, placer mines currently provide
treatment for the wastes generated from their stripping and
sluicing operations. Settling ponds, however, have been widely
shown to be effective in improving water quality by reducing
turbidity.
Techniques used for the control of suspended solids discharged
from placer mining operations, regardless of size, are not
being employed on a major scale at present. The termination
of mining operations, even with treatment facilities, does
not eliminate water-quality degradation, however, because most
VII-98
DRAFT
-------
DRAFT
Figure VII-15. SCHEMATIC DIAGRAM OF TREATMENT FACILITIES AT MILL 3101
2.15 m3/min (569 gpm)
0.19 m3/min
(50 gpm)
0.38 m3/min
(100 gpm)
0.03 m3/min (8 gpm)
SHIPPED WITH
CONCENTRATE
PROCESS WATER
I
VACUUM PUMP
2.01 m3/min (531 gpm)
0.11 m3/min (30 gpm)
ESTIMATED
MINE WATER
0.47 m3/min
(125 gpm)
DRILL WATER
0.19 m /min
(50 gpm)
RUNOFF
FROM RAIN
COOLING WATER
AND UTILITIES
NEUTRALIZATION
AS REQUIRED
0.33 to 1.88 m3/min (87 to 497 gpm)
(MONTHLY AVERAGES)
0.99 m3/min (262 gpm)
(YEARLY AVERAGE)
DISCHARGE
0.16 m3/min
(42 gpm)
EVAPORATED
VII-99
DRAFT
-------
DRAFT
operations which use impoundment usually construct the settling
or tailing pond adjacent to the stream being worked. With
erosion taking place continuously, these facilities are seldom
permanent.
Mining operations exploiting lode ores which discharge mine
water from open-pit or underground operations commonly dis-
charge directly to a receiving stream, provide process water
for a mill circuit, or discharge wastewater to a mill tailing
pond. An example of the effectiveness of settling on water
quality is discussed under Gold-Ore Milling Operations. Mill
tailing ponds have demonstrated effective treatment, primarily
for suspended-solids removal, but secondarily for heavy-metal
removal.
Open-pit gold mining operations in arid regions often have
little or no mine discharge, whereas underground mines typically
discharge water from the mines.
Milling Operations. In-plant control techniques and processes
used by the gold milling industry are processes which were
designed essentially for reagent conservation. These processes
are the reagent circuits indicated in the process diagrams of
Figures III-9 and 111-10.
In the cyanidation process used at mills 4101, 4104, and 4105,
gold is precipitated from pregnant cyanide-leach solutions
with zinc dust. The precipitate Is collected in a filter
press, and the weak, gold-barren cyanide solution which remains
is recycled back to the leaching circuit. This solution may
be used as a final weak leach, or the solution may be returned
to its initial concentration with the addition of fresh cyanide
and used as a strong leach. In these processes, recycling of
cyanide reagent effects an estimated 33- to 63-percent saving
of this reagent. Loss of cyanide from the mill circuit is
primarily through retention in the mill tailings. Recycling
of cyanide reduces the quantity of cyanide used and also reduces
the amount of reagent present in effluent from discharging
mills.
In a similar manner, mercury is typically recycled in amalga-
mation processes. Currently, amalgamation is practiced at
only one milling operation (mill 4102) . This mill uses a
barrel amalgamation process to recover gold. At this mill,
VII-100
DRAFT
-------
DRAFT
the gold is separated from the amalgam in a high-pressure
press, and the mercury is returned to the amalgamator for
reuse. Some mercury is lost from this circuit—primarily,
through retention in the mill tailings.
Ultimate recovery or removal of mercury from the waste stream
of a mill presents an extremely difficult task. To do so
requires removing a small concentration of mercury, usually
from a large volume of water. Advanced waste treatment methods,
such as ion exchange, might achieve as much as 99 percent removal,
but the expense for treating large volumes of water would be high.
Primarily as a result of this, and in light of recent stringent
regulation of mercury in effluents, the gold milling industry
has been, taking advantage of the process flexibility available
to it and has, for the most part, replaced amalgamation with
cyanidation processes for gold recovery. This process flexi-
bility is the best control currently being practiced by the
industry for minimizing or eliminating mercury waste loading.
The primary wastes emanating from a gold mill are the slurried
ore solids. For this reason, mill effluents are typically
treated in tailing ponds, which are designed primarily to
provide for the settling and collection of the suspended solids
in the mill tailings. In most cases, these operations dis-
charge from tailing ponds, and the usual practice is to decant
the water from the top of the pond at a point where maximum
clarification has been attained. In some facilities, two or
more ponds are connected in series, and wastewater is decanted
from one to another before final discharge.
Although the structure, design, and methods of ponding may
vary somewhat in accordance with local topography and volume
of wastewater, the desired goal is the same—to achieve
maximum settling and retention of solids.
To illustrate the effectiveness of settling ponds as treatment
systems in the gold-ore milling industry, the discussion
which follows outlines an operation which recovers gold and
other metals and treats wastewater by use of a tailing pond.
Mill 4102 is located in Colorado. This mill beneficiates
ore containing sulfides of lead, zinc, and copper, in
addition to native gold and silver. During 1973, 163,260
metric tons (180,000 short tons) of ore were milled to
produce lead/copper and zinc concentrates by flotation
and gold by amalgamation.
VII-101
DRAFT
-------
DRAFT
Makeup water for the mill circuit is drawn from a nearby creek.
This water is introduced into the grinding circuit for trans-
portation and flotation of the ore. Prior to entering the
flotation circuit, the ground ore is jigged to produce a
gravity concentrate. This concentrate contains most of the
gold, which is recovered by amalgamation. After amalgama-
tion, the jig concentrate is fed into the flotation circuit,
because some lead is contained in the material.
Mill tailings are discharged to a tailing pond at a rate of
2,290 cubic meters (600,000 gallons) per day. Decant from
this pond flows to a smaller polishing pond prior to final
discharge to a stream. The tailing pond and the polishing
pond have a total area of 18.2 hectares (45 acres).
Table VII-18 presents the chemical composition of mill water
and raw and treated waste load for mill 4102, which practices
amalgamation for gold and froth flotation for sulfide minerals.
These data indicate that removal of selected metals is achieved
to a degree; however, the treatment is most efficient in the
removal of suspended solids.
Mill 4101 is located in Nevada. This mill recovers gold
occurring as native gold in a siltstone host rock which
is mined from an open pit. Schuetteite (HgS04^2HgO)
also occurs in the ore body, and mercury is recovered
as a byproduct during furnacing of the gold concen-
trate. Ore milled during 1973 totaled 750,089 metric tons
(827,000 short tons). This figure normally is 770,950 metric
tons (850,000 short tons) but was lower than usual due to a
20-day labor strike.
This mill employs complete recycle of the tailing-pond decant.
However, due to consumptive losses, some makeup water is re-
quired, and this water is pumped to the mill from a well.
Water is introduced into the grinding circuit for transportation
and processing of the ore by the agitation/cyanidation-leach
method.
Mill tailings are discharged at a rate of 2,305 cubic meters
(603,840 gallons) per day to a 37-hectare (92-acre) tailing
pond. Approximately 1,227 cubic meters (321,500 gallons)
per day of tailing-pond decant are pumped back to the mill
VII-102
DRAFT
-------
DRAFT
TABLE VII-18. WASTE COMPOSITIONS AND RAW AND TREATED WASTE LOADS
ACHIEVED AT MILL 4102 BY TAILING-POND TREATMENT
PARAMETER
pH
TSS
COD
Oil ind OrMM
Cd
Cr
Cu
Toltl F.
Pb
Totil Mn
MS
Zn
MILL WASTEWATER
CONCENTRATION
(mo/ HI
9.1-'
496.000
1142
1
<002
<0.02
0.03
1.0
<01
626
00014
1.3
RAW WASTE LOAD
par unit or* milled
kg/1000 imtrie torn
-
2.871.000
66
6.8
<0.12
< 012
0.17
6
<0.6
48
0008
76
Ib/IOOOilwttom
-
5.742.000
132
116
<0.24
<024
0.34
12
<12
86
0016
1B.O
TAILING-POND EFFLUENT
CONCENTRATION
(me/11
10A*
4
2235
1
<0.02
0.06
1.2
1.8
<0.1
637
00011
0X15
TREATED WASTE LOAD
kg/1000 nntric ton
-
20
130
6
<01
OJ
7
8
<0.6
40
0.006
OJ
lb/1 000 ihort torn
-
40
260
12
<1.2
BO
0.012
04
•Vilu. In pH unlu
f'induflry diU monthly tvnigi ovw pwiod Nwnnbw 1873 through Novtmtan 1874
VII-103
DRAFT
-------
DRAFT
from a reclaim sump. No discharge from this operation results.
Potential slime problems in the mill circuit are controlled
through adjustment of the pH to 11.7 and by use of Separan
flocculant in the circuit.
Table VII-19 gives the results of chemical analysis of mill
effluent and tailing-pond decant water after treatment. No
waste loadings are given, since no discharge results. Samples
were obtained from this facility to determine the effectiveness
of treatment, even though the mill has no discharge. Note,
however, that this mill has an alkaline-chlorination unit avail-
able for use in cyanide destruction should emergency conditions
require a discharge.
Data from both mills indicate that dissolved heavy metals
are removed to some degree in the tailing pond, but more
effective technology is required for removal of these waste
constituents. Although such technology is not currently used
in the gold mining and milling Industry, it is currently
available. This technology also has special application to
mine discharges, as they usually contain relatively high
dissolved-metal loads. This technology will also be appli-
cable to those situations where sufficient reduction of
metals and cyanide in tailing-pond effluents is not being
achieved.
Conventional treatment available for dissolved heavy metals
generally involves:
(1) Coagulation and sedimentation employing alum, iron
salts, polyelectrolytes, and others.
(2) Precipitation with lime, soda ash, or sulfides.
These treatment technologies have been previously discussed
in this section. Treatment by these methods is not normally
practiced in this industry category. However, where metal
mining wastes are treated, the most common means used is to
discharge to a tailing pond, in which an alkaline pH is
maintained by lime or other reagents. Heavy-metal ions are
precipitated at elevated pH; these ions are then settled out,
together with suspended solids, and maintained in tailing
ponds.
VII-104
DRAFT
-------
DRAFT
TABLE VII-19. CHEMICAL COMPOSITIONS OF MILL WASTE WATER AND
TAILING-POND DECANT WATER AT MILL 4101 (NO
RESULTANT DISCHARGE)
PARAMETER
pH
TSS
Turbidity (JTU)
TDS
COD
Oil and Grease
Cyanide
As
Cd
Cr
Cu
Total Fe
Pb
Total Mn
Hg
Zn
CONCENTRATION (mg/ £ )
MILLWASTEWATER
12.26*
545,000
6.70
4,536
43
<1
5.06
0.05
0.10
0.06
0.17
<0.5
< 0.1
0.02
-
3.1
TAILING-POND DECANT
11.29*
12
1.0
4,194
43
<1
5.50
0.04
0.02
0.03
0.13
< OS
< 0.1
0.90
0.152
2.5
•Value in pH units.
VII-105
DRAFT
-------
DRAFT
Mercury presents a special problem for control, due to its
potential for conversion in the environment to its highly
toxic methyl-mercury form. The amalgamation process still
finds some use in the gold milling industry, and, in addition,
this metal sometimes occurs with gold in nature. Although
mercury will precipitate as the hydroxide, the sulfide is
much more insoluble. It is expected that, where dissolved
mercury occurs in mine or mill wastes, it will be treated for
removal by sulfide addition. This reaction requires alkaline
conditions to prevent the loss of sulfide ion from solution
as H2S. Theoretical considerations of solubility product and
dissociation equilibria suggest that, at a pH of 8 to 9,
mercury ion will be precipitated from solution to a concentra-
tion of less than 10 exp(-41) g/1. In practice, it is not
likely that this level can be achieved. However, by optimi-
zing conditions for sulfide precipitation, mercury should be
removed to a concentration of less than 0.1 microgram/liter
(0.1 ppb).
The conditions under which lime precipitation of heavy metals
is achieved must take into consideration auxiliary factors.
As indicated, the most important of these factors is pH.
The minimum solubility of each metal hydroxide occurs at a
specific pH; therefore, optimum precipitation of particular
metals dictates regulation and control of pH. When more than
one metal is to be precipitated, the pH must necessarily be
compromised to obtain the maximum coprecipitation achievable
for the given metals.
Another factor which must be considered is the oxidation
state(s) of the metal or metals to be treated. For example,
As(+5) is much more amenable to chemical treatment than is
As(+3). In addition, cyano-metallic or organo-metallic com-
plexes are generally much more difficult to remove by chemical
treatment than are free metal ions. Where these factors
impede chemical treatment, prior oxidation of the waste stream
can be employed to destroy the metal complexes and oxidize
metal ions to a form more amenable to chemical treatment.
This oxidation may be achieved by aeration of the waste stream
or by the addition of chlorine or ozone.
To achieve high clarification and removal of solids and chemi-
cally treated metals, it is essential to provide good sedimenta-
tion conditions in the tailing pond. Typically, this is done
in the industry by designing tailing ponds to provide ade-
VII-106
DRAFT
-------
DRAFT
quate retention time for the settling of solids and metal
precipitates. Specification of a recommended retention time
for traditional tailing-pond design is problematical, because
the influence of pond geometry, inlet/outlet details, and other
factors that ensure even distribution and an absence of
short-circuiting are of greater importance than the theoretical
retention provided. A design retention time of 30 days, based
on the average flow to be treated, is often specified and is
appropriate if short-circuiting due to turbulence or stratifi-
cation does not occur. The use of a two-cell pond is recom-
mended to increase control and reliability of the sedimentation
process.
In some cases, suspended solids or metal precipitates may
retain surface charges or colloidal properties and resist
settling. These solids and colloids can be treated for
removal by the addition of coagulating agents, which either
flocculate or act to neutralize or Insulate surface charges
and cause the suspended solids and colloids to coagulate and
settle. These agents may be such flocculants as alum
(A12_(S04_)j3) or iron salts, or such coagulants as clays,
silica, or polyelectrolytes.
Cyanide destruction has been previously discussed in this
section. The technology for oxidation and destruction of
cyanide is well-known and currently available. Where dis-
charges of cyanide have the potential to enter the environment,
complete destruction prior to discharge is recommended.
Technology For Achieving Np_ Discharge £f_ Pollutants. Elimin-
ation of point discharges is currently being achieved in the
industry by two slightly different technologies: impoundment
and recycle. Where impoundment is used, the mill tailings
are simply discharged to a pond and retained there. Recycling
exists where tailing-pond water is decanted and returned to
the mill for reuse. A mill or mine/mill complex is potentially
capable of employing either of these technologies, whereas a
mine alone may only be able to make use of impoundment.
The feasibility of impoundment is dependent on the overall
water balance of the location of the mine/mill's mine or mill.
In arid regions, the impoundment of tailings is a feasible
alternative to discharging and is, in fact, being practiced.
VII-107
DRAFT
-------
DRAFT
Where recycle systems are employed, the design must also
take water balance into consideration. In those areas where
precipitation exceeds evaporation during all or part of the
year, some system to divert runoff away from the tailing
pond is required to keep excess water in the pond to a minimum.
Also, where heavy rainfalls periodically occur, tailing ponds
must he designed to hold the excess water accumulated during
these periods. A mine/mill complex may find it necessary to
segregate the mine and mill effluents to further relieve the
recycle system of excess water. In such cases, it is expected
that the mine effluents will be treated by the chemical methods
discussed previously and then discharged.
To some extent, a mill may depend on inherent loss of water
from the system to maintain a balanced recycle system. These
losses include any or all of the following:
(1) Consumptive losses in the milling process (i.e.,
retention of mositure in the concentrate, etc.);
(2) Retention of moisture by the tailing solids in the
tailing pond;
(3) Evaporation;
(4) Seepage and percolation of water from the tailing pond.
The extent of these losses is dependent on a number of factors,
namely:
(1) Milling process employed;
(2) Evaporation rate (function of climate and topography);
(3) Type of material used to construct the tailing pond;
(4) Characteristics of tailing solids;
(5) Characteristics of soil underlying the tailing pond;
(6) Use of liners, diversion ditches, and other methods.
VII-108
DRAFT
-------
DRAFT
Given the present state of technology available and the
demonstrated status of recycle within the gold milling industry,
the maintenance of a balanced recycle system is technologically
feasible.
The feasibility of a recycle system must also consider the
effects of the reclaim water upon the mill circuit. For example,
it has been indicated previously that reclaiming cyanidation-
process water could result in a loss of gold should this water
be introduced at the ore-grinding stage.
In the Province of Ontario, it has been found that the level of
cyanide in the tailing-pond decant from active mine/mill oper-
ations approximates 0.02 to 0.5 percent of total cyanide mill
additions (Reference 59 ). However, data indicate that
the concentration of cyanide in tailing-pond decant may build
up if the decant is being reclaimed. If this occurs, the
alkaline-chlorination method can be used for cyanide destruction.
Complete destruction of cyanide can be achieved by excess
addition (8.5:1) of chlorine. On this basis, the recycling
of cyanidation-process water is considered technologically
feasible.
Recycling and zero discharge are currently being accomplished
at mill 4101, which is milling gold by the cyanide/agitation-
leach process (Figure 111-10). The overall water balance
for this mill has been presented in Figure V-22. Treatment
efficiency data for this mill, presented in Table VII-19,
indicate a buildup of dissolved solids and cyanide in the
reclaim water. However, no loss in percent recovery as a
result of recycling has been reported by this mill. In addi-
tion, the recovery rate for this mill does not differ from
that of cyanidation mill 4105, which does not recycle process
water.
Silver Ores
The discussion which follows describes treatment and control
technology currently employed in the silver-ore mining and
dressing industry. Aspects of treatment and control pertaining
to the silver-ore category are described.
Mining Operations. Wastewater treatment at silver mining
operations primarily consists of discharge of wastewater to
a mill tailing pond, or direct discharge without treatment.
VII-109
DRAFT
-------
DRAFT
Mining of silver ores primarily exploits the sulfide minerals
tetrahedrite ((Cu, Fe, Zn, Ag)12Sb4S13) and argentite (Ag£S)
and native silver. Native silver often occurs with gold,
copper, lead, and zinc minerals. Little water use is encountered
in silver-ore mining, with the exception of dredging, where
silver is recovered as a minor byproduct.
Separate treatment of mine water per ee_ is not typically
practiced in this industry; however, where practiced, treatment
is performed in conjunction with treatment of mill wastewater
in a tailing pond.
Milling Operations. As discussed in Section V, milling
processes currently employed in the silver industry are froth
flotation (about 99 percent of U.S. mill production), cyanidation
of gold ores, and amalgamation. Cyanidation and amalgamation
recovery of silver currently constitute approximately 1 percent
of U.S. silver production by milling. The occurrence of silver,
either with gold in a free state or as a natural alloy with
gold, has also resulted in production of silver at refineries.
Silver is often recovered also as a byproduct of the smelting
and refining of copper, lead, and zinc concentrates.
Cyanidation for gold and silver is currently being practiced
at mill 4105 (gold category), but wastewater treatment tech-
nology as currently practiced consists of a sand reclaimer pond
for removal of coarse solids only. Amalgamation for gold
and silver is currently limited to one known site. Waste-
water treatment at this facility has been described previously
for mill 4102.
Mill 4105, which recovers both gold and silver, currently
practices in-plant recycling of reagents, as indicated in
Section III for Gold Ores. This results in economies of
both cost and reagent use, as well as prevention of the dis-
charge of cyanide for treatment or into the environment.
In-plant control practices common to silver flotation mills
are based on good housekeeping measures, employed to prevent
spills of flotation reagents. The feed of these reagents into
a circuit Is carefully controlled, because a sudden increase
or decrease of some reagents could have adverse effects on
recovery from the flotation circuit.
VII-110
DRAFT
-------
DRAFT
Wastes resulting from silver milling are typically treated
in tailing ponds. These ponds function primarily to facilitate
the settling and retention of solids. Except in the case of
total impoundment, the clarified tailing-pond water is currently
discharged. At mill 4401, a further reduction of waste loading
is achieved by partial recycle of the tailing-pond decant water
(approximately 60- to 75-percent recycle). Mill 4402 has
achieved zero discharge through total recycle of tailing-
pond decant water. Flotation is the predominant method
currently used to concentrate silver ore. Flotation circuits
are commonly run under alkaline conditions. For example,
soda ash, caustic soda, and hydrated lime are added to the
circuit of mill 4402, and lime is added to the circuit of
mill 4401. These reagents are added to the mill circuits
to act as depressants and pH modifers and consequently make
the tailing pond alkaline. This facilitates the removal of
metals as hydroxides in the tailing pond. However, note that
the reagents producing an alkaline pH in the tailing pond are
added in the mill to control the process conditions there, and
a high degree of control over the pH in the tailing pond is
not currently practiced in the industry. To facilitate opti-
mum precipitation of metal hydroxides in the tailing pond, a
higher degree of control over the pH may be required in some
cases. Highly alkaline conditions (pH range of 10 to 11)
may be required to effect greater removal efficiency in treat-
ment facilities.
The removal of antimony from silver-milling wastewater presents
a special problem, as the hydroxide may be very unstable
and is not reported to exist. The presence of antimony in
wastewater has been noted, because it is closely associated
with silver in some ore bodies—especially, those of the
Coeur d'Alene District of Idaho. The sulfide of antimony
is highly insoluble; therefore, treatment for antimony removal
will involve sulfide precipitation. Although Na2_S is itself
toxic at high concentrations, the amount required to treat
the levels of antimony found in mine and mill wastewater
(approximately 2 to 3 mg/1) is small (approximately 1 mg/1)
and will be consumed in the precipitation reaction. Sulfide
precipitation must be carried out under alkaline conditions
to prevent the removal of sulfide ion from solution as H2J5 gas.
Cyanide is used as a pyrite depressant at mill 4401.
VII-111
DRAFT
-------
DRAFT
This mill is also recycling its process water with no apparent
adverse affects from this reagent. However, should the destruction
of cyanide become necessary for process control or as a safety
measure in treating accidental leaks from the treatment system,
the alkaline-chlorination method is an effective treatment
for the destruction of cyanide. This process has been discussed
previously in this section. An example of tailing-pond
treatment as practiced at mill 4401 is described below.
Mill 4401 is located in Idaho. Ore is brought to the
mill from an underground mine. Valuable minerals in
the ore body are primarily tetrahedrite, but chal-
copyrite and galena also occur. During 1973, 182,226 metric
tons (200,911 short tons) were milled to produce a copper/
silver concentrate.
Water used at the mill consists of both reclaim water and
makeup water, pumped from a nearby creek. This water is Intro-
duced into the grinding circuit for the transportation and
flotation of the ground ore. Mill tailings are discharged
at a rate of 3,188 cubic meters (835,200 gallons) per day to
the tailing-pond system. This system is composed of three
tailing ponds and a clarification pond. Two of the tailing
ponds are inoperative, due to extensive damage resulting from
a recent flood. Prior to this flood, tailings were distributed
to the three ponds, and their decant was pumped to the clarifi-
cation pond. This system covers a total area of 4.5 hectares
(10.9 acres). Presently, water is both discharged and recycled
back to the mill from the clarification pond. Approximately
1,649 cubic meters (432,000 gallons) per day are recycled,
while 1,141 cubic meters (299,000 gallons) per day are discharged,
Mine water is also discharged to this pond system at a rate of
553 cubic meters (145,000 gallons) per day.
A new tailing pond is under construction and is expected to be
in use soon. This pond will have an area of 6.9 hectares
(17.0 acres).
Table VII-20 gives the chemical composition of raw and treated
waste loads from mill 4401, which uses tailing pond
treatment. Decreases in several parameters, in addition
to suspended-solid removal, are noted. COD, cyanide, copper,
mercury, and nickel are all reduced significantly, while
important reductions of TOC and antimony are also observed.
VII-112
DRAFT
-------
DRAFT
TABLE VII-20. WASTE COMPOSITIONS AND RAW AND TREATED WASTE LOADS
AT MILL 4401 (USING TAILING-POND TREATMENT AND
PARTIAL RECYCLE)
PARAMETER
PH
TSS
Turbidity IJTUI
COD
TOC
Oil ind Gram
Cyinid*
At
Cd
Cr
Cu
Totil ft
Pta
Mn
Hg
Ni
Ag
Zn
Sb
MILL WASTEWATER
CONCENTRATION
(mg/D
-
555.000
20
596
220
7
DOS
<007
<002
<01
-
0.0024
014
<002
<002
185
RAW WASTE LOAD
pir unit or* millvd
kg/1000 mmic torn
-
2.497.000
-
268
100
30
023
<011
<003
<016
1 1
-
<003
<003
83
lb/1000 than torn
-
4.904.000
-
636
200
60
046
<022
<006
<032
22
-
<032
-
0022
126
<006
<006
166
TAILING-POND EFFLUENT
CONCENTRATION
(mg/U
70'
<2
06
198
175
12
0026
<007
<002
<01
008
010
<01
039
00006
006
<3
-
32
28
18
0.04
<011
<003
<016
013
016
<0.16
062
00008
01
<003
003
16
lb/1000 than toni
-
<6
-
64
66
38
0.08
<022
<0.06
<0.32
026
032
<032
1 24
00016
02
<006
006
32
•V.Iu. in pH uniti
VII-113
DRAFT
-------
DRAFT
Control and Treatment Technology To Achieve No Discharge.
Currently, two silver mills are recycling their process water.
Mill 4402 reclaims all of its tailing-pond decant, while mill
4401 presently reclaims approximately 60 percent of its
tailings-pond decant. However, operation of mill 4401 with
complete recycle could be achieved, and would be, were it not
currently less expensive to use fresh water pumped from a
nearby well, rather than recycled process water from an impound-
ment as makeup water.
The feasibility of recycle entails consideration of the overall
water balance at a given mill and possible interferences in
the mill circuit caused by the recycling of process reagents
and/or buildup of dissolved solids. Water-balance considerations
and recycling of cyanide reagent have been discussed previously
in Section VII.
Silver ores are concentrated primarily by the froth flotation
process, and it has been noted previously that recycled flota-
tion reagents might interfere with the mill circuit. However,
no published data exist which would support this position.
Recycling successfully being carried on at mill 4402 (total
recycle—no discharge) and mill 4401 (partial recycle) demon-
strates the feasibility of achieving total recycle and zero
discharge. It is expected that unwanted quantities of a
particular frother appearing in a recycle stream (from a
tailing area, etc.) can probably be reduced or eliminated by:
(1) increasing the retention time of the frother-con-
taining wastes to facilitate increased oxidation or
blodegradation before recycle to the mill; or
(2) oxidation of the frothers through application of a
degree of mechanical aeration, etc., to the waste
stream; or
(3) selecting another frother with superior breakdown
properties for use in the mill.
A further degree of control of the recycle system can be
gained by use of a two-cell pond. In this system, clarified
water from the primary pond would be decanted to the second
VII-114
DRAFT
-------
DRAFT
pond, which would be used as a surge basin for the reclaim
water. This system would lend itself to increased control
over the slime content of reclaim water. This is desirable,
since these slimes have been thought to inhibit differential-
flotation processes in some mills. In addition, the second
pond would provide a site for the implementation of mechanical
aeration, should this treatment become necessary.
Segregation of Waste Streams. At certain mine/mill complexes,
for the mill to achieve a balanced recycle system, it may be
necessary to segregate the mine and mill waste streams.
In such cases, it is expected that, prior to discharge, the
mine effluents would be chemically treated for the removal
of metals and suspended solids in settling ponds. As pre-
viously discussed, this treatment would normally involve
precipitation of metals using lime and/or sulfides.
The discussion which follows describes a silver milling
operation currently operating with recycle and zero discharge.
Mill 4402 is located in Colorado. Ore is brought to
the mill from an underground mine. Valuable minerals
in the ore body include sulfide of silver—primarily,
argentite, galena, and free or native silver. During 1973,
75,005 metric tons (82,696 short tons) of this ore were milled
to produce a lead/silver concentrate.
Process water is recycled at this mill. However, makeup water
is required, and this water is pumped from a well. Water is
introduced into the grinding circuit to facilitate trans-
portation and flotation of the ground ore. Mill tailings are
sent through two stages of cyclones to remove sands, which are
used for backfilling stopes in the mine. Cyclone overflow
is discharged to a 1.6-hectare (4-acre) tailing pond at a rate
of 1,511 cubic meters (396,000 gallons) per day. Clarified
pond water is recycled back to the mill at a rate of 962
cubic meters (252,000 gallons) per day.
A new tailing pond is being built at this mill. This pond will
have an area of 6 hectares (15 acres).
Table VII-21 demonstrates the treatment efficiency achieved
in the mill tailing pond and compares mill raw-wastewater
VII-115
DRAFT
-------
DRAFT
TABLE VI1-21. CHEMICAL COMPOSITIONS OF MILL RAW WASTE WATER
AND TAILING-POND DECANT WATER AT MILL 4402
PARAMETER
TSS
Turbidity (JTU)
COD
TOC
Oil and Grease
Cyanide
As
Cd
Cr
Cu
Total Fe
Pb
Total Mn
Hg
Ni
Ag
Zn
Sb
CONCENTRATION (ing/ 1 )
MILL RAWWASTEWATER
90,000
1.05
22.70
29.0
2
< 0.01
0.07
< 0.02
< 0.1
0.22
1.80
0.56
1.75
0.149
0.10
< 0.02
0.37
< 0.2
TAILING-POND DECANT
2
0.575
22.70
17.5
2
< 0.01
< 0.07
< 0.02
< 0.1
< 0.02
1.59
0.10
1.80
0.002
0.11
< 0.02
2.3
< 02
vri-ii6
DRAFT
-------
DRAFT
input to tailing-pond decant water recycled to the mill.
No waste loads are presented, because no discharge results.
Bauxite Ore
As discussed in Section IV, Industry Categorization, two
bauxite mines currently operating in the U.S. extract bauxite
ores from open-pit and underground mines. The characteristics
of pollutants encountered in wastewaters from these operations
are discussed in Section V. The current treatment technology
and industry practice for treatment of bauxite-mine drainage
are described below.
Lime neutralization is the only treatment method presently being
employed by the two domestic bauxite producers to treat mine
water. Both acidic and alkaline waters are treated by this
technique, but, due to the relatively small amount of alkaline
water that is treated daily (83 cubic meters, or 22,000 gallons,
per day), only acid mine-water neutralization is discussed in
detail here.
Generally, mine water and surface drainage destined for treat-
ment undergo settling in a number of natural depressions,
sumps, and settling ponds before reaching the lime-neutraliza-
tion facility; thus, suspended-solids loadings are reduced.
The addition of lime to raw mine drainage to reach elevated
pH causes precipitation of heavy metals as insoluble or
slightly soluble hydroxides. Formation of specific metal
hydroxides is controlled by pH, and removal of the suspended
hydroxides is accomplished by settling. The discussion of
this treatment technique is presented in the early portion
of Section VII under Chemical Precipitation.
Two variations of lime storage at bauxite-minewater-treatment
facilities are employed, and both systems achieve slightly
different efficiencies of pollutant removal. It is doubtful
that treatment efficiency is dependent upon lime storage
technique, but the pH and pH control of the limed solution
are probably the dominant factors in determining concentration
levels attained in settling ponds.
Figure VII-16 is a schematic flowsheet of the lime-neutraliza-
tion facility at open-pit mine 5101, which processes approxi-
mately 7,165 cubic meters (1,900,000 gallons) per day of raw
mine drainage.
VII-117
DRAFT
-------
DRAFT
Mine 5101. Open-pit mine complex 5101 is located in
Arkansas and produces about 2,594 metric tons (2,860
short tons) of high-silica bauxite daily. There are several
pits associated with the water-treatment facility, and acid
waters collected from the pits, spoils-storage areas, and
disturbed areas are directed to the treatment plant.
Mine 5101 treats the major portion of its open-pit mine drainage
through the treatment plant, as shown in Figure VII-16. Other
open-pit drainages which require intermittent pumping for
discharge will be treated by a mobile lime-treatment plant in
the near future. At this operation, about 0.45 kg (approxi-
mately 1 pound) of slurried lime is used to neutralize 3.79
cubic meters (1000 gallons) of acid mine water. This facility
has a controlling pH probe, located in the overflow from the
detention tank, which activates the automatic plant and pump
cutoffs at a high point of pH 9.0 and a low point of pH
6.0. The operating pH generally ranges from 7.5 to 8.0,
and the pH of the effluent discharge ranges from 6.3 to 7.3.
Table VII-22 lists analytical data for raw mine water (silt-
pond overflow) and treated effluent (as the discharge leaves
the overflow weir at the sludge pond).
Mine 5102. Open-pit mine 5102 is also located in
Arkansas and mines a high-silica-content bauxite deposit.
Contaminated surface drainage from outlying areas and ground-
water accumulation in the holding pond produce about 14,140
cubic meters (4,000,000 gallons) of raw drainage daily.
Surface drainage collects from an area of approximately 662
hectares (1,635 acres) of disturbed and undisturbed land.
An experimental lime-neutralization plant has been operated
at mine 5102 and processes approximately 2,650 cubic meters
(2,700,000 gallons) per day of acid mine drainage.
This mining operation presently treats less than 10 percent
of its total raw mine drainage, but full-scale operation of
a treatment plant having a capacity of 11,355 cubic meters
(3,000,000 gallons) per day is expected in mid-1975. The
new plant will operate similarly to the present plant, but
an enlarged system of settling lagoons and sludge drying
beds should provide adequate treatment efficiency.
VII-118
DRAFT
-------
DRAFT
Figure VII-16. LIME-NEUTRALIZATION PLANT FOR OPEN-PIT MINE 5102
LIME-SLURRY
STORAGE
TANKS
SLUDGE
SETTLING
POND
RAW-WATER
HOLDING
POND
CLEAR-WATER
SETTLING
POND
1.84 m°/day
(486 gpm)
1.84 m3/day
(486 gpm)
DISCHARGE
VII-119
DRAFT
-------
DRAFT
TABLE VII-22. CHEMICAL COMPOSITIONS OF RAW AND
TREATED MINE WATERS AT MINE 5101
PARAMETER
PH
Acidity
Alkalinity
Conductivity tf
TDS
TSS
Total Fe
Total Mn
Al
Ni
Zn
Sr"
Fluoride
Sulfate
CONCENTRATION (mg/£)
RAW MINE DRAINAGE
RANGE
2.8 to 3.0f
250 to 397
0
1000
560 to 617
2
7.2 to 21 .8
3.2 to 3.5
23.8 to 18.6
0.3 to 0.31
0.82 to 1.19
0.5
0.048 to 0.290
490 to 500
AVERAGE*
2.9*
324
0
1000
589
2
14.5
3.4
21.2
0.3
1.01
0.5
0.17
495
TREATED EFFLUENT
RANGE
6.0 to 6.8f
OtolO
6 to 13
1000
807 to 838
1.2 to 4.0
0.14 to 0.20
2.25 to 3.37
0.33 to 0.80
0.18 to 0.19
0.07 to 0.09
1.74
0.03 to 0.67
500 to 581
AVERAGE*
6.4*
0.5
10
1000
823
3
0.2
2.8
0.6
0.2
0.08
1.74
0.35
541
•Values based on two grab samples
f Value in pH units
"Value in mieromhos/cm
"value* based on one grab sample
VII-120
DRAFT
-------
DRAFT
The treatment used at mine 5102 involves slurried storage of
lime in large agitator tanks for eventual mixing with mine
water in the confines of a pipeline. About 0.83 kg (1.82 Ib)
of hydrated lime is used to neutralize 3.79 cubic meters
(1000 gallons) of raw mine water. This lime rate maintains
the influent to the sludge pond at a pH of 9.0 to 11.0, and
efficient from the clear-water settling pond varies from a pH
of 6.0 to 8.0.
Table VII-23 lists the chemical composition of both raw mine
water (influent to the treatment plant) and the treated efflu-
ent (discharge from clear-water settling pond).
Ferroalloy Ores
The ferroalloy-ore mining and dressing category includes,
for purposes of treatment here, operations mining and bene-
ficiating ores of cobalt, chromium, columbium and tantalum,
manganese, molybdenum, nickel, tungsten, and vanadium (one
operation extracting non-radioactive vanadium). Vanadium
obtained from milling of uranium, vanadium, and radium ores
under NRC licensing is covered as part of the uranium-ore
category. Since the subcategorization of this category is
not based upon end product recovered, but rather upon the
process used, representative mines and mills are used to
illustrate wastewater treatment and control as practiced in
ferroalloy-ore subcategories.
Currently, there are no operations mining or beneficiating
ores of chromium, cobalt, columbium , and tantalum. A mangan-
iferous ore is currently being mined at one location in the
U.S., but no wastewater results, and no milling activities
are carried on. A second manganiferous ore mine and mill
is expected to reopen in late 1975 or 1976. Consequently,
treatment and control technology currently employed in the
molybdenum, nickel, tungsten, and vanadium industries will
be used as examples here to represent treatment used in
subcategories of this category.
Mining Operations. Mining of ferroalloy ores is by both
underground and open-pit methods. Mine wastewater is
characterized by high and variable flow and dissolved heavy
metals, and is often acidic. At open-pit mines, seasonal
VII-121
DRAFT
-------
DRAFT
TABLE VI1-23. CHEMICAL COMPOSITIONS OF RAW AND
TREATED MINE WATERS AT MINE 5102
PARAMETER
pH"
Acidity*
Alkalinity
Conductivity *
TDS
TSS»
Total Fe«
Total Mn
Al»
Ni
Zn
Sr
Fluoride
Sulfate"
CONCENTRATION (mg/l)
RAW MINE
DRAINAGE
2.91"
240
0
2.21 2««
468
45
49.0
1.56
14.8
0.05
0.24
0.1
0.59
432
TREATED
EFFLUENT
7.2'
0
30
897"
630
6.6
0.29
<0.02
0.12
<0.02
<0.02
—
0.56
343
•Values based on industry samples and represent the
average of eight or more grab samples taken in 1974.
* Value in pH units
••Value in micromhos/cm
VII-122
DRAFT
-------
DRAFT
fluctuations in mine water may be extreme. At such opera-
tions, acidic streams from sulfides in mine waste dumps cdd
to the waste load of the wastewater requiring treatment.
Mine water is often used as mill process water at underground
mines. At open-pit operations, seasonal variability generally
makes mine water an unacceptable source of process water.
Treatment for suspended-solid removal is almost universally
practiced in the ferroalloy-ore mining industry. Both treatment
in tailing ponds with mill wastewater and use of separate
treatment systems such as settling ponds and clariflocculators
(variants of mechanical clarifiers in which mixing is provided
for flocculant distribution) are used. Where waste streams
are acidic, neutralization is generally practiced. Where open-
pit mining and ore stockpiling are practiced, the potential for
oxidation of metals (especially, molybdenum) increases, yield-
ing higher levels of concentration of dissolved heavy metals
and, thus, increased raw waste loads.
Examples of treatment practice are given in discussions that
follow, using mines 6103, 6104, and 6107 as examples. In
addition to these sites, mine water at mine 6102 is treated
by neutralization and by a closed-circuit mill tailing pond
from which only seasonal discharge results. Runoff from
mine 6106 is treated by settling only.
Mine 6103. This mine is an underground molybdenum mine,
in Colorado, which is still under
development. Treatment of mine water at this site during
development of the mine has included flocculant addition,
spray cooling, and solids removal in a series of three settling
ponds. Sanitary wastewater from the mine site is given tertiary
treatment in a separate facility prior to mixing with mine
water in the first settling basin. Samples of the 9,800-
cubic-meter/day (1,700-gpm) mine-water flow were obtained at
the point of discharge from the mine and at the overflow from
the third settling pond. The results of chemical analyses
of these samples of raw mine water and effluent from the
treatment system are presented in Table VII-24.
Appreciable reductions of suspended solids and the heavy metals
Cu, Mn, Pb, Zn, and Fe are evident. The influence of highly
treated sanitary waste is, apparently, reflected in elevated
COD values at the effluent from the treatment system.
VII-123
DRAFT
-------
DRAFT
TABLE VI1-24. CHEMICAL COMPOSITIONS OF RAW MINE WASTE WATER
AND TREATED EFFLUENT AT MINE 6103
PARAMETER
TSS
TDS
Oil and Grease
COD
As
Cd
Cu
Total Mn
Mo
Pb
V
Zn
Total Fe
Fluoride
CONCENTRATION (mg/H)
BEFORE TREATMENT
802.9
726
1.0
<10
<0.01
0.16
0.06
5.5
<0.1
0.19
<0.5
0.47
17.0
4.5
AFTER TREATMENT
24.3
564
1.0
67.5
<0.01
<0.01
<0.02
1.0
<0.1
0.03
<0.5
<0.02
0.17
3.7
VII-124
DRAFT
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DRAFT
Mine 6104. This mine is an underground mine, located in
California, which obtains a complex ore yielding
tungsten, molybdenum, and copper. The mine
produces approximately 2,200 metric tons (2,425 short
tons) of ore per day. Mine water pumped from the mine daily
totals 47,000 cubic meters (13,000,000 gallons), of which
approximately 7,000 cubic meters (1,848,000 gallons) are used,
untreated, as mill process water. The remainder is treated
for solids removal in a clariflocculator. Underflow from the
clariflocculator is pumped to the mill tailing pond for
further treatment. The bulk (approximately 90 percent) of
clarified overflow is discharged, with the balance used as
mill process water. Table VII-25 presents the results of
chemical analyses of raw mine water and the effluent from
the clariflocculator. A clariflocculator is used for treat-
ment because of severely limited land and space availability
in this area of very high relief (steep terrain). The use of
ammonium nitrate-based blasting agents previously contributed
to elevated nitrate and nitrogen levels in mine wastewater.
This situation has been largely alleviated by a change in
explosives used at the mine.
In addition to a significant reduction of suspended-solid
concentrations, important reductions of Pb, Mn, and Fe have
been noted.
Mine 6107. This mine is an open-pit vanadium mine, working
non-radioactive ore. This operation is located in
Arkansas, an area of high annual rainfall. The mine area is
drained by two streams, which are considered as mine waste-
water and are treated via neutralization by ammonia. Part
of the wastewater is also treated by settling behind a series
of rock dams. Attempts are being made to segregate highly
contaminated leach streams from pyritic waste piles. These
leach streams are pumped to the mill water-treatment ponds
for neutralization and settling.
Table VII-26 presents the results of chemical analyses of raw
and treated mine wastewater at mine 6107. Neutralization
and settling treatment is employed at mine discharge 005, and
neutralization treatment alone is used at discharge 004.
The presence of ammonia in the effluents reflects the use
of ammonia for neutralization. Residual levels of iron and
manganese in effluent from discharge 005 are noteworthy.
VII-125
DRAFT
-------
DRAFT
TABLE VI1-25. CHEMICAL COMPOSITIONS OF RAW AND TREATED
MINE WATERS AT MINE 6104 (CLARIFLOCCULATOR
TREATMENT)
PARAMETER
PH
TSS
Oil and Grease
COD
As
Cd
Cu
Mn
Mo
Pb
V
Zn
Fe
Fluoride
CONCENTRATION (mg/£)
RAW WASTEWATER
6.5*
33.9
2
91.3
<0.07
<0.01
<0.02
0.21
<0.1
0.14
<0.5
0.05
1.51
0.52
TREATED WASTEWATER
7.8*
3.1
2.7
91.3
<0.07
<0.01
<0.02
0.03
<0.1
0.02
<0.5
0.03
0.12
0.46
Value in pH units
VII-126
DRAFT
-------
DRAFT
TABLE VI1-26. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTEWATERS AT MINE 6107
PARAMETER
Flow
TSS
TDS
Oil and Grease
COD
Ammonia
As
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
CONCENTRATION (mg/l)
DISCHARGE 005
RAW MINE WATER
15.000 m3/day
(4.300.000 gpd)
-
366
-
31
-
<0.07
< 0.005
<0.02
6.8
-
-
0.09
-
—
TREATED EFFLUENT
(NEUTRALIZATION & SETTLING) T
15.000 m3/day
(4.300.000 gpd)
30
285
< 0.1 00
< 0.010
0.25
3.6
<1
DISCHARGE 004*
TREATED EFFLUENT
(NEUTRALIZATION ONLY) T
5,000 m3/day
(1.400.000 gpd)
15
105
<1
5
10
0.01
<0.01
<0.01
0.94
<0.10
<0.01
018
<0.10
<1
Analysis of raw mine water unavailable for Discharge 004
Company data
VII-127
DRAFT
-------
DRAFT
Milling Operations. The ferroalloy-ore milling Industry has
been subcategorlzed on the basis of process used and size,
as described In Section IV. No exemplary operations were
visited which belong to the mill subcategory representing
operations processing less than 5,000 metric tons (5,500
short tons) per year. Operations representative of the
remaining milling subcategories provide examples of the
processes and all treatment options applicable to small
operations as well. Treatment technology currently practiced
is relatively uniform throughout the ferroalloy milling
Industry, although some examples of treatment for waste
constituents peculiar to particular subcategories have been
observed.
Commonly practiced treatment includes settling, neutralization,
and recycle of process water. In addition, sites visited
were observed to practice lime precipitation, distillation,
and air stripping.
Mill 6101. This operation is a flotation mill recovering
molybdenite concentrate on a large scale (approximately 14,000
metric tons, or 15,400 short tons) per day. Mill 6101 is
located in a mountainous area of New Mexico. Approximately
22,000 cubic meters (6,000,000 gallons) of water are used
in froth-flotation processing each day. No mine water is
produced, with process water being drawn from wells and a
nearby river. Ore processing consists of crushing, grinding,
and froth flotation. (See Section V.)
Treatment at mill 6101 utilizes tailing ponds and an
additional settling pond for removal of residual suspended
solids. Flocculants are added to the tailing stream, if
required for settling prior to discharge. Limited amounts
of water are reclaimed in thickeners at the mill site.
Because the mill circuit is mildly alkaline, lime is not
required to maintain neutral pH in the effluent stream.
Because the terrain near the mine and mill site did not allow
development of a sound tailing-disposal area, water-treatment
facilities are located at a significant distance (16 km,
or 10 miles) from the mill. Tailings are delivered to the
tailing ponds as a slurry, pumped through three 16-kilometer-
long (10-mile-long) steel pipelines, two of which are 25 cm
(10 in.) in diameter, and one of which is 30.5 cm (12 in.)
in diameter.
VII-128
DRAFT
-------
DRAFT
Because of abrasive wear on the pipe, it is necessary to
rotate and replace piping frequently. The use of end-of-line
monitors in the mill control room, a change to more abrasion-
resistant neothane-lined pipe, and a large tailing-disposal
maintenance staff have essentially eliminated problems with
recurrent spills of tailings from pipe breaks, which were
experienced in the past.
Three impoundments are used at mill 6101: two tailing ponds
totaling approximately 121 hectares (300 acres) in area,
and a secondary settling pond with a 1.6-hectare (4-acre) surface
area. The older of the two tailing ponds is nearly full and
partly revegetated. The second pond contains a water pool
of approximately 160 hectares (40 acres). Seepage through the
second dam is limited by use of an asphalt liner. Discharge
from the secondary settling pond flows through a small surface
channel to the final discharge point.
In addition to the tailing and settling ponds, construction
at the tailing-disposal site includes a diversion ditch and a.
flood-control dam to regulate drainage from a mountain, northeast
of the tailing ponds. These diversion structures are sealed to
protect the tailings area from the 100-year-frequency storm.
Water recycle from the tailing basin is rendered extremely
difficult at this plant by the large separation between the
mill and tailing area, although it is technically compatible
with the recovery practice.
Table VII-27 is a compilation of company chemical data for
intake and treated discharge waters. Table VII-28 presents
data for effluent treated using a tailing pond with secondary
settling. Raw-waste characteristics for mill 6101 were pre-
sented in Section V. The effectiveness of this treatment
scheme for suspended-solid removal is evident. The alkalinity
of the mill wastewater results in the effective removal of
most heavy metals in the tailing basins and settling pond.
Significant reductions of Cd, Cu, Fe, Mn, Pb, and Zn were
noted In this treatment scheme. Only total dissolved solids
are discharged at a level in excess of 0.1 kg/metric ton
(0.2 Ib/short ton) of ore milled.
Mill 6102. At this mill, molybdenite concentrates are
recovered by flotation. Byproduct concentrates of tin, tungsten,
monazite, and pyrlte are recovered in a complex system involving
gravity separation, froth flotation, and magnetic separation.
VII-129
DRAFT
-------
DRAFT
TABLE VII-27. ANALYSES OF INTAKE AND DISCHARGE WATERS FROM
MILL 6101 (COMPANY DATA)
PARAMETER
Alkalinity
BOD (Way)
COD
TDS
TSS
Hardnen
Ammonia (As N)
Nitrate
Phosphorus
Al
Sb
As
Ba
Be
B
Cd
Ca
Cn
Co
Cu
AVERAGE
CONCENTRATION
(ma/ H)
INTAKE
40
<30
<50
260
65
155
0.6
0.1
<0.01
0.24
<0.1
—
< 0.001
< 0.002
<0.1
< 0.002
103
<0.01
< 0.005
0.02
DISCHARGE
30
<30
<50
600
100
800
1.0
0.1
0.04
0.2
<0.1
—
< 0.001
< 0.002
<0.1
< 0.002
277
<0.01
< 0.005
0.02
PARAMETER
Fa
Pb
Mg
Mn
Ag
Mo
Ni
K
Se
Ag
Na
Sn
Ti
Zn
Sulfate
Chloride
Fluoride
Cyanide
Thiocyanate
AVERAGE
CONCENTRATION
(mg/£)
INTAKE
0.4
< 0.005
10
0.9
< 0.0001
0.01
0.02
1
< 0.005
< 0.001
3
< 0.01
< 0.08
0.05
100
2
0.2
—
^
DISCHARGE
0.16
< 0.005
30
0.9
< 0.0001
2
0.017
31
< 0.005
< 0.001
50
<0.01
<0.08
<0.06
1000
2
1.5
—
0.6
VII-130
DRAFT
-------
DRAFT
TABLE VI1-28. CHEMICAL COMPOSITION OF WASTE WATER AND WASTE LOADING
FOR MILL 6101
PARAMETER
TSS
TDS
Oil and Grease
COD
Total Cyanide
At
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
CONCENTRATION
(ing/ ft)
4.3
2.272
3
19.8
0.03
0.02
<0.01
<0.02
1.3
4.0
0.13
0.02
0.10
3.4
TOTAL WASTE
kg/day
73
39,000
51
340
0.51
0.34
< 0.2
< 0.3
22
68
2.2
0.34
1.7
58
Ib/day
160
86.000
112
760
1.1
0.76
<0.4
<0.7
48
150
4.8
0.75
3.7
130
WASTE LOAD
per unit ore milled
kg/1000 metric tons
5.2
2300
3.6
24
0.036
0.024
<0.01
< 0.02
1.6
4.9
0.16
0.024
0.12
4.1
lb/1 000 short tons
10
5,600
7.2
48
0.072
0.048
<0.03
< 0.04
3.2
9.8
0.32
0.048
0.24
8.2
VII-131
DRAFT
-------
DRAFT
Monazlte and pyrite concentrates are currently delivered to
the tailing impoundment for disposal; they are not shipped.
Ore processed is 39,000 metric tons (43,000 short tons) per
day. This mill is located in Colorado in a mountainous
area.
This operation uses water on a complete-recycle basis for
ten months of the year. During this period, due to consump-
tive losses in the mill, seepage losses, and evaporation from
tailing and water-storage ponds, the net water balance for
the system is negative. During the remaining two months
(usually May and June), heavy influx of water to the mill
tailing ponds from melting snow accumulations has necessitated
discharge of water from the system. The amount and duration
of this discharge have varied widely from year to year, de-
pending on meteorological conditions. The general flows of
water during normal operation and during purge periods are
presented schematically in Figure VII-17.
In addition to snow-melt Influx, water is drawn for the system
from a well and a small lake (domestic water supply), mine
drainage, and collection structures on a number of area
streams when needed. Diversion structures are currently
being greatly expanded and modified to provide diversion for
most of the area runoff around existing and new tailing ponds.
Drainage from a number of old mine workings (not owned by
the operator of mine 6102) to the tailing-disposal area has
complicated the diversion process. Drainage of low quality
is being segregated and channeled into the tailing ponds
rather than being diverted to the receiving stream. Water
leaves the system through consumptive losses in the mill,
evaporation from pond areas, seepage, and the aforementioned
discharge during peak runoff. With the completion of diver-
sion structures, discharge will be substantially reduced, or
possibly eliminated.
Within the water system, a complex pattern of pumping and
gravity flow is used to provide water treatment and recycle.
Three major impoundments, as well as a number of smaller
seepage and settling ponds, are currently involved.
A large man-made lake serves as the major holding basin for
water to be recycled to the mill. It receives decant water
from two active tailing ponds. From this lake, water is
pumped to two 7,570-cublc-meter (2,000,000-gallon) holding
tanks at the mill site.
VII-132
DRAFT
-------
DRAFT
Figure VII-17. WATER-FLOW SCHEMATIC DIAGRAM FOR MILL 6102
90.000 m3/day
(25.000.000 gpd)
RUNOFF
AVERAGE OF j
11.000 m3/d«v I
(3.100.000 gpd)
RANGE OF j
0 to 140,000 m3/day I
(0 to 38.500,000 gpd) '
MINE
WATER
\
3.600
(1.000
»
LEGEND
NORMAL-OPERATION FLOW
PURGE-WATER FLOW (INTERMITTENT)
VII-133
DRAFT
-------
DRAFT
Two mill tailing ponds, 303 hectares (750 acres) and 182
hectares (450 acres) In area, are Interconnected and also
connected to the mill water reservoir by a series of decant
structures.
Tailing ponds have not been treated with any deliberate sealant.
Seepage through the toe dam Is collected In seepage ponds and
pumped back up to the tailing ponds. The allowance of seepage
in this fashion is intended to limit hydrostatic pressures on
the dam and enhance safety.
Mine water is treated by lime-slurry addition in lagoons before
being pumped to the tailing pond and entry into the mill water
system. About 1,364 kg (3,000 Ib) per day of lime are con-
sumed in treating the average mine water flow of 3,600 cubic
meters/day (700 gpm).
Construction of a major new tailing pond is presently under-
way. This pond will have an area of 485 hectares (1,200
acres) and is expected to serve the mill for the next 35 to 50
years. Concurrent with this tailing-pond construction, a
number of supporting projects are underway, including develop-
ment of the extensive diversion structures mentioned previously.
Samples were collected at a number of points in the water-
management system. Since no discharge occurred at the time
of the site visit and sampling, analysis of these samples
does not provide direct measure of discharge quality. Table
VII-29 presents results of analyses of tailing-pond decant
water for the pond from which discharge occurs during spring
runoff and also shows the concentration of pollutants in
mill-storage process water after further settling. Table
VII-30 presents company data for discharge quality during
spring runoff and also shows calculated waste loads. Raw
waste characteristics and loading for mill 6102 are presented
in Section V.
Comparison of data in Tables VII-29 and VII-30 shows that
appreciably higher concentrations of many pollutants are
observed in the effluent streams during purge periods than
are found in the tailing ponds during normal operation.
This flushing effect—presumably, resulting from flows higher
than the design capacity of the treatment system—negates,
to a large extent, the benefits derived from recycle in terms
VII-134
DRAFT
-------
DRAFT
TABLE VI1-29. CHEMICAL COMPOSITION AND CALCULATED WASTE LOAD FOR
MILL 6102 TAILING-POND SURFACE WATER, WITH ANALYTICAL
DATA FOR MILL-RESERVOIR WATER
PARAMETER
TSS
TOS
Oil »nd Graas*
COO
At
Cd
Cu
Mn
Mo
Pb
V
Zn
Fa
Cyanlda
Fluoride
TAILING-POND SURFACE WATER
CONCENTRATION
Img/ 1 1
-
1340
0
11.9
0.01
<0.01
004
3.2
12 5
<002
<05
0.10
206
0.02
14.9
TOTAL WASTE
kg/day
-
175.000
0
1.070
0.90
< 0.90
3.6
288
3.600
< 1.8
<45
9.0
180
1.8
1.340
Ib/day
-
390.000
0
2.400
20
< 2
7.9
630
7.900
< 4
<100
20
400
4.0
2300
CALCULATED WASTE LOAD
par unit ore milled
k 9/1 000 metric tons
-
4.500
0
27
0.023
< 0.02
0.092
7.4
92
< 0.05
< 1
0.23
4.6
0.046
34
lb/1000 ihort toni
-
9,000
0
54
0046
<0.05
0.18
15
180
<0.09
<2
046
9.2
0.092
69
MILL-RESERVOIR
WATER
CONCENTRATION
(mg/£ I
14
1336
20
19.8
0.01
< 0.01
020
4.3
-
<0.02
< as
0.47
45
0.04
20
TABLE VI1-30. CHEMICAL COMPOSITION AND WASTE LOADING FOR DISCHARGE
AT MILL 6102 (COMPANY DATA)
PARAMETER
TSS
TDS
COD
Oil and Grease
Total Fe
Total Mn
Zn
Cd
Mo
Cu
Cyanide
Fluoride
CONCENTRATION
< 0.01
19.09
0.125
-
20.7
TOTAL WASTE
kg/day
1^00
18.000
230
10
110
48
6.4
<0.1
210
1.4
-
230
Ib/day
3,300
40,000
510
22
240
110
14
< 02
460
3.1
-
510
WASTE LOAD
per unit ore milled
kg/1000 metric tons
38
460
5.9
026
2.8
12
0.16
< 0.003
5.4
0.036
-
5.9
lb/1000 short tons
76
920
12
0.52
5.6
2.4
0.32
< 0.006
11
0.072
-
12
VII-135
DRAFT
-------
DRAFT
of removal of many pollutants. As a result, yearly average
effluent loads per ton of ore are, in most cases, comparable
to those achieved at mill 6101 without recycle from the
tailing pond. Significant advantage is seen in the recycle
system, however, in the removal of pollutants such as IDS, which
are not effectively removed by the standard alkaline preci-
pitation and settling treatment. Significantly greater advan-
tage is expected to be realized from the recycle system as
further development of diversion ditches appreciably decreases
the volume of purge flow, resulting in improvements in quality,
as well as decreased quantity, of effluent.
Mill 6106. This operation is engaged in the processing of
nickel ore (garnierite) to produce ferronickel. Mill 6106
is located in Oregon and processes approximately
4,535 metric tons (5,000 short tons) of ore per day. This
mill is representative of physical ore processors.
Water used in beneficiation and smelting of nickel ore at
mill 6106 is extensively recycled, both within the system and
from external water treatment. The bulk of the plant water
use is in the smelting operation, since wet-beneficiation
processes are not practiced. Water is used for ore-belt
washing, in scrubbers on ore driers, in cooling, and for slag
granulation. Water recycled within the process is treated in
two settling ponds, arranged in series. The first of these,
4.8 hectares (12 acres) in area, receives a process water
influx of 12.5 cubic meters (3,300 gallons) per minute, of
which 9.9 cubic meters (2,600 gallons) per minute are returned
to the process. Overflow to the 5.2-hectare (13-acre) second
pond amounts to 1.2 cubic meters (320 gallons) per minute.
This second pond also receives runoff water from the open-
pit mine site which is highly seasonal, amounting to zero
for approximately six months and reaching as high as 2,200
cubic meters (580,000 gallons) per day during the (winter)
rainy season. The lower pond has no surface discharge during
the dry season, inputs being balanced by evaporation and sub-
surface flow to a nearby creek. A sizeable discharge results
from runoff inputs during wet weather. Average discharge
volume over the year amounts to 460 cubic meters (120,000
gallons) per day.
This mill was visited during a period of zero discharge, and
samples collected reflect this condition. Samples were
collected from the influent to the first settling pond and
VII-136
DRAFT
-------
DRAFT
from its overflow, as well as from the surface waters of the
lower settling pond. Analytical data for the influent to the
treatment system are reported in Section V. Data for influent
to the second settling pond from the first pond, and for its
surface waters, are presented in Tables VII-31 and VII-32.
In general, the analyses of these samples were in agreement
with data furnished by the company for corresponding conditions.
In Table VII-33, average effluent loads based on company data
for the period of discharge are furnished. Since influent from
mine runoff could not be determined, no accurate measure of
treatment effectiveness is available. It is evident, however,
that effluent loads are quite low.
As Table VII-31 shows, the first settling pond alone is highly
effective in reducing concentrations of heavy metals in the
effluent stream. The recycle of substantial portions of the
process water delivered to this pond still further diminishes
the effluent load. The surface discharge from the second
settling pond is lower in most metals than the overflow from
the first pond, even though substantial mine runoff also enters
the second pond. The alkaline pH (average of 8.7) prevalent
in these basins enhances treatment effectiveness in retaining
heavy metals.
Mill 6107. At this operation, vanadium is recovered from
non-radioactive ore in a hydrometallurgical operation involving
salt roasting, leaching, solvent extraction, and precipitation.
Approximately 1,140 metric tons (1,250 short tons) of ore are
processed per day, requiring the use of 7,600 cubic meters
(1,900,000 gallons) of process water. At this operation,
representative of the leaching^mill subcategory, three distinct
mill wastewater streams are discharged.
Two of three effluents associated with mill 6107 contain
primarily noncontact water. One is primarily spring water
and natural drainage, with some seepage inputs from a process-
water reclamation pond and occasional spills of process water.
The other receives non-contact cooling water. Treatment of
these waste streams consists only of segregation from process
water and area runoff. Analytic data for these effluents are
presented in Tables VII-34 and VII-35.
The main wastewater stream from mill 6107 receives inputs from
several process units and air-pollution control devices, as
well as contaminated drainage from the mill area. Essentially
VII-137
DRAFT
-------
DRAFT
TABLE VII-31. CHEMICAL COMPOSITION AND TREATED WASTE LOADS FOR
OVERFLOW FROM FIRST SETTLING POND AT MILL 6106
PARAMETER
Cd
Co
Cu
F»
Mn
Ni
Pb
Zn
CONCENTRATION
(mg/£)
<0.01
<0.02
0.96
0.02
0.07
<0.1
0.03
TOTAL WASTE
kg/day
<0.02
<0.08
<0.03
1.4
0.03
0.11
<0.2
0.045
Ib/day
< 0.04
<0.02
< 0.4
0.099
WASTE LOAD
per unit ore milled
kg/1000 metric tans
< 0.004
<0.02
< 0.007
0.31
0.0066
0.024
<0.04
0.0099
lb/1000 thart tons
< 0.04
<0.01
0.62
0.013
0.048
<0.09
0.020
TABLE VII-32. CHARACTERISTICS OF SURFACE WATER FROM
SECOND SETTLING POND AT MILL 6106
PARAMETER
TSS
TDS
Oil and Grease
Cd
Cu
ft
Mn
Nl
Pb
Zn
COntCENTRATION
(mg/£)
63
184
27
< 0005
< 002
0.47
< 0.02
0.03
< 0.05
0.009
TOTAL WASTE
kg/day
29
85
1.2
<0002
<0009
0.22
<0009
0014
<002
0.0041
tb/dsy
6.4
187
2.6
< 0.004
<002
0.48
<0.02
0031
<004
00090
WASTE LOAD
POT unit or* milled
kg/1000 metric toni
064
187
026
< 00004
< 0.002
O.O4B
<0002
00031
<004
00009
lb/1000 short Mm
1 3
37
0.53
< 00009
<0004
0097
< 0.004
0.0062
<0.09
O.O018
per unit product
kg/1000 rrwtfic (am
36
1.000
14
C002
<0.1
26
< 0.1
0.18
< 2
0.05
lb/1000 short torn
69
2.000
29
< 005
<0i
036
<5
010
VII-138
DRAFT
-------
DRAFT
TABLE VI1-33. CHEMICAL COMPOSITION AND TREATED WASTE LOADS FROM
FINAL EFFLUENT FOR MINE/MILL 6106 DURING RAINY
SEASON (COMPANY DATA)
PARAMETER
TSS
TDS
Cu
Fe
Mn
Ni
Zn
CONCENTRATION*
img/ 9 )
30.8
166
0.003
0.12
0.007
0.038
0.006
TOTAL WASTE-
kg/day
14
76
0.0014
0.055
0.0032
0.017
0.0028
Ib/day
31
170
0.0031
0.12
0.0070
0.037
0.0062
WASTE LOAD
per unit ore processed*
kg/1000 metric tons
3.1
17
0.00031
0.012
0.0007
0.0037
0.00062
lb/1000 short tons
6.2
34
0.00062
0.024
0.0014
0.0074
0.0012
Approximate average for periods of discharge
•Yearly averages
TABLE VI1-34. CHEMICAL COMPOSITION AND WASTE LOADING FROM
AREA RUNOFF AND RECLAMATION-POND SEEPAGE
AT MILL 6107 (COMPANY DATA)
PARAMETER
pH
TSS
TDS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Cu
Mn
Mo
Pb
Zn
Fa
Fluoride
CONCENTRATION
(mg/jn
6.4»
10
1.705
<1
6
1.0
0.02
<0.01
< 0.01
<0.01
5.8
< 0.1
<0.01
0.04
< 0.1
< 1
TOTAL WASTE
kg/day
-
52
8.900
< 5
31
5.2
0.10
< 0.05
< 0.05
< 0.05
30
< 0.5
< 0.05
0.21
< 0.5
< 5
Ib/day
-
104
18,000
<10
62
10.4
021
< 0.1
< 0.1
< 0.1
60
< 1
< 0.1
0.42
< 1
<10
WASTE LOAD
per unit ore milled
kg/1000 metric tons
-
46
7.800
< 4
27
4.6
0.088
< 0.04
< 0.04
< 0.04
26
< 0.4
< 0.04
0.19
< 0.4
< 4
lb/1000 short tons
-
92
16.000
< 9
54
9.2
0.18
< 0.09
< 0.09
< 0.09
53
< 0.9
< 0.09
0.38
< 0.9
< 9
•Value in pH units
VII-139
DRAFT
-------
DRAFT
TABLE VII-35. CHEMICAL COMPOSITION AND WASTE LOADING FOR
COOLING WATER EFFLUENT AT MILL 6107
(COMPANY DATA)
PARAMETER
pH
TSS
TDS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Cu
Mn
Mo
Pb
Zn
Fa
Fluoride
CONCENTRATION
(mg/£>
7.2*
20
695
< 1
IB
10
0.010
< 0.01
<0.01
< 0.01
0.54
<0.10
< 0.01
0.18
< 0.10
< 1
TOTAL WASTE
kg/day
-
42
1.500
< 2
32
21
0.021
< 0.02
< 0.02
<0.02
1.1
< 02
< 02
<2
Ib/day
-
82
3.300
< 4
70
46
0.046
< 0.04
< 0.04
<0.04
2.4
<0.4
<0.04
0.84
< 0.4
< 4
WASTE LOAD
per unit ore milted
kg/1000 metric tons
-
37
1.300
< 2
2B
18
0.018
<0.02
<0.02
<0.02
0.97
< 0.2
< 0.02
0.34
< 0.2
< 2
lb/1 000 short toru
-
74
2,600
< 4
56
38
0.036
< 0.04
< 0.04
< 0.04
1.9
< 0.4
< 0.04
0.67
< 0.4
< 4
•Value in pH units
VII-140
DRAFT
-------
DRAFT
all streams entering this waste stream bear very high con-
centrations of dissolved salts, as well as a variety of other
contaminants, including ammonia and various heavy metals.
The complex system of inputs and treatment and holding ponds
feeding this discharge is illustrated in Section V. The
main process effluent from washing, leaching, and solvent
extraction is treated by ammonia addition prior to discharge
to a 5.3-hectare (13-acre) holding pond, where it is joined
by scrubber bleed water from ore dryers and treated sanitary
wastewater, both of which have first been treated for solids
removal in a holding pond. Bleed water from a roaster/scrubber
is treated by settling in a primary pond before delivery to
a 2.8-hectare (7-acre) holding pond, adjacent to that containing
process effluent. Discharge from these two ponds is staged
to avoid the formation of calcium sulfate precipitates, which
would result from their combination. Further, discharge is
adjusted by impoundment in accordance with flow in the receiving
water to comply with permit stipulations on the maximum allow-
able salinity increase in the receiving water (25 mg/1). The
volume of this effluent is limited somewhat by recycle of
water from the tailing pond to the washing circuit, recycle
within the solvent-extraction/precipitation operation, and
recycle of scrubbing water to the greatest extent practical.
In general, further reuse of water is limited by the extremely
high concentrations of dissolved solids in the effluent water.
Data for the process wastewater after ammonia treatment, and
for the drier scrubber bleed after solids removal, are pre-
sented in Tables VII-36 and VII-37. The two waste streams are
combined in one holding pond for staged discharge. Since this
pond was not discharging during sampling, only company data
are presented in Table VII-38.
Table VII-39 presents data for treated effluent from the holding
pond receiving wastewater from roaster/scrubbers after primary
settling. Table VII-40 presents additional company data for
the same discharge. Average characteristics of total process
effluent (company data) are presented in Table VII-A1.
Mill 6104. At mill 6104, a complex ore is processed by flo-
tation and leaching operations to yield molybdenum and copper
concentrates and ammonium paratungstate. The mill is located
in California. Mill wastewater is
treated by lime addition to a pH of 9.5 and subsequent
impoundment in a tailing pond, from which clarified
VII-141
DRAFT
-------
DRAFT
TABLE VI1-36. CHEMICAL COMPOSITION AND WASTE LOADING FOR
PROCESS EFFLUENT AFTER AMMONIA TREATMENT
AT MILL 6107
PARAMETER
PH
TDS
Oil and Grease
COD
As
Cd
Cr
Cu
ivtn
Mo
Pb
V
Zn
Fe
Fluoride
CONCENTRATION
(mg/£)
IN WASTEWATER
8.8»
40.284
5
443
0.13
0.039
0.2
0.13
52
< 0.1
<0.05
31.5
0.47
0.3
4.55
TOTAL WASTE
kg/day
—
85.000
11
930
0.27
0.082
0.42
0.27
109
<0.2
<0.1
66
0.99
0.63
9.6
Ib/day
—
190,000
24
2.000
0.59
0.18
0.92
0.59
240
<0.4
<0.2
145
2.2
1.4
21
WASTE LOAD
per unit ore processed
kg/1000
metric tons
—
75,000
10
820
0.24
0.072
0.37
0.24
96
<0.2
<0.1
58
0.87
0.56
8.5
lb/1000
short tons
—
150.000
20
1,640
0.48
0.144
0.74
0.48
192
<0.4
< 0.2
116
1.74
1.12
17
•Value in pH units
VII-142
DRAFT
-------
DRAFT
TABLE VII-37. CHEMICAL COMPOSITION AND WASTE LOADING FOR
DRIER SCRUBBER BLEED WATER AFTER SETTLING
TREATMENT AT MILL 6107
PARAMETER
PH
IDS
Oil and Grease
COD
As
Gd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Fluoride
CONCENTRATION
(mg/£)
IN WASTEWATER
7.7"
10.852
3
34.27
<0.07
< 0.005
0.1
0.08
13.0
<0.1
<0.05
37.5
0.17
0.75
1.2
TOTAL WASTE
kg/day
—
10,000
2.8
32
<0.07
<0.05
0.094
0.075
12
<0.09
<0.05
35
0.16
0.71
1.1
Ib/day
-
22.000
6.2
70
<0.15
<0.1
0.21
0.17
26
<0.2
<0.1
77
0.35
1.6
2.4
WASTE LOAD
per unit ore processed
kg/1000
metric tons
-
8,800
2.5
28
<0.06
< 0.004
0.083
0.066
11
<0.08
<0.04
31
0.14
0.63
0.97
lb/1000
short tons
-
16.600
5
56
<0.12
< 0.008
0.166
0.122
22
<0.16
<0.08
62
0.28
1.26
1.94
'Value in pH units
VII-143
DRAFT
-------
DRAFT
TABLE VI1-38. CHEMICAL COMPOSITION AND WASTE LOADING FOR
HOLDING-POND EFFLUENT (PROCESS WATER AND
DRIER SCRUBBER BLEED) AT MILL 6107
(COMPANY DATA)
PARAMETER
Ammonia
Ca
Cd
Cu
Mn
Mo
V
Zn
Ni
Fe
Sulfate
Chloride
CONCENTRATION
(mg/4)
IN WASTEWATER
2,030
450
0.08
0.23
38
16
31
0.83
0.96
0.23
12.200
7,800
TOTAL WASTE
kg/day
6.500
1.400
0.26
0.73
120
51
99
2.7
3.1
0.73
39,000
25,000
Ib/day
14,000
3,100
0.57
1.6
260
110
220
5.9
6.8
1.6
86.000
55,000
WASTE LOAD
per unit ore processed
kg/1000
metric tons
5,600
1.200
0.23
0.64
110
45
87
2.4
2.7
0.64
34,000
22.000
lb/1000
short tons
11,200
2.400
0.46
1.28
220
90
174
4.8
5.4
1.28
68,000
44.000
VII-144
DRAFT
-------
DRAFT
TABLE VII-39. CHEMICAL COMPOSITION AND WASTE LOADING FOR
ROASTER SCRUBBER BLEED WATER AFTER SETTLING
AT MILL 6107
PARAMETER
pH
TSS
TDS
Oil and Grease
COD
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Fluoride
CONCENTRATION
(mg/A)
IN WASTEWATER
7.9"
121
57,690
3
1.859
< 0.07
< 0.005
0.2
< 0.03
5.5
< 0.1
<0.05
15
5.95
0.25
6.0
TOTAL WASTE
kg/day
—
209
100,000
5.2
3.200
<0.1
< 0.009
0.35
<0.05
9.5
<0.2
<0.09
26
10
0.43
10
Ib/day
—
460
220,000
11
7,000
< 0.3
< 0.02
tt 77
< 0.1
21
< 0.4
< 0.2
57
23
0.95
23
WASTE LOAD
per unit ore processed
kg/1000
metric tons
—
180
88,000
4.6
2,800
<0.09
< 0.008
0.31
<0.04
8
<0.2
<0.08
23
8.8
0.38
8.8
lb/1000
short tons
-
360
176,000
9.2
5,600
<0.18
< 0.016
0.62
<0.08
16
<0.4
<0.16
46
17.6
0.76
17.6
'Value in pH units
VII-145
DRAFT
-------
DRAFT
TABLE VI1-40. CHEMICAL COMPOSITION AND WASTE LOADING FOR
ROASTER SCRUBBER BLEED WATER AFTER SETTLING
AT MILL 6107 (COMPANY DATA)
PARAMETER
Ammonia
Ca
Cd
Cu
Mn
Mo
V
Zn
Ni
Fa
Sulfate
Chloride
CONCENTRATION
(mg/£)
IN WASTEWATER
360
26.000
0.42
0.31
11
1.1
14
8.4
1.0
0.93
500
36.000
TOTAL WASTE
kg/day
620
45.000
0.73
0.54
19
1.9
24
15
1.7
1.6
820
62,000
Ib/day
1,400
99,000
1.6
1.2
42
4.2
53
33
3.7
3.5
1,900
140.000
WASTE LOAD
per unit ore processed
kg/1000
metric tons
550
40,000
0.64
0.48
17
1.7
21
13
1.5
1.4
760
55,000
lb/1000
short tons
1,100
80,000
1.28
0.96
34
3.4
42
26
3.0
2.8
1/120
110,000
VII-146
DRAFT
-------
DRAFT
TABLE VI1-41. CHEMICAL COMPOSITION AND WASTE LOADING FOR
AVERAGE TOTAL PROCESS EFFLUENT AT
MILL 6107 (COMPANY DATA)
PARAMETER
pH
TSS
TDS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
CONCENTRATION
(mg/£)
IN WASTEWATER
6.7»
180
44,000
<1
70
1.200
0.020
0.30
0.090
0.26
28
11
<0.1
4.00
0.50
1
TOTAL WASTE
kg/day
-
890
220,000
<5.0
340
5,900
0.098
1.5
0.44
1.3
140
54
<0.5
20
2.5
4.9
Ib/day
-
2,000
480,000
<10
750
13.000
0.22
3.3
0.97
2.9
310
120
< 1
44
5.5
11
WASTE LOAD
per unit ore processed
kg/1000
metric tons
-
780
190,000
< 4
300
5.200
0.09
1.3
0.39
1.1
120
48
<0.4
18
2.2
4.3
lb/1000
short tons
-
1,560
380,000
< 8
600
10,400
0.18
2.6
0.78
2.2
240
96
< 0.8
36
4.4
8.6
•Value in pH units
VII-147
DRAFT
-------
DRAFT
water exits by percolation and evaporation. Treatment practiced
on segregated waste streams from the leaching and solvent-
extraction processes is representative of advanced treatment
applicable to leaching operations. Waste streams from chemical
processing of scheelite flotation concentrates are treated
by distillation in a two-stage evaporator/crystallizer and by
stripping with air for ammonia removal prior to combination
with tails from other operations for liming and delivery to
the tailing ponds.
Samples of the solvent-extraction effluent and the precipitation
waste before treatment were not obtained. Since there was no
surface discharge, and since there was no pool of water in
the tailing pond at the time of the visit to this site, no
sample of clarified mill discharge water could be obtained.
Limitations met by this discharge may be assumed to be indica-
tive of its quality and are tabulated below.
Concentration
Parameter (mg/1)
Sodium 600
Chloride 1000
Sulfate 1000
Total Nitrogen 5
(Organic, NH3., W03)
Nitrate 2
These values are consistent with the observed 2,290 mg/1
TDS content of the combined tailing stream (See Waste Character-
istics, Section V), reflecting the substantial removal of
dissolved salts—especially, sodium sulfate—from the effluent.
Mercury Ores
Historically, water has found little use in the mercury-ore
mining and dressing industry. In the past, the mined ore
was primarily fed directly into a retort or furnace, and the
mercury was recovered by roasting. When beneficiation has
been employed, it has normally been limited to crushing and/or
grinding. As a result, water-treatment technology or facilities
have not been typically required in this industry.
VII-148
DRAFT
-------
DRAFT
Mining Operations. Water is not used in mercury mining
operations and is discharged where it accumulates. When mines
are not located adjacent to a mill, or when their effluents
(if any) are to be segregated from the mill wastewater, it
will be necessary to discharge these waters, unless total
impoundment is possible. Treatment of this wastewater is
necessary for removal of suspended solids and heavy metals.
The mercury ion is best treated for removal by sulfide pre-
cipitation. Other technologies for the removal of heavy-
metal waste constituents are the chemical precipitation and/or
flocculation methods and settling ponds, which have been
discussed previously in this section.
Milling Operations. Mercury ore can be concentrated by
gravity methods and by froth flotation. However, these methods
have not been employed widely, since direct retorting of the
ore is an efficient and effective method for recovering mer-
cury. In addition, most mercury ores are not amenable to
gravity separation, since mercury minerals tend to be crushed
finer than the gangue, with resultant excessive lo_ss of these
minerals in the slimes. However, as lower-grade mercury ores
become mined, it is expected that beneficiation processes
will become increasingly important and necessary in this
industry.
Mill 9201. This operation is located in the state
of California. Operation of this mill is
seasonal, with closure of the mine/mill during the rainy
season (winter), when muddy roads make access difficult.
A sandstone ore containing cinnabar (HgS) is mined from an
open pit and brought to the mill. During 1973, 30,000 short
tons (27,210 metric tons) of ore were milled by gravity methods
to produce a cinnabar concentrate. No discharge results from
the mine.
This mill operates on a total-recycle system, with no discharge
resulting. Water is used in a gravity-separation process,
and the mill tailings are discharged at a rate of 1,665 cubic
meters (436,000 gallons) per day to a 1-hectare (2.5-acre)
tailing pond. Seperan NP-10, a flocculant, is added to the
waste stream to increase solids settling. Clarified pond water
is decanted and returned to the mill for reuse. About 16
cubic meters (4,300 gallons) per day of makeup water are
VII-149
DRAFT
-------
DRAFT
required, and this is obtained from a nearby reservoir.
The efficiency of the treatment system is presented in Table
VII-42. No waste loadings have been computed, because no
discharge results from this operation.
Mill 9902. This operation is located in Nevada.
Although this mill is not yet active, it is due
to begin operation during 1975. Mercury ore, cinnabar,
will be concentrated by froth flotation.
This mill proposes to employ a recycle system; should this
type of operation pose problems, the ponding area will be
increased, and a combination of impoundment and evaporation
will be used. Presently, this operation plans to employ two
20-hectare (50-acre) ponds if recycling is used, and four
20-hectare (50-acre) ponds if impoundment is required. As
a result, no discharge is expected to result.
Uranium, Radium, and Vanadium Ores
The discussion that follows describes treatment and control
technology in current use in the uranium, radium, and vana-
dium (byproduct recovery under NRC licensing) ore mining
and dressing industry. Aspects of treatment and control
which are characteristic of this category are described.
Mining Operations. Uranium mining in the U.S. is conducted
primarily in the arid states. Relatively few mines in this
industry discharge any water. Where it is practical, mine
wastewater is used as process feed water for milling. It
then becomes a mill effluent and is impounded, and subsequently
is lost to evaporation and seepage or percolation. At the
operations employing the best treatment and control technology
in this industry, uranium values are frequently extracted
from minewater by ion exchange (IX) methods. In addition,
where dry mines are proximate to mines discharging wastewater,
the discharge is often recycled to the dry mines to effect
in-aitu leaching. Evaporation and seepage losses in this
process often reduce water volume to a point where no discharge
results. Further treatment of wastewater destined for
natural waterways always includes settling.
High values of Ra226 observed in mine wastewater indicate
that coprecipitation treatment is necessary to reduce radium
VII-150
DRAFT
-------
DRAFT
TABLE VII-42. CHEMICAL COMPOSITIONS OF MILL WASTEWATER AND
TAILING-POND SURFACE WATER AFTER TREATMENT
AT MINE/MILL 9201 (NO DISCHARGE, RECYCLE OF-
TREATED WATER)
PARAMETER
PH
TSS
TDS
COD
Oil and Grease
SiO2
Al
Cd
Cr
Cu
Total Fe
Pb
Total Mn
Hg
Ni
Sr
Zn
Sb
Mo
Fluoride
Sulfate
CONCENTRATION <1
9.8
10.4
< 0.005
0.04
<0.02
<0.5
<0.1
50.0
-
0.68
0.60
0.14
<0.5
<0.2
0.61
100
TAILING-POND
DECANT
65*
76
144
2723
2
9.3
0.5
< 0.005
0.02
<0.02
0.87
<0.1
0.10
0.125
0.10
0.10
0.03
<05
-------
DRAFT
values to acceptable values. Values of Ra226 In mine waste-
water currently range from approximately 100 to 400 picocuries
per liter, while technology currently being employed in mill
wastewater treatment nearly always attains reduction to a
level of below 3 picocuries per liter; under favorable con-
ditions existing in well-designed treatment systems, levels
of 1 picocurie per liter have been obtained.
To employ treatment technology recommended here for radium
reduction, in mine wastewaters, it may sometimes be necessary
to add sulfate ion to the wastewater stream to allow coprecipi-
tation with barium chloride. If ferrous sulfate is added
at a level of 100 mg/1, some molybdenum is also coprecipitated
with ferric hydroxide, and sulfate ion is liberated to effect
radium coprecipitation.
Mine 9401. This operation currently obtains uranium ore
from four underground mines in New Mexico, one of which
contributes a significant amount of mine water to adjacent
mines after treatment by ion exchange (for uranium extraction)
for in-situ leaching. The total flow in the ion-exchange
plant is 9,300 cubic meters (2,455,200 gallons) per day.
Evaporation losses in surface distribution channels apparently
cancel the excess influx from the one wet mine, so no dis-
charge results. If there were a discharge from the ion-
exchange system, this discharge would exhibit high levels of
suspended solids (530 mg/1) and COD (750 mg/1).
The ion-exchange process at this operation illustrates that
an IX system which is optimized for one particular ion (e.g.,
uranyltrisulfuric ion) is relatively ineffective for removing
even similar ions. As shown in the table below, only vana-
dium follows uranium in being extracted.
Element In Out
U 25 1
As 0.03 0.04
Pb 0.02 0.11
V 1.0 less than 0.5
Fe 0.47 0.51
Mo 0.5 0.77
Be 0.01 0.01
Al 0.55 0.55
B 0.15 0.19
VII-152
DRAFT
-------
DRAFT
Element In Out
Ca 93 96
Mg 45 45
K 25 25
Na 200 200
Sr 0.87 0.124
Zn 0.034 0.064
However, in this case, uranium and vanadium are reduced to
levels of 1 mg/1 or less. With some compromises in treatment
efficiency for uranium and vanadium, other metals can be re-
moved .
Mine 9402. A group of several mines discharging 11,500
cubic meters (3,036,000 gallons) of water per day is located
near a mill which uses approximately two-thirds of the mine
discharge as mine process water. This operation is also
located in New Mexico. Two types of treatment are used.
At one mine, mine water is treated for suspended-solid removal
by a series of three settling ponds and then is discharged.
Table VII-43 presents the chemical compositions of raw and
treated wastewaters resulting from this mine.
A second group of mines feeds a treatment system consisting
of an ion-exchange plant (for removal of uranium values)
and a large settling pond.
Initial concentrations varying from 2 to 12 mg/1 of U_3_OJ[ were
treated by use of an eight-column anlonic-exchange system,
which recovers 98 percent of the influent uranium. At lower
concentrations, this process is known to be less effective
than at higher concentration. An example of this loss effi-
ciency can be cited for the case of mercury removal by ion-
exchange methods, as shown in Figure VII-18. The fact that
uranium shows a similar behavior is illustrated by the data
points for uranium that have also been plotted on this graph.
Additional data on the efficiency of IX processes are avail-
able to members of the industry. These data currently are
proprietary, for competitive reasons.
Table VII-44 presents the results of treatment by ion exchange
and settling at mine 9402(002).
VII-153
DRAFT
-------
DRAFT
TABLE VII-43. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTEWATERS AT MINE 9402 (001)
PARAMETER
PH
TSS
COD
TOC
As
Cd
Cr
Cu
Hg
Mo
Ni
Pb
V
Zn
Ra
U
CONCENTRATION (mg/£)
RAW WASTE WATER
8.1*
289
<10
45
0.02
—
—
<0£
—
0.5
-
0.13
2.1
—
230"
4.14
TREATED WASTEWATER
7.4*
17
15.9
19.5
0.02
0.003*
0.01 *
0 to 0.01 f
0.001 f
0.8
0.04*
0.1
1.7
0.002f
65"
1.1
Value in pH units Company data Value in picocuries/£
Figure VII-18. ION EXCHANGE FOR MERCURY AND URANIUM AT LOW
LOADINGS AND CONCENTRATIONS
0.1
1.0 10 100
EQUILIBRIUM CONCENTRATION (mg/£)
VII-154
1000
10,000
DRAFT
-------
DRAFT
TABLE VII-44. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTEWATERS AT MINE 9402 (002)
PARAMETER
PH
TSS
COD
TOC
As
Cd
Cr
Cu
Hg
Mo
Mi
Pb
V
Zn
Ra
Th
U
CONCENTRATION (mg/l)
RAW WASTE WATER
7.7*
—
734
20.5
<0.01
<0.02
<0.02
<0.5
0.0004
0.5
<0.01
0.18
<0.5
<0.5
3.260**
<0.1
13.31
DISCHARGE FROM
TREATMENT (IX)
8.1*
-
405
20.5
0.02
<0.02
<0.02
<0.5
0.0004
0.1
<0.01
0.11
<0.5
<05
105"
<0.1
4.55
DISCHARGE FOLLOWING
SETTLING BASIN *
8.4*
20 to 30
1.6
-
-
0.002
0.01
0.01
0.001
0.9
0.04
0.05
-
0.022
370"*
-
1
Company data
^Value in pH units
Value in picocuries/£
VII-155
DRAFT
-------
DRAFT
Milling Operations. Treatment for suspended-solid removal,
neutralization of pH, precipitation of hazardous pollutants,
coprecipitation of parameters in very low concentrations, and
for the recovery of values exists in milling operations of
the uranium industry. Some treatment is used to permit dis-
charge, while, in most instances, treatment facilitates recycle
and/or impoundment. Approximately 90 percent of the uranium
milling industry has no point discharges. Two of the remain-
ing milling operations have lateral seepage from tailing
impoundments that is collected and discharged. One operation
is currently modifying its entire process to attain
zero discharge. This is expected to be accomplished
by increased use of recycling and by minor process
modifications.
Mill 9401. This operation is located in New Mexico and
extracts uranium and vanadium byproducts by alkaline leaching
processes. (See Section III.) The mill incorporates two
recycle loops: one Involving recarbonization of leach, which
leaves all water characteristics relevant to discharges
essentially unchanged, and another loop that returns decant
water from tailings by means of an ion-exchange column.
The IX process recovers uranium that was rejected to tailings
and solubilized there; however, this loop also does not
improve water quality.
As discussed in Section III, the alkaline-leach process used
at this mill involves a purification step that adds sodium
and sulfate Ions to the water. If water were recycled indefin-
itely, these ions would Increase in the tailing ponds. Evapo-
ration there would eventually permit crystallization of sodium
sulfate, and the formation of crystals in other portions of
the loop would prevent the use of the recycle liquor, even
for such operations as repulping of tailings.
Some of the saline liquor apparently disappears by seepage,
as is also suggested by slightly elevated sulfate levels
reported for test wells near the tailing area. The existence
of seepage has allowed indefinite recycling of a significant
portion of the flow through the plant. The mill has no point
discharge.
If seepage is controlled, measures must be taken to separate
sodium sulfate resulting from the purification process from
the other recyclable liquors. A separate, lined evaporation
pond would serve this function.
VII-156
DRAFT
-------
DRAFT
Mill 9402. The mines and mill are located near each other
at this operation in New Mexico, and some water from the
mines is used in the acid-leach process, while the remainder
is discharged. The mill itself has no point discharge.
Like most acid-leach operations, the mill cannot practice
recycle from tailing decant liquor because high concentrations
of solutes interfere with the process of concentrating values.
The effect of evaporation on the tailings that are pumped
through a sequence of four sequential ponds is illustrated
in Figure VII-19. If vertical seepage or discharge were to
result from this operation, neutralization of the acid waste
liquors to prevent discharge of innocuous salts and resolu-
bilized heavy metals would be necessary.
Lateral seepage from the first tailing pond is controlled
by pumping from a second seepage collection "pond," at the
toe of the dam, to safer storage in a third pond, which is
at a higher elevation than the first tailing pond. From there,
water may be pumped to one of two smaller ponds at even higher
elevation. This arrangement of ponds provides protection
against failure of any one dam, except for the main tailing dam.
Failure of the dams retaining the upper ponds would dump
their contents into the larger, lower ponds, rather than
into the environment.
Mill 9403. This mill is located in Utah. Mines supplying
this operation are completely separated from the mill and
were not visited. The mill uses alkaline leach and has exten-
sive byproduct operations. Its discharge to a river is
expected to be reduced in volume by a factor of ten or
eliminated in late 1975. Complete recycle is technically possible
but would require expensive alterations to waste-treatment
facilities. Land suitable for construction of a pond large
enough to remove waste liquor by evaporation and seepage is
several kilometers away and is located at an elevation several
hundred meters higher.
The present mill treats river water (to reduce hardness),
raw wastewaters (to remove suspended and settleable solids),
and decant water from the tailing pond (to remove radium by
BaC12_ coprecipitation). The water-softening scheme is not
properly an effluent treatment, but it illustrates a large-
VII-157
DRAFT
-------
DRAFT
Figure VII-19. CHEMICAL CHANGES IN A SEQUENCE OF TAILING
IMPOUNDMENTS AT MILL 9402
3.0
ALL HEAVY-METAL CONCENTRATIONS NORMALIZED TO RAW MILL WASTEWAT R
2.0 •
cc
z
UJ
o
8
o
UJ
N
oc
o
INFLUENT
POND-1
EFFLUENT
POND-2
EFFLUENT
LOCATION
POND-3 POND-5
EFFLUENT EFFLUENT
VII-158
DRAFT
-------
DRAFT
scale technique for reducing calcium and magnesium, by reducing
calcium carbonate content from approximately 500 mg/1 to
35 mg/1. Table VII-45 shows the effect of tailing-pond and
coprecipitation treatments on effluent characteristics.
Mill 9404. This mill located in New Mexico is located approx-
imately 100 km (60 miles) from the mine that furnishes ore.
The mill uses acid leaching, and recycle is not practical.
A tailing pond, 3 kilometers (2 miles) from the mill, evaporates
wastewater and concentrates the solutes. The tailing area
covers a somewhat porous stratum. For this reason, a deep
well was drilled to a depth of 770 meters (2,530 feet) Into
porous strata containing water unfit for other use, and
decant wastewater from the pond is occasionally injected
into this well, following filtering to remove suspended solids
that might plug the well. There is no point discharge at
this mill.
Metal Ores. Not Elsewhere Classified
This group contains ore mining and dressing operations which
vary considerably in their size, methods of mining and bene-
ficiation, and location. Relatively few operations are
represented in this diverse group, with primary production
for antimony, beryllium, platinum, and rare-earth ores repre-
sented by one mine and mill each. Tin and zirconium ores
are obtained as byproducts, while antimony is also obtained
as a byproduct of both silver mining and milling and lead and
zinc smelting.
Antimony Ores
There currently exists only one operation (mine/mill 9901)
which is mining and milling ore primarily for its antimony
content. Mill 9901 discharges tailings from its flotation
circuit to a tailing pond and achieves zero discharge by
Impoundment of tailings in this pond. The operators of this
mill also indicate that recycling of tailing-pond process
water would not be expected to pose any problems, should
recycling become desirable at this mill. However, if this
water were to be recycled, additional settling treatment
would be necessary to reduce its slime content. Therefore,
the impoundment area would require either expansion or
redesign to facilitate a recycle system.
VII-159
DRAFT
-------
DRAFT
TABLE VI1-45. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTEWATERS AND EFFLUENT WASTE LOADING
AT MILL 9403 (SETTLING AND BaCI2 COPRECIPITATION)
PARAMETER
pH
TSS
COD
TOC
As
Cd
Cr
Cu
Hg
Mo
Ni
Pb
V
Ra
Th
U
CONCENTRATION (mg/H)
RAW
WASTEWATER
9"
111.000
27.8
<1
1.4
0.04
<0.02
1.1
0.0016
0.25
0.52
0.69
<0.5
111*
_
3.9
TREATED
EFFLUENT
9«
31
71.4
20
2.8
<0.02
<0.02
<0.5
0.0002
3.3
<0.01
0.13
7.4
4.09f
<0.1
2.5
EFFLUENT WASTE LOAD
kg/day
—
161
370
100
15
<0.1
<0.1
<2.6
0.001
17
<0.05
0.67
38
21.2"
<0.5
13
Ib/day
—
354
814
220
33
0.22
0.22
<5.7
0.0022
37
<0.11
1.47
84
-
<1.1
29
kg/metric ton
of concentvttu
-
120
270
74
10
-
-
-
0.0007
12
-
OM
30
15.8ft
—
10
Ib/short ton
of concontrstB
-
240
540
148
20
-
-
-
0.0014
24
-
0.96
60
14.4""
-
21
•Value in pH units
Value in picocuries/£
"Value in microcurim/day
*Value in microcuries/metric ton
•Value in microcuries/short ton
VII-160
DRAFT
-------
DRAFT
No effluents are currently being discharged to the surface
from mine 9901. However, this operation has been active
for only a few (three to five) years; as the. mine is developed
more extensively, a discharge may result from the influx of
ground water. If discharged, the mine wastewater may
potentially contain suspended solids and solubilized metals,
which will require treatment prior to final discharge of the
effluent. Treatment technologies potentially available for
application at this mine are chemical precipitation and
flocculation methods and use of settling basins, previously
discussed.
Chemical precipitation of metal hydroxides by lime addition
will successfully remove most of the heavy metals (i.e.,
arsenic and zinc) present in this ore body. Lime will also
create the alkaline conditions necessary for the successful
removal of antimony by sulflde precipitation.
Beryllium Ores
Only one operation in the beryllium mining and milling industry
is known to use water in a milling process. The limited amount
of beryl mined domestically is, for the most part, concentrated
by crude hand-cobbing methods. However, bertrandite, mined
from an open pit, is processed at mill 9902 by a sulfuric
acid leach process. This mill is achieving zero discharge
by impoundment of the mill tailings in a tailing pond. Water
is removed from the pond by natural evaporation and possible
percolation into the subsurface. No discharge exists from
the open-pit mine at this time.
Platinum-Group Metals
The bulk of production of the platinum-group metals results
from recovery as byproducts from copper ore during refining
operations. These metals are also being recovered by an
operation (mine/mill 9904) which seasonally mines a placer
deposit in Alaska. This placer, located alongside a major
river, is mined by a dredge, floated on a impoundment con-
structed over the deposit. The heavy minerals are concentrated
by gravity-separation methods; therefore, waste loading of
the process water includes primarily suspended solids. These
process wastes are discharged to the dredge pond, where some
VII-161
DRAFT
-------
DRAFT
settling of the solids occurs. The suspended-solid content
of the pond water is further reduced as it filters through
a sand barrier prior to final discharge.
The relatively unsophisticated methods described above are
typical of the best existing treatment at precious-metal
placer operations. As such, this treatment is designed to
reduce suspended-solid loadings of final discharges. Since
recycle is usually not practicable at a placer operation of
this type, use of the treatment described is necessary.
Therefore, efficient treatment can be maximized by optimizing
conditions for settling and/or filtration of the process
wastes. Long-range control of solids should take the location
of the treatment facilities into consideration. These facilities
should, when possible, be located at a distance from a stream,
which would afford protection from seasonally high waters.
Table VII-46 presents the chemical composition and treated
waste load for mine/mill 9904.
Rare-Earth Ores
Currently, only one operation mines a lode deposit for its
rare-earth mineral content. This operation (mine 9903) mines
bastnaesite from an open pit and concentrates the ore in a
flotation circuit. The flotation concentrate is further
upgraded in a leach circuit before final processing in a
solvent-extraction plant. Presently, the flotation tailings
are discharged to a tailing pond, and the clarified pond
water is recycled back into the flotation circuit. Process
wastes from the leach circuit are separately discharged to
an evaporation pond. The efficiency of tailing-pond treatment
of the water to be recycled is presented in Table VII-47.
The rare-earth mineral monazite is recovered primarily as a
byproduct of titanium operations. Treatment technology employed
at these operations is discussed under Titanium in this section.
Segregation of Waste Streams. Because mine/mill 9903 is
located in an arid region, water is a scarce commodity at this
site. It is primarily for this reason that water is recycled
from the tailing pond back to the flotation circuit. The
leach-circuit wastes are not combined with the water to be
recycled, as this waste contains very high dissolved-solid
VII-162
DRAFT
-------
DRAFT
TABLE VII-46. CHEMICAL COMPOSITION OF TREATED EFFLUENT AND
WASTE LOAD FROM MINE/MILL 9904 (PLATINUM)
PARAMETER
COD
TSS
Fe
Pb
Zn
Fluoride
CONCENTRATION (mg/fc)
IN WASTEWATER
7.6
30
0.17
0.01
0.03
0.95
TREATED WASTE LOAD
per unit of ore milled
kg/1000 metric tons
0.11
0.43
0.002
0.0001
0.0004
0.01
lb/1000 short tons
0.22
0.86
0.004
0.0002
0.0008
0.02
TABLE VI1-47. CHEMICAL COMPOSITIONS OF RAW WASTEWATER
AND TREATED RECYCLE WATER AT MILL 9903
(NO DISCHARGE)
PARAMETER
PH
TSS
TDS
TOC
Cr
Total Mn
V
Y
La
Ce
Pr
Nd
Sm
Eu
Gd
Th
Fluoride
CONCENTRATION img/SL)
RAW WASTEWATER
9.02*
360,000
14,476
3.100
0.35
0.5
<0.3
—
—
—
—
—
—
—
—
—
365
TREATED
RECYCLE WATER
7.58*
17.300
9.576
1,400
0.03
4.5
<0.3
0.014
1.32
2.75
0.27
0.51
41
< 0.001
0.006
< 0.001
55
Value in pH units
VII-163
DRAFT
-------
DRAFT
concentrations! which would undoubtedly cause interference
in the flotation circuit. At this mill, the waste streams
have been segregated, then, to facilitate recycle.
Tin Ores
Tin is obtained as a byproduct of molybdenum mining and milling
at one location in the United States. No separate discharges
result from tin mining or processing.
Titanium Ores
Titanium ores mined and milled in the United States occur in
two modes: as a hard rock deposit and as placer or heavy-
sand deposits of ilmenite, rutile, and leucoxene. The methods
of mining and beneficiation of both types of deposits are
described in detail in Section III. The treatment and control
technologies employed at exemplary operations in this ore •
category are described below.
Mine/Mill 9905. In the U.S., one operation is presently mining
a lode deposit for titanium minerals (primarily, ilmenite) .
At this operation, ore mined from an open-pit mine is crushed
and floated to concentrate the ilmenite. Prior to flotation,
magnetite associated with the ilmenite is magnetically separated
from the ore.
Process wastes, largely from the flotation circuit, are discharged
to a formerly used open-pit quarry, which serves as a tailing
pond. Clarified overflow from this pit is recycled back into
the mill circuit. Tailing-pond treatment-efficiency data are
presented in Table VII-48. No chemicals are added for treatment
purposes, although the process water has an alkaline pH.
Although this mill employs a recycle system, rain and runoff
which collect in the recycle system occasionally result in a
seasonal discharge. Diversion ditching is not presently used
at this mill. If diversion ditching or other systems were
installed to prevent excess water from collecting, a seasonal
discharge might not occur at mill 9905.
Water is currently discharged from open-pit mine 9905. Prior to
final discharge, this water is retained for settling for a short
time in a small pond. Improved treatment of this mine water
could be attained by increased retention time in a pond, and by
VII-164
DRAFT
-------
DRAFT
TABLE VI1-48. CHEMICAL COMPOSITIONS OF RAW WASTE WATER
AND TREATED RECYCLE WATER AT MILL 9905
PARAMETER
Conductivity
Turbidity (JTU)
TSS
IDS
TOC
Oil and Grease
As
Cd
Or
Cu
Total Fe
Pb
Total Mn
Hg
Ni
V
Ti
Zn
Nitrate
CONCENTRATION (mg/£)
RAW WASTE WATER
660*
2.2
26,300
518
9.0
2.0
<0.01
<0.002
0.58
0.43
630
<0.05
5.9
0.004
1.19
2.0
2.08
7.6
0.68
TREATED
RECYCLE WATER
490*
056
2
526
125
2.0
0.01
< 0.002
0.02
<0.03
<0.02
<0.05
0.3
< 0.0002
<0.01
<0.5
<0.2
< 0.002
0.50
Value in micromhos/cm
VII-165
DRAFT
-------
DRAFT
treatment with lime or other precipitating agent to ensure
optimum metal and fluoride removal.
Mine/Mills 9906 and 9907. These operations recover titanium
minerals (ilmenlte and rutlie) and the zirconium mineral
zircon from sand placers. Similar operations also recover the
rare-earth mineral monazite.
As these placer deposits are located Inland, the typical practice
is to construct a pond over the ore body and to mine the placer
by dredging. The heavy-mineral sands are upgraded by gravity
methods or use of a wet mill, and the heavy minerals in the bulk
concentrate are separated and concentrated by electrostatic
and magnetic methods in a dry mill.
Process wastes emanating from the wet mill are discharged to
the dredge pond. However, as discussed in Section V, the primary
waste constituents of the dredge-pond effluents are the colloidal
organic materials, of high coloring capacity, present in the
ore body. These materials are flocculated by reducing the pH
to 3.5 with sulfuric acid. The water then flows through a
large pond system, where the coagulated sludge settles. The
clarified overflow from this system is neutralized with lime
prior to final discharge to the receiving stream. Both acid
and lime are fed by automatically controlled equipment. Reagents
are added to the waste stream in flumes designed to create
turbulent mixing. The treatment efficiency of this system is
presented in Tables VII-49 and VII-50 for operations 9906 and
9907, respectively. Waste-load reduction data are presented
in Tables VII-51 and VII-52.
Potential Control Technology a£ Sand Placer Operations.
Water used in the wet mill at these placer mines is drawn from
the dredge pond; therefore, in this sense, process water is
recycled. However, some fresh water is required for use as
pump seals, as wash water in the finisher spirals, and in
"laminar flows" (gravity-separation devices), and this water is
drawn from a well.
A degree of waste-load reduction could be achieved by partial
recycle of the treated dredge-pond effluent back to the wet
mill for use in the finisher spirals or laminar flows. Treated
water would be suitable to replace the fresh water now used
VII-166
DRAFT
-------
DRAFT
TABLE VI1-49. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTEWATERS AT MILL 9906
PARAMETER
PH
Conductivity
Color
TDS
TSS
COD
TOC
Oil and Grease
Al
As
Cr
Cu
Total Fe
Total Mn
Hfl
Ti
Zn
CONCENTRATION (mg/fc)
RAW WASTE WATER
125?.
51,400
1.606
11,000
1,337.6
972.0
400
69
0.05
0.03
<0.03
4.9
0.036
-
<0.2
0.014
TREATED EFFLUENT
V*
280t
75"
96
11
14.4
63
1.0
2.8
0.01
<0.01
<0.03
0.25
<0.01
0.0002
-------
DRAFT
TABLE VI1-50. CHEMICAL COMPOSITIONS OF RAW AND TREATED
WASTEWATERS AT MILL 9907
PARAMETER
pH
Conductivity
Color
TDS
TSS
COD
TOC
Oil and Grease
At
As
Cr
Cu
Total Fe
Total Mn
HB
Ti
Zn
CONCENTRATION (mg/fc)
RAWWASTEWATER
40f
16.240**
370
209
361.6
321.2
40
15
0.03
<0.01
<0.03
0.93
<0.01
0.0024
0.40
< 0.002
TREATED EFFLUENT
6.4*
255f
••
13
172
4
12.8
3.8
1.0
1.0
0.01
<0.01
<0.03
0.12
0.04
0.003
<0.2
0.025
Value in pH units
Value in micromhos/cm
Value in cobalt units
VII-168
DRAFT
-------
DRAFT
TABLE VM-51. WASTEWATER COMPOSITION AND TREATED WASTE LOAD
WITH ACID FLOCCULATION AND SETTLING AT MILL 9906
PARAMETER
PH
TDS
TSS
COD
TOC
Oil and Grease
Al
As
Cr
Cu
Total Fa
Total Mn
Hg
Ti
Zn
CONCENTRATION (mg/i)
IN WASTEWATER
7.7*
96
11
14.4
6.8
1.0
2.8
0.01
<0.01
<0.03
0.25
<0.01
0.0002
<0.2
0.017
TREATED WASTE LOAD .
per unit concentrate produced
kg/1000 metric tons
—
4.130
473
620
290
43
120
0.43
<0.43
<1.3
11
<0.43
0.009
<8.6
0.73
lb/1000 short tons
—
8,260
946
1,240
680
86
240
0.86
<0.86
<2.6
22
<0.86
0.018
<17.2
1.46
Total amount of ore milled unavailable
tValue in pH units
VII-169
DRAFT
-------
DRAFT
TABLE VII-52. WASTEWATER COMPOSITION AND TREATED WASTE LOAD
WITH ACID FLOCCULATION AND SETTLING AT MILL 9907
PARAMETER
PH
TDS
TSS
COD
TOC
Oil and Grease
Al
As
Cr
Cu
Total Fe
Total Mn
Hg
Ti
Zn
CONCENTRATION (mg/£)
IN WASTEWATER
6.4*
172
4
12.8
3.8
1.0
1.0
0.01
<0.01
<0.03
0.12
0.04
0.0003
<02
0.025
TREATED WASTE LOAD
per unit concentrate produced*
kg/1000 metric tons
7.050
164
520
150
41
41
0.41
<0.41
<1.2
4.9
1.6
0.01
<032
1
lb/1000 short tons
14.100
328
1,040
300
82
82
0.82
<0.82
<2.4
9.8
3.2
0.02
<1.6
2
Total amount of ore milled unavailable
Value in pH units
VII-170
DRAFT
-------
DRAFT
in the wet mill. The primary reason why this practice is
not currently employed is that water can be drawn from wells at
less expense than required to recycle treated water.
Zirconium Ores
No primary operations for zirconium ores exist in the United
States. Zirconium is obtained as a byproduct of heavy-mineral
sand placer operations for titanium. No separate discharge or
waste loading can be assigned to this metal.
NON-WATER-QUALITY ENVIRONMENTAL ASPECTS
The effects of these treatment and control technologies on
air pollution, noise pollution, and radiation are usually small
and are not of any significance. The uranium-ore mining and
milling industry may produce wastes which are not compatible
with environmental health and which may require additional
handling safeguards, such as stabilization of tailing-disposal
areas, treatment-lagoon lining, etc.
Large amounts of solid waste in the form of solids and sludges,
created by settling and chemical precipitation, may result from
treatment used for settling, neutralization, and precipitation.
Dredging is often employed to clear tailing impoundments and
settling basins of unwanted solids and sludges. Often, these
solids and sludges are left in the impoundments and eventually
dry. Acquisition and dedication of extensive land areas for
treatment facilities such as large tailing-disposal areas
represent an adverse effect which is difficult to avoid.
Detailed preplanning and adoption of methods which reduce land
areas needed can alleviate this problem to some extent. At
many locations, land availability and topography present no
immediate problems and the space for dumping dredged solids
or for constructing treatment facilities presents no difficulty.
VII-171
DRAFT
-------
DRAFT
SECTION VIII
COST, ENERGY, AND NONWATER-QUALITY ASPECTS
INTRODUCTION
The costs of water-pollution control in the ore mining and
dressing industry are directly related to the implementation
of new and improved treatment methods. Capital and annual
costs incurred by representative operations in each of the
industries, and in subcategories within the industries, are
presented in this Section. These costs represent incremental
costs to attain specified effluent treatment levels. For
example, if the prevailing current practice encompasses use
of tailing ponds, the capital and operating costs associated
with such ponds are not included. The costs of any additional
treatment facility or activity necessary to meet the pres-
cribed standards, however, are included.
Separate capital and annual costs to achieve BPT, BAT, and
New Source Performance Standards are tabulated for typical
plants in each industry and in each industry subcategory.
Again, these are always expressed as incremental costs. These
costs are then combined in a summary table to show the total
costs incurred to attain the specified effluent levels. All
costs are expressed in 1972 dollars. The Marshall and Stevens
Equipment Cost Index for mining and milling is used where
cost adjustments are required.
A summary of the costing methodology employed is presented in
the section which follows. A detailed description of the cost
categories, factors, relationships, data sources, and assump-
tions utilized in computation of the industry costs is con-
tained in Supplement B. The selected approach entailed
the derivation and validation of costs for the various faci-
lities, activities, and materials which, in combination, form
the specified treatment processes.
Where applicable and practical, the costs are developed as a
function of variables which are generally known for specific
facility operations. Supplement B is organized to facilitate
the computation of treatment costs for other specified plant
operations.
VIII-1
DRAFT
-------
DRAFT
SUMMARY OF METHODS USED
Capital Coats
Capital costs include all costs incurred for the construction,
procurement, and installation of required treatment facilities
and equipment.
The major facility and equipment categories used to compute
capital costs are:
Impoundments
Settling Ponds/Lagoons
Tailing Ponds
Tailing-Pond Distribution System
Treatment Processes/Facilities/Equipment
Clarifiers/Thlckeners
Lime Neutralization and Precipitation
Hydrated-Lime System
Pebbled-Lime System
Coagulation/Flocculation (including Ferric Sulfate
Treatment)
Sulfide-Precipitation Treatment
Ion Exchange
Aeration
Barium Chloride Coprecipitation
Ammonia Stripping
Recarbonation/Sulfur Dioxide Addition
Transport Systems
Pipes
Pumps
Land
Other Costs
Contingency
Contractor Fee
The cost of impoundment is computed as a function of the
volume contained, total depth, and dike dimensions. Large
variations in costs are encountered for the construction of
an Impoundment of given size. A major factor is the local
topography. For example, very little dike construction may
be necessary where advantage is taken of an existing ground
depression. In other areas, dikes may have to be constructed
along the entire perimeter. In estimating impoundment costs
VIII-2
DRAFT
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DRAFT
for typical plants, it has been assumed that dikes must be
constructed around the entire perimeter. Detailed data are
presented in Supplement B, however, which permit estimation
of costs for specific lagoon and dike designs.
It is assumed that cyclones are used at tailing ponds to sep-
arage solids from the waste streams.
Thickener and clarifier costs are based on vendor quotations.
Costs are determined as a function of capacity.
Treatment costs vary with the characteristics and magnitude
of the waste streams. Two types of lime neutralization/pre-
cipitation facilities are considered. One uses hydrated lime,
introduced as a slurry; the other, pebbled lime, stored dry.
The first is practical for operations characterized by
flows of less than 18,925 cubic meters (5,000,000 gallons)
per day. The second is generally used to treat waste streams
of higher volume.
The major components of the hydrated-lime system are tanks,
a slurry mixer and feeder with associated instrumentation,
pumps, and a building to house the latter two components.
Lime storage consists of a 15- to 30-day supply. Treatment
facility costs are computed for application of 0.45 and 0.90
kg of lime per 3,785 liters (1 and 2 lb/1000 gal) of effluent
flow.
The pebbled-lime system consists of storage silo(s), lime
feeders and slakers, mixing tanks, and pumps. Storage silos
are designed to accommodate a 15-day supply of lime. Lime
feeders and slakers with feed rates of 455 to 1,818 kg (1,000
to 4,000 Ib) per hour are used, together with mixing tanks
of sufficient size for 15-minute retention. Costs are
developed for treatment systems designed to add 0.9 or 1.4 kg
of lime per 3,785 1 (3.785 cubic meters) (equivalent to 2
and 3 lb/1,000 gal) of waste water.
In some instances, slightly larger applications of lime than
previously noted are necessary where either hydrated- or
pebbled-lime facilities are used. No changes in facilities
are made in these cases. Rather, it is assumed that the
lime storage facilities are resupplied more frequently. The
increased application of lime thus is reflected only in
increased operating costs.
VIII-3
DRAFT
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DRAFT
Many variations of coagulation and flocculation are possible.
One basic system is considered in this study. It consists
of a mixing tank(s), two holding tanks, and two positive-
displacement pumps. The flocculant is mixed to provide a
0.5-percent solution. The mixture is then transferred to a
holding tank, where the solution is diluted to 0.1 percent.
One of the holding tanks is used to feed the solution into the
waste stream while a new batch of solution is made up in
the other. The pumps are used to transfer the solution from
the mixing tank to the holding tank and to meter the solu-
tion into the waste stream.
Ferric sulfate treatment is essentially similar to coagulation/
flocculation. A three percent solution is mixed directly in
two holding tanks and metered into the waste stream. Each tank
holds a one-day supply of solution. The need for the mixing
tank and one pump is eliminated.
Coagulation/flocculation and ferric sulfate systems are tailored
to individual plant requirements, as shown in Supplement B.
An important aspect to be noted here is that there are tradeoffs
between equipment sizes and the number of batches of solution
mixed daily.
The cost of installing a sodium sulfide treatment system
generally is very low. In many instances, this system consists
of a 208-liter (55-gallon) drum, from which the sulfide solution
trickles into the waste stream. The amount needed depends
on the characteristics of the waste stream; generally, it is
of the order of 1 to 2 mg/1 (1 to 2 ppm).
The cost of an ion-exchange unit is a function of the amount
of resin needed, which, in turn, depends on the daily flow,
the characteristics of the wastewater, and the specific
standard to be achieved. The amount of resin required is
determined for each plant where this treatment is employed.
The ion-exchange unit costs include purchase costs of the main
unit, and ancillary equipment, as well as installation costs.
Two applications of aeration are considered in the study: one
for mixing, the other for oxidation. The former is designed
to raise the DO level in the wastewater. Its cost is deter-
mined on the basis of the volume of water to be agitated.
The latter application consists of the chemical addition of
oxygen, where the amount of oxygen required is a function of
chemical change to be achieved. The cost in this case is
computed from the amount of oxygen which must be added to
the water.
VIII-4
DRAFT
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DRAFT
Barium chloride coprecipitation treatment costs are based on
industry sources. The main components of a ammonia stripper
are a plastic mixing tank containing caustic soda, a metering
pump, and a packed column. This treatment is used in only one
instance. Costs are from Reference
Both recarbonation and sulfur dioxide addition utilize a holding
tank sized for five minutes of retention. Carbon dioxide or
sulfur dioxide is bubbled through the wastewater while it is
contained in the holding tank.
Piping and pump requirements depend on the average flow rates,
the characteristics of the waste stream, and the distance
over which the waste stream must be transported. Pipe and
pump sizes and costs for waste streams which contain a signi-
ficant amount of solids are based on a flow rate of 1m (3.3 ft)
per second and on the use of slurry pumps. Wastewater which
carries relatively little solid material is assumed to be
pumped at a rate of 2 m (6.6 ft) per second utilizing water
pumps. The cost of a standby pump is included in all cases.
All facilities are assumed to be located on rural land. The
cost used is $1,755 per hectare ($730/acre).
Contingency and contractor fees are included as 13 percent
of the capital costs.
Annual Costs
The cost categories included are:
Annual capital recovery
Facility repair and maintenance
Equipment repair and maintenance
Operating personnel
Material
Energy (Power)
Taxes
Insurance
Annual capital recovery, as defined for this study, includes
the cost of both capital and the depreciation. The cost of
capital is computed at 8 percent. The assumed useful lives
of facilities and equipment are 20 and 10 years, respectively.
Annual capital recovery costs are computed as follows.
VIII-5
DRAFT
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DRAFT
CA - B (r) (1 + r)exp
where
B
r
n
((1 + r)exp n) - 1
= Initial cost
- True annual interest rate
= Useful life in years
Annual land cost is also included in the capital recovery
cost. This cost is computed as an opportunity cost at an
annual rate of 10 percent.
Facility repairs and maintenance are included as 3 percent
of initial capital cost, excluding contingency and fee. The
rate applied to equipment is 5 percent of initial installed
cost per year. This is an average cost applicable to mining
and milling equipment.
One exception to the above rates is the maintenance and repair
of tailing ponds. Extensive effort is required for periodically
raising the distribution pipes, moving the cyclones, and reshaping
the upper portions of the dike(s). The annual cost is estimated
at 30 percent of the initial cost of the distribution system
(Reference 68).
Operating personnel are assigned for specific tasks which
must be performed at the treatment facilities. A cost of $9.00
per hour, which includes fringe benefits, overhead, and super-
vision, is applied.
Material costs are a function of the type of treatment process
employed, the volume of the wastewater which must be treated,
its characteristics, and the effluent levels which must be
attained. Representative delivered material costs are:
Pebbled Lime
Hydrated Lime
Sodium Sulfide
Flocculant
Alum
Ion-Exchange
(IX) Resins
Ferric Sulfate
Barium Chloride
$ 30.80/metric ton
38.50/metric ton
0.22/kg
2.20/kg
0.07/kg
2,500/cubic meter
49.50/metric ton
805.00/metric ton
28.00/short ton
35.00/short ton
0.10/lb
1.00/lb
0.03/lb
70.80/cubic foot
45.00/short ton
730.00/short ton
VIII-6
DRAFT
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DRAFT
Energy costs are based on the cost per horsepower-year, com-
puted as follows:
Cy = HP x 0.7457 x Hr x Ckw
E x P
where
Cy = Cost per year
HP = Total horsepower of motors
E = Efficiency factor
P = Power factor
Hr - Annual operating hours
Ckw = Cost per kilowatt hour
Efficiency and power factors are each assumed Co be 0.9;
the cost per kilowatt hour, $0.012.
The computed cost is increased by 10 percent to account for
miscellaneous energy usage.
Annual taxes are computed as 2.5 percent of land costs.
Insurance is estimated at 1 percent of capital cost.
The discussions which follow are presented by ore mining/
milling category and subcategory. Subcategories in which
no operations currently have discharges are not discussed
inn this section.
VIII-7
DRAFT
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DRAFT
WASTEWATER-TREATMENT COSTS FOR IRON-ORE CATEGORY
Iron-Ore Mines
There are 39 major Iron-ore-producing mines currently in opera-
tion. Ore production ranges from 65,300 to 40,634,000 metric
tons (73,000 to 44,800 short tons) annually, with mine
wastewater ranging from 0 to 80,000 cubic meters (0 to 21,000,000
gallons) per day.
An existing mine with an annual ore production of 8,460,000
metric tons (9,400,000 short tons) and a wastewater flow of
47,520 cubic meters (12,500,000 gallons) per day was chosen
to represent this subcategory.
Two levels of technology are considered. The total cost of each
level is shown in Table VIII-1.
Wastewater Treatment/Control
Level A_:_ Coagulation/Flocculation, Settling, and Discharge
The mine wastewater is treated with 25 mg/1 of alum and 1 mg/1 of
flocculant for suspended-solid removal. The treated effluent is
then retained for two days in a settling pond before discharge.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Flocculation system -
1 mixing tank of 1900-liter (500-gallon) capacity
2 holding mix tanks of 9,500-liter (2,500-gallon)
capacity
Piping - Flow @ 1 m (3.3 ft)/sec through 60-cm (2-ft) x
250-meter (820-foot) pipe
Pumps - 2 positive-displacement
Pond - 4-meter (13-foot) dike height
6-meter (20-foot) top width
143,000-cubic-meter (37,777,000-gal) capacity
Land - 4.2 hectares (10 acres)
VIII-8
DRAFT
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DRAFT
TABLE VIII-1. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MINE 1105
SUBCATEOORV: Iron-Ore Mines
PLANT SIZE; 8,460,000
PLANT APE; 7 YEARS
METRIC TONS (9.400.000 SHORT TONS) PER YEAR OF ore mined
PLANT LOCATION; Mesabi Range
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
192.5
21.1
88.6
1.3
111.0
0.013
B
384.6
49.7
241.4
15.9
307.0
0.036
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
Dissolved Fe
Total Mn
CONCENTRATION (mg/£) (ppm)
RAW?
(UN-
TREATED)
30
10
2.1
6.9
AFTER TREATMENT TO LEVEL
A
20
10
1.0
3.0
B
20
10
0.5
2.0
C
D
E
"ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0:807
'HISTORICAL DATA
LEVEL A: COAGULATION/FLOCCULATION, SETTLING, AND DISCHARGE
LEVEL B: LEVEL A PLUS LIME PRECIPITATION
VIII-9
DRAFT
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DRAFT
Operating-Cost Assumptions for Level A:
Coagulant - 415.8 metric tons (457.4 short tons)/year
Flocculant - 16.67 metric tons (18.34 short tons)/year
Operating personnel - 5 mixes/day @ 1 hr/mix
Power - 9.7 kW (13 hp)
Capital Investment;
Facilities
Lagoon $ 122,000
Contingency and contractor's fee 15,860
Total facility cost $ 137,860
Land 7,350
Equipment
Flocculation/Coagulation unit 14,900
Piping 27.000
Equipment subtotal 41,900
Contingency and contractor's fee 5,445
Total equipment cost 47,345
Total Capital Investment 192,555
Annual Cost;
Amortization
Facility $ 14,040
Equipment 7,055
Total Amortization $ 21,095
VIII-10
DRAFT
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DRAFT
Operation and Maintenance (O&M)
Land $ 735
Operating personnel 15,750
Facility repair and maintenance 3,660
Equipment repair and maintenance 2,095
Materials 64,260
Taxes 185
Insurance 1,925
Total O&M costs $ 88,610
Electricity 1.325
Total Annual Cost $ 111.030
Level B; Level A plus Lime Precipitation
In addition to level-A technology, the wastewater is treated with
0.9 kg of pebbled lime per 3.785 cubic meters (2 lb/1000 gallons)
of wastewater before entering the settling pond. The incre-
mental cost for lime precipitation is shown below.
The capital and operating costs and assumptions for attaining
level B are shown below.
Capital-Cost Components and Assumptions for Level Bj^
Lime precipitation system
Operating-Cost Assumptions for Level B;
Lime - 4,000 metric tons (4,410 short tons)/year
Operating personnel - 2 hr/shift, 6 hr/day
Power - 108 kW (145 hp)
Capital Investment;
Equipment
Lime precipitation unit $ 170,000
Contingency and contractor's fee 22,100
Total equipment cost $ 192,100
Total Capital Investment $ 192.100
VIII-11
DRAFT
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DRAFT
Annual Cost:
Amortization
Operation and Maintenance (O&M)
Operating personnel
Equipment repair and maintenance
Materials
Insurance
Electricity
Total O&M costs
Total Annual Cost
$ 18,900
8,500
123,480
1.920
$ 28,630
152,800
14.570
$ 196.000
Iron-Ore Mills Employing Chemical/Physical Separation
There are 34 iron-ore mills in this subcategory. The amount
of ore milled ranges from 364,000 to 6,600,000 metric tons
(402,000 to 7,236,000 short tons) annually. The daily mill
wastewater ranges from 0 to 22,320 cubic meters (0 to 5,900,000
gallons) .
The representative mill operation employing a chemical and/or
physical process mills 5,000,000 metric tons (5,550,000 short
tons) of ore annually. The wastewater flow is 13,435 cubic
meters (3,550,000 gallons) per day.
Two levels of technology are considered for this subcategory.
The total cost of each level is shown in Table VIII-2.
Wastewater Treatment/Control
Level AJ:_ Floeculation, Settling, and Discharge
The wastewater is treated with 5 mg/1 of flocculant and flows,
by gravity, to a settling pond. The retention time is assumed
to be two days before discharge.
The capital and operating costs and assumptions for attaining
this level are shown below.
VIII-12
DRAFT
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DRAFT
TABLE VIII-2. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 1107
8UBCATEOORY; Iron-Ore Mills Employing Chemical/Physical Separation
PLANT SIZE; 5.000.000 METRIC TONS ( 5 .500 .000 SHORT TONS) PER YEAR OF ore milled
PLANT AGE; 17 YEARS PLANT LOCATION; Michigan
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
65.0
7.5
80.1
1.3
88.9
0.018
B
181.0
24.8
139.3
13.3
177.4
0.035
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
Dissolved Fe
Total Mn
WASTE LOAD (kg/1000 metric tons of product [ore milled] *)
RAW
(UN-
TREATED)
56.2**
13.2
2.06**
5.14"
AFTER TREATMENT TO LEVEL
A
26.2
13.2
1.32
2.7
B
26.2
13.2
0.66
2.7
C
0
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.907
fTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
"MILL 1108 DATA
"MILL no? DATA
LEVEL A: FLOCCULATION. SETTLING, AND DISCHARGE
LEVEL B: LEVEL A PLUS LIME PRECIPITATION
VIII-13
DRAFT
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DRAFT
Capital-Cost Assumptions for Level A;
Pond - 3-meter (10-foot) dike height
6-meter (20-foot) top width
40,300-cubic-meter (10,646,000-gal) capacity
Flocculation system -
1 mixing tank @ 1,900-liter (500-gallon) capacity
2 holding tanks @ 9,500-liter (2,500-gallon) capacity
2 positive-displacement pumps
Piping - Flow @ 1 meter (3.3 feet)/sec through 32-cm
(1-ft) x 100-meter (328-foot) pipe
Land - 1.6 hectares (4 acres)
Operating-Cost Assumptions for Level Aj_
Flocculant - 23.45 metric tons (25.8 short tons)/year
Operating personnel - 8 mixes/day @ 1 hour/mix
Power - 9.7 kW (13 hp)
Capital Investment;
Facilities
Lagoon $ 34,100
Contingency and contractor's fee 4,435
Total facility cost $ 38,535
Land 2,800
Equipment
Flocculation unit 14,900
Piping 6.100
Equipment subtotal 21,000
Contingency and contractor's fee 2,730
Total equipment cost 23,730
Total Capital Investment $ 65,065
VIII-14
DRAFT
-------
D* I"
^.^%t I
Annual Cost:
Amortization
Facility S 3,925
Equipment 3,535
Total amortization $ 7,460
Operation and Maintenance (O&M)
Land 280
Operating personnel 25,200
Facility repair and maintenance 1,025
Equipment repair and maintenance 1,050
Materials 51,805
Taxes 70
Insurance 65G
Total O&M costs 80,080
Electricity 1.325
Total Annual Cost $ 88,865
Level Bj_ Level A_ plus Lime Precipitation
In addition to level-A tecnnology, the wastewater is treated with
0.9 kg of hydrated lime per 3.785 cubic meters (2 lb/1000 gal)
cf wastewater before entering the settling pond.
The capital and operating costs and assumptions for attaining
this level and this size of operation are shown below.
Capital-Cost Components and Assumptions for Level B;
Lime precipitation system
Operating-Cost Assumptions for Level B;
Lime - 1,127 metric tons (1,240 short tons)/year
Operating personnel - 1 hr/shift, 3 hr/day
Power - 81 kW (108 hp)
VIII-15
DRAFT
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DRAFT
Capital Investment:
Equipment
Lime precipitation unit $ 102,650
Contingency and contractor's fee 13,345
Total equipment cost $ 115,995
Total Capital Investment $ 115.995
Annual Cost;
Amortization
Equipment $ 17,285
Total amortization $ 17,285
Operation and Maintenance (O&M)
Operating personnel 9,450
Equipment repair and maintenance 5,135
Materials 43,490
Insurance 1,160
Total O&M costs 59,235
Electricity 12.000
Total Annual Cost $ 88,520
VIII-16
DRAFT
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DRAFT
WASTEWATER-TREATMENT COSTS FOR COPPER-ORE CATEGORY
Copper Mines
There are 55 major copper-producing mines currently in opera-
tion. Ore production ranges from 130,320 to 34,500,000 metric
tons (143,600 to 38,000,000 short tons) annually. Mine waste-
water ranges from 0 to 30,522 cubic meters (0 to 8,064,000
gallons) per day.
A representative copper mine produces 16,550,000 metric tons
(18,250,000 short tons) a year and has an average daily waste-
water flow of 2,725 cubic meters (720,000 gallons).
One level of technology is considered. The total cost for this
level is shown in Table VIII-3.
Wastewater Treatment/Control
Level A: Lime Precipitation, Settling, Recarbonation, and
Discharge
The mine drainage is treated with 0.9 kg of hydrated lime per
3.785 cubic meters (2 lb/1000 gal) to precipitate dissolved
metals. The treated effluent is then retained in a settling
pond for two days. Recarbonation is required for pH adjust-
ment before discharge.
The capital and operating cost components and assumptions for
attaining this level are shown below.
Capital-Cost Components and Assumptions for Level A_:
Pond - 3-meter (10-foot) dike height
3-meter (10-foot) top width
8,500-cubic meter (2,245,000-gal) capacity
Lime precipitation system
Recarbonation system -
1 holding tank, 5-minute retention, 9,500-liter
(2,510-gallon) capacity
1 ejector
VIII-17
DRAFT
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DRAFT
TABLE VIII 3. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MINE 2120
SUBCATEOORY: Copper Mines
PLANT SIZE: 16,550,000 METRIC TONS (18 ,250 .OOCSHORT TONS) PER YEAR OF ore mined
PLANT AGE: 19 YEARS PLANT LOCATION: Montana
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTStSI/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
108.1
15.3
24.0
5.0
44.3
0.003
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
As
Cd
Pb
Hg
Fe
Zn
Cu
CONCENTRATION (mg/JU (ppm)
RAW
(UN-
TREATED)
40
1
0.07
0.175
0.1
0.0005
6.0
31.3
5.30
AFTER TREATMENT TO LEVEL
A
20
1
0.05
0.005
0.02
0.0002
0.1
0.05
0.05
B
C
D
•
E
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0.907
LEVEL A: LIME PRECIPITATION, SETTLING. RECARBONATION, AND DISCHARGE
VIII-18
DRAFT
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DRAFT
Piping - Flow @ 2 meters (6.6 feet)/sec through 14-cm
(5.5-in.) x 1000-meter (3,280-foot) pipe
Land - 0.54 hectare (1.33 acres)
Operating-Cost Assumptions for Level A;
Lime - 228.6 metric tons (251.5 short tons)/year
Operating personnel - 1 hr/shift, 3 hr/day
Power - 37 kW (50 hp)
C02_ - can be reclaimed from milling operations; thus,
no additional cost
Capital Investment;
Facilities
Lagoon $ 12,000
Contingency and contractor's fee 1,560
Total facility cost $ 13,560
Land 975
Equipment
Lime precipitation unit 45,000
Recarbonation 3,800
Piping 34.000
Equipment subtotal 82,800
Contingency and contractor's fee 10,765
Total equipment cost 93,565
Total Capital Investment $ 108,100
Annual Cost;
Amortization
Facility $ 1,380
Equipment 13,945
Total amortization $ 15,325
VIII-19
DRAFT
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DRAFT
Operation and Maintenance (O&M)
Land
Operating personnel
Facility repair and maintenance
Equipment repair and maintenance
Materials
100
9,450
360
4,140
8,820
Taxes
Insurance
Electricity
Total O&M costs
Total Annual Cost
$ 23,975
5.000
$ 44,300
Copper Mills Using Froth Flotation (Precipitation Minus
Evaporation - -76 cm tp_ Positive) (-30 in. t£ Positive)
There are five mills in this subcategory. Ore production
ranges from 1,211,000 to 17,714,000 metric tons (1,336,000
to 19,530,000 short tons) each year. The daily wastewater
flow ranges from 21,760 to 95,000 cubic meters (5,750,000 to
25,000,000 gallons).
An existing operation that annually mills 8,000,000 metric tons
(8,840,000 short tons) with a daily wastewater flow of 95,000
cubic meters (25,000,000 gallons) was chosen to represent
this subcategory.
Two levels of technology are considered for this subcategory.
The total cost of each level is shown in Table VIII-4.
Wastewater Treatment/Control
Level A; Lime Precipitation. Polyelectrolyte Addition,
Settling, and Discharge
Approximately 70 percent of the mill effluent is treated with
1.36 kg of pebbled lime per 3.785 cubic meters (3 lb/1000 gal)
of wastewater to precipitate heavy metals from acid solution.
This is later mixed with the remaining effluent. In addition,
VIII-20
DRAFT
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DRAFT
TABLE VI11-4. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 2121
8UBCATEOQRY: Copper Mills Using Froth Flotation
PLANT SIZE: 8.000.000
PLANT AGE: 20 YEARS
.METRIC TONS ( 8 ,840 ,000 SHORT TONS) PER YEAR OF Ore milled
PLANT LOCATION; Michigan
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
523.7
64.8
342.2
21.5
428.5
0.054
B
1,921.0
286.3
104.2
90.0
480.5
0.06
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
Cyanide
Pb
Zn
Cd
Cu
Hg
Fe
WASTE LOAD (kg/1000 metric tons of product (ore milled]1)
RAW
(UN-
TREATED)
-
10
0.035
0.35
0.35
0.07
2.6
0.0027
18
AFTER TREATMENT TO LEVEL
A
35
3.5
0.035
0.2
0.2
0.02
0.2
0.0003
0.35
B
0
0
0
0
0
0
0
0
0
c
D
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0507
fTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED), MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: LIME PRECIPITATION, POLYELECTROLYTE ADDITION. SETTLING. AND DISCHARGE
LEVEL B: TOTAL RECYCLE (ZERO DISCHARGE)
VIII-21
DRAFT
-------
DRAFT
polyelectrolytes are added during upset conditions (spring and
summer) to increase flocculation. The effluent is retained for
two days In a settling pond before discharge. The capital and
operating cost components and assumptions for attaining this
level are shown below.
Capital-Cost Components and Assumptions for Level A£_
Pond - 4-meter (13-foot) dike height
6-meter (20-foot) top width
300,000-cubic-meter (79,252,000-gal) capacity
Lime precipitation system
Polyelectrolyte feed system - data supplied from
Industry surveys.
Piping - Flow @ 2 meters (6.6 feet)/sec through 84-cm
(33-in.) x 100-meter (328-foot) pipe
Land - 11 hectares (27 acres)
Operating-Cost Assumptions for Level Aj_
Lime - 8,100 metric tons (8,910 short tons)/year
Polyelectrolyte - 45.35 metric tons (50 short tons)/year
@ $900/metric ton
Operating personnel - 8 hr/day
Power - 160 kW (215 hp)
Capital Investment;
Facilities
Lagoon $ 194,000
Contingency and contractor's fee 25,220
Total facility cost $ 219,220
Land 19,250
VIII-22
DRAFT
-------
DRAFT
Equipment
Lime precipitation unit $ 230,000
Polyelectrolyte feed system 9,000
Piping 13.400
Equipment subtotal 252,400
Contingency and contractor's fee 32.810
Total equipment cost $ 285.210
Total Capital Investment $ 523.680
Annual Cost:
Amortization
Facility $ 22,330
Equipment 42,505
Total amortization $ 64,835
Operation and Maintenance (O&M)
Land 1,925
Operating personnel 25,200
Facility repair and maintenance 5,820
Equipment repair and maintenance 12,620
Materials 290,900
Taxes 480
Insurance 5,235
Total O&M costs 342,180
Electricity 21.500
Total Annual Cost $ 428,515
Level Bj_ Total Recycle (Zero Discharge)
Total recycle includes additional pumps and piping for recircu-
lating the impounded water from the tailing pond. The capital
and operating costs and assumptions for attaining this level
are shown below.
Capital-Cost Components and Assumptions for Level Bj_
Piping - Flow @ 2 meters (6.6 feet)/sec through 84-cm
(33-in.) x 10,000-meter (32,800-foot) pipe
Pumps - 9 75-kW (100-hp) plus 9 standbys
VIII-23
DRAFT
-------
DRAFT
Operating-Cost Assumptions for Level Bj_
Power - 675 kW (900 hp)
Capital Investment:
Equipment
Piping $1,340,000
Pumps 360,000
Equipment subtotal 1,700,000
Contingency and contractor's fee 221,000
Total equipment cost $1,921,000
Total Capital Investment $1.921.000
Annual Cost;
Amortization
Equipment $ 286.290
Total amortization $ 286,290
Operation and Maintenance (O&M)
Equipment repair and maintenance 85,000
Insurance 19,210
Total O&M costs 104,210
Electricity 90.000
Total Annual Cost $480,500
VIII-24
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR LEAD- AND ZINC-ORE CATEGORY
Lead/Zinc Mines With No Solubility Potential
There are 12 mines in this subcategory. Ore production ranges
from 143,300 to 2,280,000 metric tons (158,000 to 2,514,200
fchort tons) annually. Minfe wastewater flow ranges from 6,810
to 49,200 cubic meters (1,800,000 to 13,000,000 gallons) per day.
A hypothetical mine was selected as the representative for
this subcategory. It is assumed to have a wastewater flow of
18,925 cubic meters (5,000,000 gallons)' a day and an annual ore
production of 630,000 metric tons (700,000 short tons).
One level of technology is considered. The total cost of achieving
this level is shown in Table VIII-5.
Wastewater Treatment/Control
Level A£ Sedimentation Lagoon, Secondary Settling, and
Discharge
Since there is no solubilization potential for heavy metals, no
precipitation is necessary. However, suspended-solid concentra-
tions present a problem. The recommended technology includes
use of two settling ponds: one large pond with a 10-day
retention and a smaller polishing pond with a 2-day retention.
Capital and operating cost components and assumptions for attain-
ing this level are shown below.
Capital-Cost Components and Assumptions for Level Aj_
Pond A - 4-meter (13-foot) dike height
6-meter (20-foot) top width
250,000-cubic-meter (66,043,000-gallon) capacity
Pond B - 3-meter (10-foot) dike height
3-meter (10-foot) top width
50,000-cubic-meter (13,209,000-gal) capacity
Piping - from mine to pond A, 1000 meters (3,280 feet);
from pond A to pond B, 500 meters (1,640 feet).
Flow @ 2 meters (6.6 feet)/sec through
37.5-cm (14.8-in.) pipe.
VIII-25
DRAFT
-------
DRAFT
TABLE VIII-5. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL) -
LEAD/ZINC, NO SOLUBILITY
8UBCATEOQRY: Lead/Zinc Mines With No Solubility Potential
PLANT SIZE: 630,000 METRIC TONS (700,000 SHORT TONS) PER YEAR OF Ore mined
PLANT AOE:N/A YEARS PLANT LOCATION: N/A
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
413.6
46.7
19.5
8.2
74.4
0.12
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
COD
Oil and Grease
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
CONCENTRATION (mg/JU (ppm)
RAW
(UN-
TREATED)
138
630
29
0.015
0.2
0.05
1.0
0.0001
0.5
4.9
0.7
AFTER TREATMENT TO LEVEL
A
20
50
10
0.01
0.1
0.05
1.0
0.0001
0.5
0.1
0.1
B
C
D
E
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0.907
LEVEL A: SEDIMENTATION LAGOON. SECONDARY SETTLING, AND DISCHARGE
VIII-26
DRAFT
-------
DRAFT
Pumps - from mine to pond A - 1 plus standby,
13,140 1(3,469 gal)/minute each
Operating-Cost Assumptions for Level Aj_
Power - 60 kW (80 hp)
Capital Investment;
Lagoon(s) $ 225,800
Contingency and contractor's fee 29.355
Total facility cost $ 255,155
Land 19,425
Equipment
Piping 105,000
Pumps 18.000
Equipment subtotal 123,000
Contingency and contractor's fee 15,990
Total equipment cost 138,990
Total Capital Investment $ 413.570
Annual Cost;
Amortization
Facility $ 25,990
Equipment 20.715
Total amortization $ 46,705
Operation and Maintenance (O&M)
Land 1,945
Facility repair and maintenance 6,775
Equipment repair and maintenance 6,150
Taxes 485
Insurance 4,135
Total O&M costs 19,490
Electricity 8,165
Total Annual Cost $ 74,360
VIII-27
DRAFT
-------
DRAFT
Lead/Zinc Mines With Solubility Potential
There are 16 known mines in this subcategory. Annual ore
production ranges from 143,300 to 669,240 metric tons (158,000
to 737,860 short tons). Mine wastewater flow ranges from 950
to 131,050 cubic meters (251,000 to 34,623,500 gallons) per day.
A hypothetical mine was selected as representative for this
subcategory. It is assumed to have a wastewater flow of 18,925
cubic meters (5,000,000 gal) per day and an annual ore production
of 630,000 metric tons (700,000 short tons).
Two levels of technology are considered. The total cost of
achieving these levels is shown in Table VIII-6.
Wastewater Treatment/Control
Level A; Lime Precipitation, Settling, and Discharge
Acid mine wastewater has the potential for solubilization of
undesired metals. The technology utilized for this occurrence
is lime precipitation and settling. Since the mine drainage is
acid, a concentration of 1.36 kg of pebbled lime per 3.785
cubic meters (3 lb/1000 gal) of wastewater is required to raise
pH sufficiently high for precipitating metals. The treated water
is then retained for a minimum of 10 days before discharge.
Pumps are not listed as a separate item, since they are integral
parts of the lime precipitation unit. Capital and operating
cost components and assumptions for attaining this level are
shown below.
Capital-Cost Components and Assumptions for Level Ar_
Pond - 4-meter (13-foot) dike height
6-meter (20-foot) top width
250,000-cubic-meter (66,043,000-gal) capacity
Land - 9 hectares (22 acres)
Lime precipitation system
Piping - Flow @ 2 meters (6.6 feet)/sec through 37.5-cm
(14.8-in.) x 1000-meter (3,280-foot) pipe
VIII-28
DRAFT
-------
DRAFT
TABLE VI11-6. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL) -
LEAD/ZINC, SOLUBILITY
SUBCATEOORV: Lead/Zinc Mines With Solubility Potential
PLANT SIZE: 630,000 METRIC TONS (700.000 SHORT TONS) PER YEAR OFOre mined
PLANT AGE: N/A YEARS PLANT LOCATION: N/A
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS IS1000) TO ATTAIN LEVEL
A
407.3
49.1
115.5
10.9
175.5
0.28
B
671.5
88.5
129.8
11.9
230.2
0.37
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
COD
Oil and Grease
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
CONCENTRATION (mg/£) (ppm)
RAW
(UN-
TREATED)
58
95
15.0
0.055
0.042
0.06
2.5
0.0007
57.2
0.3
38.0
AFTER TREATMENT TO LEVEL
A
20
50
10
0.05
0.2
0.05
1.0
0.0001
0.5
0.2
0.5
B
20
50
10
0.01
0.01
0.05
1.0
0.0001
0.5
0.1
0.15
C
D
E
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT, MULTIPLY COSTS SHOWN BY 0.907
LEVEL A: LIME PRECIPITATION. SETTLING. AND DISCHARGE
LEVEL B: HIGH-DENSITY SLUDGE PROCESS
VIII-29
DRAFT
-------
DRAFT
Operating-Cost Assumptions for Level A;
Lime - 2,380 metric tons (2,625 short tons)/year
Operating personnel - 2 hr/shift, 6 hr/day
Power - 80 kW (107 hp)
Capital Investment
Facilities
Lagoon $ 174.000
Contingency and contractor's fee 22,620
Total facility cost $ 196,620
Land 15,750
Equipment
Lime precipitation unit 102,500
Piping 70.000
Equipment subtotal 172,500
Contingency and contractor's fee 22,425
Total equipment cost 194,925
Total Capital Investment $ 407.295
Annual Cost;
Amortization
Facility $ 20,025
Equipment 29.050
Total amortization $ 49,075
VIII-30
DRAFT
-------
DRAFT
Operation and Maintenance (O&M)
Land
Operating personnel
Facility repair and maintenance
Equipment repair and maintenance
Materials
Taxes
Insurance
Total O&M costs
1,575
18,900
5,220
8,625
73,500
3,625
4.070
Total Annual Cost
High-Density Sludge Process
In addition to lime and settling as described for level A, a
high-density sludge process has been suggested for enhanced
removal of dissolved metals.
This process has been costed as a separate item. The incre-
mental cost for implementing this system is shown below. The
total cost for this system must be added to level-A costs,
since lagoons and lime precipitation are necessary for the
operation of this technology. Capital and operating cost
components and assumptions for attaining this level are shown
below.
Capital-Cost Components and Assumptions for Level IJj_
Clarifier - 8-hr retention, 6,350-cubic-meter (1,680,000-gal)
capacity.
Underflow from clarifier is 10% of inflow, and
50% of underflow is discharged to settling pond
with overflow; thus, 5% of underflow is recir-
culated through lime precipitation unit.
Slurry Pump - 660 liters (174 gal)/minute
Pipe - Flow @ 1 meter (3.3 ft)/sec through 12.5-cm
(4.9-in.) x 50-meter (164-foot) pipe from clarifier
to precipitation unit.
Operating-Cost Assumptions for Level
Power - 7.5 kW (10 hp)
$ 115,515
10.900
$ 175.490
VIII-31
DRAFT
-------
DRAFT
Capital Investment;
Equipment
Clarifier
Piping
Pumps
Equipment subtotal
Contingency and contractor's fee
Total equipment cost
Annual Cost;
Amortization
Equipment
Total amortization
Operation and Maintenance (O&M)
Equipment repair and maintenance
Insurance
$ 226,800
1,500
5.500
233,800
30.395
$ 264.195
39.375
11,690
2.640
39,375
Electricity
Total O&M costs
Total Annual Cost
14,330
1.000
$ 54,705
Lead/Zinc Mills
There are 21 known major lead/zinc mills in operation. The
amount of ore milled by these operations ranges from 195,840
to 2,520,000 metric tons (215,920 to 2,778,390 short tons)
annually. The daily mill wastewater flow ranges from 0 to
15,120 cubic meters (0 to 4,000,000 gallons).
A hypothetical mill was selected as representative for this
subcategory. It is assumed to have an annual milling capacity
of 630,000 metric tons (700,000 shor tons), with a daily
wastewater flow rate of 5,678 cubic meters (1,500,000 gallons).
VIII-32
DRAFT
-------
DRAFT
Two alternative levels of technology are considered for this
subcategory. The total cost of each level is shown in Table
VIII-7.
Wastewater Treatment/Control
Level A^ Tailing Pond. Secondary Settling, and Discharge
The best practiced technology consists of use of a tailing
pond, followed by a secondary settling pond. A minimum 10-
day retention time in the tailing pond and a 2-day retention
time in the secondary settling pond are recommended. The
tailing distribution system consists of piping, around the
perimeter of the tailing pond, and cyclones, located at 100-
meter (328-foot) Intervals along one length of the tailing
dam.
Capital and operating cost components and assumptions for
attaining level A are shown below.
Capital-Cost Components and Assumptions for Level A;
Tailing pond - 3-meter (10-foot) dike height
3-meter (10-foot) top width
4,245-meter (13,925-ft) perimeter
Settling Pond - 3-meter (10-foot) dike height
3-meter (10-foot) top width
15,000-cubic-meter (3,963,000-gal) capacity
Land - 101 hectares (250 acres)
Distribution system - 4,245 meters (13,924 feet) of
(7.9-in.) pipe
12 cyclones @ $1,800 each
Piping - Flow at 1 meter/sec through 30 cm pipe:
from mill to tailing pond, 1000 meters (3,280 ft);
from tailing pond to lagoon, 500 meters (1,640 ft)
Slurry pumps - 1 plus standby, 3,900 1 (1,042-gal)/minute
VIII-33!
DRAFT
-------
DRAFT
TABLE VIII-7. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MILL (HYPOTHETICAL) -
LEAD/ZINC
SUBCATEOORY: Lead/Zinc Mills
PLANT SIZE; 650,000
PLANT AGE:N/AjYEARS
.METRIC TONS<700,OOP SHORT TONS) PER YEAR OF Ore milled
PLANT LOCATION: N/A.
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
1,117.0
116.6
124.7
2.5
243.8
0.38
B
1,199.0
128.8
129.1
6.5
264.4
0.42
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
COD
Oil and Grease
Cyanide
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
WASTE LOAD (kg/1000 metric tons of product [ore milled) f)
RAW
(UN-
TREATED)
285
1,770
21.1
0.11
0.02
1.77
0.19
1.16
0.0026
0.5
6.9
17.2
AFTER TREATMENT TO LEVEL
A
15
130
13
0.04
0.01
0.05
0.03
0.25
0.0004
0.5
0.07
0.1
B
0
0
0
0
0
0
0
0
0
0
0
0
C
D
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOVIff » . 0.9C7
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED) MULTIPLY VALUES SHOWN BY 2.".
LEVEL A: TAILING POND. SECONDARY SETTLING. AND DISCHARGE
LEVEL B: TOTAL RECYCLE (ZERO DISCHARGE)
VIII-34
DRAFT
-------
DRAFT
Operating-Cost Assumptions for Level At
Tailing-pond distribution system maintenance @ 30% of
distribution cost
Power - 18.6 kW (25 hp)
Capital Investment;
Facilities
Tailing pond $ 420,255
Lagoon 19.940
Facility subtotal 440,195
Contingency and contractor's fee 57,225
Total facility cost $ 497,420
Land 176,750
Equipment
Distribution system 284,790
Piping 93,000
Pumps 14.000
Equipment subtotal 391,790
Contingency and contractor's fee 50.935
Total equipment cost 442.725
Total Capital Investment $ 1.116,895
Annual Cost;
Amortization
Facility $ 50,665
Equipment 65.980
Total amortization $ 116,645
VIII-35
DRAFT
-------
DRAFT
Operation and Maintenance (O&M)
Land $ 17,675
Facility repair and maintenance 600
Equipment repair and maintenance 5,350
Tailing pond and distribution maintenance 85,435
Taxes 4,420
Insurance 11,170
Total O&M costs $ 124,650
Electricity 2.500
Total Annual Cost $ 243,795
Level Bj_ Total Recycle (Zero Discharge)
Total recycle can be attained only after impoundment systems
as described for level A have been constructed. Thus, the
costs cited for level B are the incremental costs for imple-
menting total recycle. The equipment includes decant pumps
and piping. Costs for implementing total recycle are shown in
Table VIII-7.
Capital-Cost Components and Assumptions for Level B_
Decant Pumps - water pumps - 3,900 1 (1,042 gal)/minute,
1 plus standby
Piping - Flow @ 2 meters (3.3 feet)/sec through 21-cm
(8.3-in.) pipe, 1,500 meters (4,920 feet) long
Ope rat ing-Cost As s umpt ion s for Level B:
Power - 30 kW (40 hp)
VIII-36
DRAFT
-------
DRAFT
Capital Investment:
Equipment
Piping
Pumps
Equipment subtotal
Contingency and contractor's fee
Total equipment cost
Annual Cost:
Amortization
Equipment
Total amortization
Operation and Maintenance (O&M)
Equipment repair and maintenance
Insurance
Electricity
Total O&M costs
Total Annual Cost
$ 64,500
8.000
72,500
9.425
12.210
3,625
820
$ 81.925
12,210
4,445
4.000
$ 20,655
VIII-37
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR GOLD-ORE CATEGORY
Gold Mines (Alone)
Three known mines operating alone without discharge to mill
treatment facilities exist in this subcategory, only two of
which are discharging. The range of ore mined is 163,000
to 478,000 metric tons (180,000 to 527,000 short tons)
annually. The average daily discharge for these operations
is 3,785 cubic meters (1,000,000 gallons).
A hypothetical mine with an annual ore production of 320,000
metric tons (353,000 short tons) and with a discharge of
3,785 cubic meters (1,000,000 gallons) per day was chosen
to represent this subcategory.
Two levels of technology are considered. The incremental
costs for the representative gold mine to attain levels A
and B are shown in Table VIII-8.
Wastewater Treatment/Control
Level Aj_ Sedimentation (Settling Pond)
Level A consists of a sedimentation pond with a one-day
retention. It is assumed that mine dewatering pumps already
have been installed.
The capital and operating costs and assumptions for attain-
ing this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Sedimentation pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 5,700 cubic meters
(1,506,000 gal)
Piping - Flow @ 2 meters/sec (6.6 feet) through pipe
measuring 17 cm (6.7 in.) x 1000 meters
(3,300 feet)
VIII-38
DRAFT
-------
DRAFT
TABLE VIII-8. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL) -
GOLD
SUBCATEQORv; Gold Mines (Alone)
PLANT SIZE: 320.000
PLANT AOE;N/A YEARS
METRIC TONS (353,000 SHORT TONS) PER YEAR op ore mined
PLANT LOCATION; N/A
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATINO AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
53.8
7.4
2.3
9.7
0.03
B
121.2
17.3
28.1
4.4
49.8
0.16
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Cu
Hg
In
Fe
CONCENTRATION (mg/JU (ppm)
RAW
(UN-
TREATED)
25
0.06
0.001
6
12
AFTER TREATMENT TO LEVEL
A
20
0.06
0.001
4
8
B
20
0.05
0.0005
0.15
1.0
C
D
E
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0507
LEVEL A: SEDIMENTATION (SETTLING POND)
LEVEL B: SEDIMENTATION. LIME PRECIPITATION. SECONDARY SETTLING, AND DISCHARGE
VIII-39
DRAFT
-------
DRAFT
Capital Investment;
Facilities
Lagoon $ 9,000
Contingency and contractor's fee 1»170
Total facility cost $ 10,170
Land 700
Equipment
Piping 38,000
Contingency and contractor's fee 4,940
Total equipment cost 42,940
Total Capital Investment $53,810
Annual Cost;
Amortization
Facility $ 1,035
Equipment 6_,400
Total amortization $ 7,435
Operation and Maintenance (O&M)
Land 70
Facility repair and maintenance 270
Equipment repair and maintenance 1,900
Taxes 20
Insurance 55
Total O&M costs 2.315
Total Annual Cost $ 9,750
VIII-40
DRAFT
-------
DRAFT
Level B; Sedimentation, Lime Precipitation, Secondary Sjettlingt
and Discharge
Level-B technology utilizes a sedimentation pond with a reten-
tion time of one day and a smaller settling pond with a 6-hour
retention period. The mine water has a pH of 6; thus, addition
of 0.9 kg of hydrated lime per 3.785 cubic meters (2 lb/1,000
gal) of water would raise the pH sufficiently for precipitation
of metals.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level B_:
Sedimentation pond - dike height of 3m (10 ft)
top width of 3 m (10 ft)
capacity of 5,700 cubic meters
(1,506,000 gal)
Settling pond - dike height of 4 m (13 ft)
top width of 3 m (10 ft)
capacity of 1,425 cubic meters (376,000 gal)
Land - 0.5 hectare (1.24 acres)
Lime precipitation system
Piping - Flow @ 2 meters (6.6 feet)/sec through pipe measuring
17 cm (6.7 in.) x 1,100 meters (3,600 feet)
Operating-Cost Assumptions for Level B_:
Lime - 317 metric tons (350 short tons)/year
Operating Personnel - 1 hr/shift, 3 hr/day
Power - 30 kW (40 hp)
VIII-41
DRAFT
-------
DRAFT
Capital Investment;
Facilities
Lagoon(s) $ 12.275
Facility subtotal 12,275
Contingency and contractor*s fee 1.595
Total facility cost $ 13,870
Land 875
Equipment
Lime precipitation unit 54,400
Piping 41.800
Equipment subtotal 94,200
Contingency and contractor's fee 12.245
Total equipment cost 106,445
Total Capital Investment $121.190
Annual Coat:
Amortization
Facility $ 1,415
Equipment 15,865
Total amortization $ 17,280
Operations and Maintenance (O&M)
Land 90
Operating personnel 9,450
Facility repair and maintenance 370
Equipment repair and maintenance 4,710
Materials 12,250
Taxes 20
Insurance 1,210
Total O&M Costs 28,100
Electricity A»4PP_
Total Annual Cost $49,780
VIII-42
DRAFT
-------
DRAFT
Gold Mills ojr Mine/Mills (Cyanidation Process)
There are three known mills practicing cyanidation, with one
of these operations employing both flotation and cyanidation.
The range of ore milled in this subcategory is 476,000 to
1,400,000 metric tons (527,000 to 1,550,000 short tons) per
year. The mill wastewater ranges from 490 to 22,710 cubic
meters (130,000 to 6,000,000 gallons) per day.
The representative mill has an annual production of 1,400,000
metric tons (1,550,000 short tons) per year with a daily
wastewater flow rate of 22,710 cubic meters (6,000,000 gallons).
Two levels of technology are considered. The incremental
costs of achieving these levels are shown in Table VIII-9.
Wastewater Treatment/Cont ro1
Level A:_ Total Recycle (Zero Discharge)
Total recycle for this subcategory entails use of an impound-
ment system, a distribution system, piping, and pumps.
Typically, an operation in this subcategory discharges its
entire effluent with no treatment or impoundment. Thus, the
capital and annual operating costs are high. The costs shown
in Table VIII-9 assume that adequate land is available.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level A^
Tailing pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
perimeter of 8,700 meters (28,536 ft)
Secondary settling pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 8,600 cubic meters
(2,272,000 gal)
Land - 421 hectares (1,040 acres)
Distribution system - 8,700 meters (28,536 feet) of pipe
measuring 60 cm (2 ft) in diameter
25 cyclones @ $1,980 each
VIII-43
DRAFT
-------
DRAFT
TABLE VI 1 1-9. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 4105
suBCATEOORY: Gold
or Mine/Mills (Cyanidation Process)
PLANT SIZE; 1.400,000
PLANT AGE: 9 YEARS
METRIC TONS ( 1 ,550 ,000 SHORT TONS) PER YEAR OF Ore milled
PLANT LOCATION; South Dakota
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS(S)/METRIC TON OF PRODUCT*
COSTS (€1000) TO ATTAIN LEVEL
A
3,136.7
309.5
433.6
17.5
760.6
0.54
B
3,142.6
310.4
434.6
17.5
762.5
0.55
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Cyanide
Cu
Hg
Zn
Fe
WASTE LOAD (kg/1000 metric ton* of production milled] *)
RAW
(UN-
TREATED)
3,160
0.52
17.0
0.036
2
660
AFTER TREATMENT TO LEVEL
A
0
0
0
0
0
0
B
0
0
0
0
0
0
c
D
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED), MULTIPLY COSTS SHOWN BY 0407
fTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2 0
LEVEL A: TOTAL RECYCLE (ZERO DISCHARGE)
LEVEL B: TOTAL RECYCLE (WITH ALKALINE CHLORINATION)
VIII-44
DRAFT
-------
DRAFT
Diversion ditching - around 1 length and 1 width of
tailing pond; total length of 4,350
meters (14,268 ft)
Piping - Flow at 1 m (3.3 ft)/sec through pipe measuring
60 cm (2 ft) x 1,100 meters (3,600 feet)
(tailings)
Flow at 2 m (6.6 ft)/sec through pipe measuring
41 cm (16 in.) x 1,100 meters (3,600 feet)
(recycle)
Pumps - slurry: 15.8 cubic meters (4,174 gal)/minute
water: 15.8 cubic meters (4,174 gal)/minute
Operating-Cost Assumptions for Level A_:
Power - 130 kW (175 hp)
Distribution system maintenance @ 30% of system cost
Capital Investment;
Facilities
Tailing pond $ 861,300
Lagoon 12,000
Diversion ditching 7,180
Facility subtotal 880,480
Contingency and contractor's fee 114,465
Total facility cost $ 994,945
Land 735,000
Equipment
Distribution system 989,100
Piping 202,400
Pumps 53,000
Equipment subtotal 1,244,500
Contingency and contractor's fee 161,785
Total equipment cost 1,406.285
Total Capital Investment $ 3,136.230
VIII-45
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Facility $ 101,340
Equipment 208,130
Total amortization $ 309,470
Operation and Maintenance (O&M)
Land 73,500
Facility repair and maintenance 575
Equipment repair and maintenance 13,055
Distribution system maintenance 296,730
Taxes 18,375
Insurance 31.360
Total O&M costs 433,595
Electricity 17.500
Total Annual Cost $ 760,565
Level B; Total Recycle (Zero Discharge) (with Alkaline
Chlorination)
Level B is the same as level A with the addition of alkaline
chlorination. Level-B costs are shown in Table VIII-9.
The incremental capital and operating costs and assumptions for
attaining this level via alkaline chlorination are shown below.
Capital-Cost Assumptions for Level B£
Chlorine - 6,755 kg (14,861 lb)/yr
-------
DRAFT
Capital Investment;
Equipment
Chlorinator $ 5.660
Equipment subtotal 5,660
Contingency and contractor's fee 735
Total Capital Investment $ 6,395
Annual Cost;
Amortization $ 9*5
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 285
Materials 745
Insurance 5_
Total O&M costs 1*035
Total Annual Cost $ 1.980
Gold Mills (Amalgamation Process)
One known mill utilizes the process of amalgamation. It mills
163,000 metric tons (180,000 short tons) yearly and discharges
2,271 cubic meters (600,000 gallons) of wastewater daily. Three
levels of technology are considered. The total costs of achieving
these levels are shown in Table VIII-10.
Wastewater Treatment/Control
Level Aj_ Sulfide Precipitation, Lime Precipitation, and Discharge
The mill in this subcategory has adequate impoundment systems
for sedimentation purposes. To achieve level A, further treat-
ment is necessary.
The addition of 0.9 kg of hydrated lime per 3.785 cubic meters
(2 lb/1000 gal) and 1.5 mg/1 of sodium sulfide is recommended for
achieving level A.
VIII-47
DRAFT
-------
DRAFT
TABLE VIII-10. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 4102
SUBCATEQORY; Gold Mills (Amalgamation Process)
PLANT SIZE; 163,000 METRIC TONS (180,000 SHORT TONS) PER YEAR QFOre milled
PLANT APE; 45 YEARS
PLANT LOCATION; Colorado
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS (SI/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
45.3
6.7
22.7
2.0
31.4
0.19
B
213.5
31.8
12.8
-
44.6
0.27
C
41.5
6.2
1.9
1.5
9.6
0.06
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Cyanide
Cu
HE
Zn
Fe
WASTE LOAD (kg/1000 metric tons of product (ore milled] T)
RAW"
(UN-
TREATED)
2,871
__
3.1
0.0064
7.5
8.7
AFTER TREATMENT TO LEVEL
A
120
_-
0.29
0.003
0.87
5.8
B
120
0.0lt1
0.29
<0.0001
0.87
5.8
C
0
0
0
o
o
0
D
E
*ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED), MULTIPLY COSTS SHOWN BY 0.907
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
••HYPOTHETICAL - BASED ON OPERATIONS VISITED IN THE SUBCATEGORY
"PRESENCE OF CYANIDE DUE TO PROCESS CHANGE
LEVEL A: SULFIDE PRECIPITATION. LIME PRECIPITATION. AND DISCHARGE
LEVEL B: PROCESS CHANGE FROM AMALGAMATION TO CYAN I DAT I ON
LEVEL C: TOTAL RECYCLE (ZERO DISCHARGE)
VIII-48
DRAFT
-------
DRAFT
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Sodium sulfide distribution system - consists of a drum
with a capacity of 0.2 cubic meter (55 gallons),
mounted near the effluent stream. A solution is
mixed daily by adding the proper amount of water to
the sodium sulfide and allowing it to drip into the
effluent stream.
Lime precipitation system
Operating-Cost Assumptions for Level A:
Sodium sulfide - 1,192 kg (2,627 lb)/year
Lime - 190 metric tons (210 short tons)/year
Operating personnel - 1 hr/shift, 2 hr/day for lime
operator
1 hr/day for sulfide operator
Capital Investment;
Equipment
Lime precipitation unit $ 40,000
Sulfide precipitation unit 100
Equipment subtotal 40,100
Contingency and contractor's fee 5.210
Total equipment cost $ 45,310
Total Capital Investment $ 45,310
VIII-49
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Equipment $ 6,735
Total amortization $ 6,735
Operation and Maintenance (O&M)
Operating personnel 12,600
Equipment repair and maintenance 2,000
Materials 7,610
Insurance 450
Total O&M costs 22,660
Electricity 2.000
Total Annual Cost $ 31,395
Level B^; Process Change from Amalgamation to Cyanidation
An alternative to precipitation for this subcategory would be
to change the milling process from amalgamation to cyanidation.
The costs incurred for this process change are difficult to
obtain and estimate. However, data were provided for a similar
change for an operation whose mill-circuit volume is 10 times
greater than the one in this subcategory. 'To estimate the cost
for the process change, an application of the six-tenths-factor
rule was used.
Note that a mill with a water flow of 22,710 cubic meters
(6,000,000 gal)/day incurred a capital investment cost of
$850,000 for the process change. Applying the six-tenths-
factor rule to an operation whose water flow is 2,271 cubic
meters (600,000 gal)/day resulted in a capital investment cost
of $213,510. No assumptions were made as the the amounts
of materials, operating labor, and power that would be re-
quired, as these data are not available. Equipment repair and
maintenance were assumed to total 5 percent of capital invest-
ment. Amortization was assumed over a 10-year period. The
costs are shown in Table VIII-10.
The capital and operating costs for attaining this level are
shown below.
VIII-50
DRAFT
-------
DRAFT
Capital Investment;
Equipment
Process change $ 213,510
Annual Coat;
Amortization $ 31,820
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 10,675
Insurance 2,135
Total O&M costs 12.810
Total Annual Cost $ 44>630
Level C^ Total Recycle (Zero Discharge)
To achieve total recycle, additional pumps and piping would be
necessary to recirculate the wastewater. The capital and opera-
ting cost components and assumptions for attaining this level are
shown below.
Capital-Cost Components and Assumptions for Level C;
Piping - Flow @ 2 meters (6.6 feet)/second through pipe
measuring 13 cm (5.1 in.) x 1000 meters (3,300 feet)
Pumps - water pumps with capacity of 15.77 cubic meters
(4,166 gal)/minute
Operating-Cos t Assumptions for Level C:
Power - 11.2 kW (15 hp)
VIII-51
DRAFT
-------
DRAFT
Capital Investment:
Equipment
Piping $ 32,000
Pumps 4,700
Equipment subtotal 36,700
Contingency and contractor's fee 4,770
Total Capital Investment $ 41,470
Annual Cost;
Amortization $ 6,170
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 1,835
Insurance 40
Total O&M costs 1,875
Electricity 1,500
Total Annual Cost $ 9,545
Gold Mills (Flotation)
The one mill which exists in this subcategory processes 478,000
metric tons (527,000 short tons) of ore annually. The flow from
the mill is 490 cubic meters (130,000 gallons) per day. A dis-
charge from the tailing pond occurs for only two months of the
year and amounts to 545 cubic meters (144,000 gallons) per day.
Two alternative treatment levels are considered. The costs of
achieving these levels are shown in Table VIII-11.
Wastewater Treatment/Control
Level A.; Diversion Ditching, Lime Precipitation, and Alkaline
Chlorination
Adequate impoundment systems exist for the mill in this sub-
category. Lime precipitation is recommended for the precipitation
VIII-52
DRAFT
-------
DRAFT
TABLE VI11-11. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 4104
SUBCATEOORV: Gold Mills (Flotation)
PLANT SIZE; 478.000
PLANT APE; 39 YEARS
__METRIC TONS (527.000 SHORT TONS) PER YEAR OF Ore nil led
PLANT LOCATION; Washington
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS (SI/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
20.3
3.5
12.1
1.0
16.6
0.035
B
31.2
4.5
12.6
1.0
18.1
0.037
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Cyanide
Hg
Cu
Zn
Fe
WASTE LOAD (kg/1000 metric torn of product [ore milled] f)
RAW **
(UN-
TREATED)
2,000
2.4
0.0001
0.02
0.15
0.6
AFTER TREATMENT TO LEVEL
A
6
0.001
0.0001
0.01
0.045
0.3
B
0
0
0
0
0
0
c
D
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0407
tTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2JO
••HYPOTHETICAL • BASED ON OPERATIONS VISITED IN SUBCATEGORY
LEVEL A: DIVERSION DITCHING. LIME PRECIPITATION, AND ALKALINE CHLORINATION
LEVEL B: LEVEL A PLUS SETTLING POND - NO DISCHARGE
VIII-53
DRAFT
-------
DRAFT
of metals. The recommended dosage is 0.9 kg of hydrated lime
per 3.785 cubic meters (2 lb/1000 gal) of wastewater. Control
is also needed to divert seasonal runoff that results in tailing-
pond overflow.
Cyanide is used in the flotation process. Should an accidental
discharge occur, chlorination of the cyanide solution would be
necessary. The amount of chlorine needed would depend upon
the amount of cyanide in the wastewater. Since discharge of
cyanide is not a typical occurence, no estimate of the amount
of chlorine has been made.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level Aj^
Diversion ditching - total of 1000 meters (3,280 feet)
Alkaline chlorinator - V-notch type; data supplied from
surveyed operation
Lime precipitation - 15-day supply of lime slurry.
Mix tank with capacity of 7.4 cubic
meters (1,955 gal) for slurry storage.
Mix tank with capacity of 5.2-cublc
meters (1,374 gal) for 15-minute
retention.
Slurry pump - 0.34 cubic meter (90 gal)/minute
Operating-Cost Assumptions for Level Aj_
Lime - 41 metric tons (46 short tons)/year
Operating personnel - 3 hr/day
Power - 7.5 kW (10 hp)
VIII-54
DRAFT
-------
DRAFT
Capital Investment;
Facilities
Diversion ditching $ 1,650
Contingency and contractor's fee 215
Total facility cost $ 1,875
Equipment
Lime precipitation unit 6,400
Aklaline chlorinator 5,660
Pumps 4,200
Equipment subtotal 16,260
Contingency and contractor's fee 2,115
Total equipment cost 18,375
Total Capital Investment $ 20.250
Annual Cost;
Amortization
Facility $ 190
Equipment 3,315
Total amortization $ 3,505
Operation and Maintenance (O&M)
Operating personnel 9,450
Facility repair and maintenance 50
Equipment repair and maintenance 815
Materials 1,610
Insurance 200
Total O&M costs 12,125
Electricity 1,000
Total Annual Cost $ 16,630
Level B; Level A_ plus Settling Pond - No Discharge
To avoid discharge of the seasonal runoff, an additional settling
pond will be necessary. The runoff would be collected in the
VIII-55
DRAFT
-------
DRAFT
settling pond and stored for use as mill process water.
A five-day retention time is assumed.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level B;
Pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 5,700 cubic meters (1,506,000 gal)
Land - 0.4 hectare (1 acre)
Capital Investment!
Facilities
Lagoon $ 9,000
Contingency and contractor's fee 1,170
Total facility cost $ 10,170
Land 700
Total Capital Investment $ 10.870
Annual Cost;
Amortization $ 1,035
Operation and Maintenance (O&M)
Land $ 70
Facility repair and maintenance 270
Taxes 20
Insurance 110
Total O&M costs 470
Total Annual Cost $ 1.505
VIII-56
DRAFT
-------
DRAFT
Gold Mine/Mills Employing Gravity Separation
There are 58 known washing facilities operating in conjunction
with the 68 known placer mining operations. The amount of
material washed at these facilities totals 698,445 cubic meters
(913,000 cubic yards) per year. The wastewater flow is 11,355
to 15,140 cubic meters (3,000,000 to 4,000,000 gallons) per day.
A hypothetical operation based on an arithmetric average of the
68 operations was selected as representative for this subcategory.
The annual material handled for the representative operation is
10,270 cubic meters (13,425 cubic yards). Assuming a specific
gravity of 2.65 for this material, the total weight handled is
27,215 metric tons (30,000 short tons) each year. The assumed
daily water flow is 13,247 cubic meters (3,500,000 gallons).
Four alternative levels of technology are considered.
The capital and operating costs of achieving these levels are
shown in Table VIII-12.
Wastewater Treatment/Control
Level Aj_ Settling Pond
The recommended treatment system for level A consists of a settling
pond for removal of suspended solids. The capital and operating
costs and assumptions for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level A:
Settling pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 7,380 cubic meters (1,950,000 gal)
Land - 0.4 hectare (1 acre)
VIII-57
DRAFT
-------
DRAFT
TABLE VIII-12. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR TYPICAL
follNE/MILL (HYPOTHETICAL)- GOLD/GRAVITY
8UBCATEQORY; Gold Mine/Mills Employing Gravity Separation
PLANT SIZE; 27,215 METRIC TONS( 30,000 SHORT TONS) PER YEAR OF Ore milled
PLANT AOEJVA YEARS PLANT LOCATION: N/A
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS ($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
12.9
1.2
0.6
1.8
0.066
B
34.4
5.1
9.5
4.0
18.6
0.68
C
47.3
6.3
10.1
4.0
20.4
0.75
D
57.5
7.8
40.5
4.1
52.4
1.93
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
WASTE LOAD (kg/1000 metric tons of product (ore milled) f)
RAW
(UN-
TREATED)
3.600
AFTER TREATMENT TO LEVEL
A
36
B
36
c
33
D
30
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.907
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED), MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: SETTLING POND
LEVEL B: DISTRIBUTION SYSTEM
LEVEL C: SETTLING POND AND DISTRIBUTION SYSTEM
LEVEL D: SETTLING POND, DISTRIBUTION SYSTEM, AND FLOCCULATION
VIII-58
DRAFT
-------
DRAFT
Capital Investment;
Facilities
Lagoon $ 10,800
Contingency and contractor's fee 1,405
Total facility cost $ 12,205
Land 700
Total Capital Investment $ 12,905
Annual Cost:
Amortization $ 1,245
Operation and Maintenance
Land $ 70
Facility repair and maintenance 325
Taxes 20
Insurance 130
Total O&M costs 545
Total Annual Cost $ 1,790
Level B: Distribution System
An alternative to level-A treatment would be to construct and
utilize a process-water distribution system. The purpose would
be to deliver dredge wastewater to all mine workings for filtra-
tion. The capital and operating costs and assumptions for
attaining this level are shown below.
Capital-Cost Components and Assumptions for Level Bji_
Piping - Flow @ 1 m (3.3 ft)/sec through pipe measuring
45 cm (17.7 in.) x 100 meters (330 feet)
Pumps - slurry type (plus one standby)
Operating-Cost Assumptions for Level Bj_
Power - 30 kW (40 hp)
Distribution system maintenance @ 30% of system capital cost
VIII-59
DRAFT
-------
DRAFT
Capital Investment;
Equipment
Piping $ 8,400
Pumps 22.000
Equipment subtotal 30,400
Contingency and contractor's fee 3.950
Total Capital Investment $ 34,350
Annual Cost;
Amortization $ 5,120
Operation and Maintenance (O&M)
Distribution system maintenance $
Insurance
Total O&M costs
Electricity
Total Annual Cost $ 18,585
Level Cj_ Settling Pond and Distribution System
Level C is the sum of levels A and B. Total invested capital
and annual operating costs for this level are shown in Table
VIII-12.
Level V}_ Settling Pond, Distribution System, and Flocculation
Level D is the same as level C plus the addition of a floccu-
lant for further suspended-solid removal. It is assumed that
2 mg/1 of flocculant is added. A simple flocculant feed system
is all that is needed. The incremental capital and operating
costs and assumptions for this system are shown below.
The total system cost is shown in Table VIII-12.
VIII-60
DRAFT
-------
DRAFT
Capital-Cost Components and Assumptions for Level D;
Flocculant feed system
Operating-Cost Assumptions for Level Dj_
Operating personnel - 3 hr/day
Flocculant - 9,267 kg (20,430 lb)/year
Power - 0.75 kW (1 hp)
Capital Investment;
Equipment
Flocculant feed system $ 9,000
Contingency and contractor's fee It 170
Total Capital Investment $ 10,170
Annual Cost;
Amortization $ 1,515
Operation and Maintenance (O&M)
Operating personnel $ 9,450
Equipment repair and maintenance 450
Materials 20,430
Insurance 100
Total O&M costs $ 30,430
Electricity 100
Total Annual Cost $ 32,045
VIII-61
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR SILVER-ORE CATEGORY
Silver-Ore Mines
There are five known major silver mines in operation. The
range of ore mined is 75,280 to 1,428,000 metric tons (83,000
to 1.574,000 short tons) annually. The mine wastewater
ranges from 246 to 4,920 cubic meters (65,000 to 1,300,000
gallons) dally.
Three of these mines are associated with mills. The remaining
two are mines alone.
A hypothetical mine, based on an arithmetic average of the
five known mines, was selected as representative for this
subcategory. The annual ore mined is 181,400 metric tons
(200,000 short tons). The average daily discharge amounts to
1,700 cubic meters (450,000 gallons). Three levels of tech-
nology are considered. The total costs of achieving these
levels are shown in Table VIII-13.
Wastewater Treatment/Control
Level A_^ Sedimentation (Settling Pond)
It is assumed that a typical silver mining operation has little
or no effluent treatment or control. Level-A technology
requires the construction of a settling pond with a 10-day
retention capacity and adequate piping. No costs are shown
for pumps, since mine dewatering facilities are already installed.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level Aj^
Settling pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 25,500 cubic meters (6,736,000
gallons)
Land - 1.3 hectares (3.2 acres)
Piping - Flow @ 2 m (6.6 ft)/sec through pipe measuring
12 cm (4.8 in.) x 1000 meters (3,280 feet)
VIII-62
DRAFT
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DRAFT
TABLE VI11-13. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL)-
SILVER
Silver-Ore Mines
SUBCATEOORV:
PLANT SIZE: 181.400
PLANT AOE;N/A YEARS
METRIC TONS (2 00, OOP
PLANT LOCATION:
SHORT TONS) PER YEAR OF ore mined
a. COSTS OP TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS(S)/METRIC TON OF PRODUCT*
COSTS 1*1000) TO ATTAIN LEVEL
A
65.6
8.0
3.0
11.0
0.06
B
114.6
15.0
20.5
2.0
37.5
0.21
c
114.7
15.0
23.4
2.0
40.4
0.22
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
As
Cu
Fe
Pb
Sb
Zn
CONCENTRATION (mg/£) (ppm)
RAW
(UN-
TREATED)
25
0.1
0.1
3.0
0.2
1.0
0.2
AFTER TREATMENT TO LEVEL
A
20
0.09
0.09
2.5
0.19
1.0
0.19
B
20
0.07
0.05
1.0
0.1
1.0
0.15
C
20
0.07
0.05
1.0
0.1
0.5
0.15
D
E
"ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0.907
LEVEL A: SEDIMENTATION (SETTLING POND)
LEVEL B: SEDIMENTATION. LIME PRECIPITATION. AND SECONDARY SETTLING
LEVEL C: LEVEL B PLUS SULFIDE PRECIPITATION
VIII-63
DRAFT
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DRAFT
Capital Investment;
Facilities
Lagoon $ 26,000
Contingency 3,380
Total facility cost $ 29,380
Land 2,275
Equipment
Piping 30,000
Contingency and contractor's fee 3,900
Total equipment cost 33,900
Total Capital Investment $ 65.555
Annual Cost;
Amortization
Facility $ 2,990
Equipment 5.050
Total amortization $ 8,040
Operation and Maintenance (O&M)
Land 20
Facility repair and maintenance 780
Equipment repair and maintenance 1,500
Taxes 55
Insurance 655
Total O&M costs 3.010
Total Annual Cost $ 11,050
Level Bj^ Sedimentation, Lime Precipitation, and Secondary
Settling
The incremental cost to achieve level B is the cost for a lime
precipitation system, additional piping, and a secondary settling
pond. The costs associated with sedimentation are shown under
Level A.
VIII-64
DRAFT
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DRAFT
The recommended treatment consists of the addition of 0.9 kg
of hydrated lime per 3.785 cubic meters (2 lb/1000 gallons) of
mine wastewater. The mine wastewater is then retained for one
day in a settling pond before discharge. The incremental
capital and operating costs and assumptions for attaining
level B are shown below. The total system cost is shown in
T*ble VIII-13.
Capital-Cost Components and Assumptions for Level B;
Lime precipitation system
Piping - Flow @ 2 m (6.6 ft)/sec through pipe measuring
12 cm (4.7 in.) x 100 meters (328 feet)
Settling pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 2,550 cubic meters (674,000 gal)
Land - 0.21 hectare (0.5 acre)
Operating-Cost Assumptions for Level B;
Lime - 142 metric tons (157.5 short tons)/year
Power - 14.9 kW (20 hp)
Operating personnel - 3 hr/day
Capital Investment;
Facilities
Lagoon $ 5,100
Contingency and contractor's fee 665
Total facility cost $ 5,765
Land 365
VIII-65
DRAFT
-------
DRAFT
Equipment
Lime precipitation system
Piping
Equipment subtotal
Contingency and contractor's fee
Total equipment cost
Total Capital Investment
35,000
3,000
38,000
4 . 940
$ 42,940
$ 49,070
Annual Cost:
Amortization
Facility $
Equipment
Total amortization
Operation and Maintenance (O&M)
Land 35
Operating personnel 9,450
Facility repair and maintenance 155
Equipment repair and maintenance 1,900
Materials 5,510
Taxes
Insurance
Total O&M costs
Electricity
Total Annual Cost
Level C_^ Level B_ plus Sulfide Precipitation
Level-C technology includes the addition of sodium sulfide plus
level-B technology.
Further removal of metals is attained by the addition of 2 mg/1
of sodium sulfide. The incremental capital and operating costs
and assumptions for sulfide precipitation are shown below. The
total cost to achieve level C is shown in Table VII1-13.
17,550
2>000
$ 26,535
VIII-66
DRAFT
-------
DRAFT
Capital-Cost Components and Assumptions for Level C;
Sulfide precipitation system
Operating-Cost Assumptions for Level C:
Sodium sulfide - 1,191 kg (2,625 lb)/year
Operating personnel - 1 hr/day
Capital Investment;
Equipment
Sulfide precipitation system $
Contingency and contractor's fee
Total Capital Investment $ 115
Annual Cost;
Amortization 15
Operation and Maintenance (O&M)
Operating personnel
Equipment repair and maintenance
Materials
Total O&M costs
Total Annual Cost
Silver Mills Employing Cyanidation, Amalgamation. Gravity
Separation, and Byproduct Recovery
Five subcategories based on milling process have been identified
for the silver milling industry. The subcategories are essen-
tially identical to those of the gold industry. Four of the
silver milling subcategories (cyanidation, amalgamation, gravity
separation, and byproduct recovery) are represented by the same
operation and require the same control and treatment technology
as the gold milling industry. The capital and annual operating
costs of implementing the required treatment technologies for
these subcategories are shown in Tables VIII-9, VIII-10, and
VIII-12.
VIII-67
DRAFT
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DRAFT
The remaining subcategory and applicable treatment technologies
are identified in the section which follows.
Silver Mills Employing Flotation Process
There are four major mills in this subcategory. These mills
mill ore in the range of 75,280 to 182,300 metric tons (83,000
to 201,000 short tons) annually. Daily wastewater flow from
these mills ranges from 1,500 to 3,160 cubic meters (396,000
to 835,000 gallons).
An existing flotation mill which mills 180,000 metric tons
(200,000 short tons) of ore and has a daily water flow rate
of 3,160 cubic meters (835,000 gallons) was selected as a repre-
sentative operation. Typically, mills in this subcategory
recycle 70 percent of their wastewater and discharge the
remaining 30 percent.
One level of technology is considered. The cost of implementing
this level is shown in Table VIII-14.
Wastewater Treatment/Control
Level A; Total Recycle (No Discharge)
Total recycle for this subcategory entails the implementation
of additional pumps and pipes to recirculate the effluent that
is normally discharged. In this case, it is approximately 946
cubic meters (250,000 gallons) a day. Also, diversion ditching
is recommended to avoid tailing-pond overflow resulting from
seasonal runoff.
Capital-Cost Components and Assumptions for Level Aj_
Piping - Flow @ 1 m (3.3 ft)/sec through pipe measuring
11 cm (4.3 in.) in diameter
Water pumps - 0.66 cubic meter (174 gal)/minute
VIII-68
DRAFT
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DRAFT
TABLE VIII-14. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 4401
8UBCATEOOHY; Silver Mills Employing Flotation Process
PLANT SIZE; 180,000 METRIC TONS (200.000 SHORT TONS) PER YEAR OF Ore milled
PLANT AGE:
23
YEARS
PLANT LOCATION; Idaho
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
39.0
5.7
2.1
0.3
8.1
0.045
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Cyanide
As
Cu
Hg
Pb
Sb
Zn
Fe
WASTE LOAD (kg/1000 metric tons of product (ore milled]*)
RAW"
(UN-
TREATED)
8,400
0.02
<0.3
1.0
0.6
6.2
8.3
4.1
8.3
AFTER TREATMENT TO LEVEL
A
0
0
0
0
0
0
0
0
0
B
C
D
E
"ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED), MULTIPLY COSTS SHOWN BY 0.807
fTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
••HYPOTHETICAL - BASED ON OPERATIONS VISITED IN THE SUBCATEGORY
LEVEL A: TOTAL RECYCLE (ZERO DISCHARGE)
VIII-69
DRAFT
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DRAFT
Diversion ditching - 1000 meters (3,300 feet) long
Operating-Cost Assumptions for Level A_j_
Power - 2.2 kW (3 hp)
Capital Investment:
Facilities
Diversion ditching $ If650
Contingency and contractor's fee 215
Total facility cost $ 1,865
Equipment
Piping 30,000
Pumps 2.900
Equipment subtotal 32,900
Contingency and contractor's fee 4,260
Total equipment cost 37,180
Total Capital Investment $ 39.045
Annual Cost;
Amortization
Facility $ UO
Equipment 5.540
Total amortization $ 5,730
Operation and Maintenance (O&M)
Facility repair and maintenance 50
Equipment repair and maintenance 1,645
Insurance __ 390
Total O&M costs 2,085
Electricity —300
Total Annual Cost $ 8,115
VIII-70
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR BAUXITE CATEGORY
Bauxite Mines
There are currently two bauxite mines in operation in the U.S.
Both operations treat a portion of their mine drainage with
lime and then allow the effluent to settle in a series of ponds.
Of the two sites (both visited), one was chosen as the industry
representative. Note that mines in this subcategory typically
have more than one discharge, and some of these discharges are
treated. The remaining wastewater is discharged directly to
nearby streams. It has been recommended that all discharges
be treated.
The representative mine produces 861,650 metric tons (950,000
short tons) of ore yearly. The average untreated mine drainages
for the representative operation consist of three discharges with
flow rates of 17,000, 7,570, and 3,785 cubic meters (4,500,000,
2,000,000, and 1,000,000 gallons, respectively) per day into
pits. Each discharge must be treated separately because of the
great distance between each pit. One level of technology is
considered for this subcategory. The incremental cost of
implementing this level is shown in Table VIII-15.
Wastewater Treatment/Control
Level A_L Lime Precipitation and Secondary Settling
The typical bauxite mine has dewatering pumps, pipes, and primary
settling ponds. The installation of additional piping, a
lime precipitation system, and secondary settling ponds for
each discharge is needed to achieve level A.
The addition of 0.9 kg of hydrated lime per 3.785 cubic meters
of mine water (2 lb/1000 gallons), followed by a 2-day retention
in the secondary settling ponds, is considered adequate treat-
ment for this subcategory.
The capital and operating costs and assumptions for attaining
this level are shown below.
VIII-71
DRAFT
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DRAFT
TABLE VIII-15. WATER EFFLUENT TREATMENT COSTS AMD RESULTING
WASTE-LOAD CHARACTERISTICS FOR MINE 5t32
Bauxite Mines
SUBCATEQORV:
PLANT SIZE; 861,650
PLANT APE; 7S YEARS
_METHIC TON81 950.000 SHORT TONSI PER YEAR OF ore mined
PLANT LOCATION! Arkansas
•. COSTS OP TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS (S1000) TO ATTAIN LEVEL
A
383.2
51.7
149.5
25.3
226.5
0.26
B
t
t
t
t
t
t
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Al
Fe
Mn
Ni
t
Zn
CONCENTRATION (mo/A) Cppm)
RAW
(UN-
TREATED!
161.0
47.8
39.2
7.5
0.11
0.23
.
AFTER TREATMENT TO LEVEL
A
20
0.35
0.30
1.5
0.10
0.05
B
20
0.30
0.30
0.50
0.05
0.05
C
D
E
*OHE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0407
*NO COST DIFFERENCE
LEVEL A: LIME PRECIPITATION AND SECONDARY SETTLING
LEVEL B: LIME PRECIPITATION AND SECONDARY SETTLING WITH OPTIMUM pH CONTROL
VIII-72
DRAFT
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DRAFT
Capital-Cost Components and Assumptions for Level A;
Three lime precipitation units -
17,000 cubic meters (4,500,000 gal)/day
7,570 cubic meters (2,000,000 gal)/day
3,785 cubic meters (1,000,000 gal)/day
Three secondary settling ponds -
all have dike height of 3 m (10 ft) and are 3
meters (10 ft) wide
capacities of 50,000 cubic meters (13,209,000 gal)
25,000 cubic meters (6,604,000 gal)
12,000 cubic meters (3,170,000 gal)
Piping - Flow @ 2 m (6.6 ft)/sec through pipes measuring:
36 cm (14 in.) x 100 meters (328 feet)
24 cm (9.4 in.) x 100 meters (328 feet)
17 cm (6.7 in.) x 100 meters (328 feet)
Land - 4.3 hectares (10.6 acres)
Operating-Cost Assumptions for Level A;
Lime - 2,380 metric tons (2,625 short tons)/year
Power - 186 kW (250 hp)
Operating personnel - 3 hr/day/unit = 12 hr/day
Capital Investment:
Facilities
Lagoon(s) $ 80,200
Contingency and contractor's fee 10.425
Total facility cost $ 90,625
Land 7,525
VIII-73
DRAFT
-------
DRAFT
Equipment
Litne precipitation units $ 236,650
Piping 15.600
Equipment subtotal 252,250
Contingency and contractor's fee 32.795
Total equipment cost $ 285.045
Total Capital Investment $ 383.195
Annual Cost:
Amortization
Facility $ 9,230
Equipment 42.480
Total amortization $ 51,710
Operation and Maintenance (O&M)
Land 750
Operating personnel 37,800
Facility repair and maintenance 2,405
Equipment repair and maintenance 12,615
Materials 91,875
Taxes 190
Insurance 3,830
Total 08M costs $ 149,465
Electricity 25«365
Total Annual Cost $ 226,540
VIII-74
DRAFT
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DRAFT
WASTEWATER-TREATMENT COSTS FOR FERROALLOY-ORE CATEGORY
Ferroalloy-Ore Mines
There are seven ferroalloy mines in this subcategory. The
annual ore production ranges from 16,560 to 14,000,000 metric
tons (18,220 to 15,500,000 short tons). The range of daily
wastewater discharged is 0 to 51,840 cubic meters (0 to
13,700,000 gallons).
A hypothetical mine, based on the industry average, was selected
as representative. This mine is assumed to have an annual
ore production of 1,800,000 metric tons (1,990,000 short tons),
with a daily discharge of 3,275 cubic meters (865,000 gallons).
The current level of technology for this subcategory includes
flocculation, neutralization, and settling or clarifying. A
further level of technology has been recommended. The total
costs of achieving this level are shown in Table VIII-16.
Wastewater Treatment/Control
Level A; Lime Precipitation and Secondary Settling
The necessary equipment includes a lime precipitation unit and
a settling pond. The addition of 0.9 kg of hydrated lime per
3.785 cubic meters (2 lb/1000 gallons) of wastewater is con-
sidered sufficient for precipitation of metals. The wastewater
is then retained for one day in a settling pond before discharge.
The capital and operating costs and assumptions for attaining
this, level are shown below.
Capital-Cost Components and Assemptions for Level A^
Lime precipitation system
Settling pond - dike height of 3 meters (10 feet)
top width of 3 meters (10 feet)
capacity of 4,900 cubic meters (1,295,000 gal)
Land - 0.35 hectare (0.85 acre)
Piping - Flow @ 2 meters (6.6 feet)/second through pipe
measuring 16 cm (6.3 in.) x 100 meters (328 ft)
VIII-75
DRAFT
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DRAFT
TABLE VIII-16. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL) -
FERROALLOY
SUBCATEOORY; Ferroalloy-Ore Mines
PLANT SIZE: 1.800,000
PLANT AQE:N/A YEARS
.METRIC TONS ( 1.990,000 SHORT TONS} PER YEAR OF ore mined
PLANT LOCATION: N/A
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTSI$)/METRIC TON OF PRODUCT*
COSTS (SIMM) TO ATTAIN LEVEL
A
93.8
14.0
25.1
12. S
51.6
0.028
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
V
Zn
Fluoride
CONCENTRATION (mg/i)
RAW
(UN-
TREATED)
50
5
1
0.07
0.5
5
2
0.15
2
0.3
15
AFTER TREATMENT TO LEVEL
A
20
3
0.4
0.05
0.05
1.0
0.5*
0.1
1.0*
0.15
10
B
C
D
6
'
"ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0*07
fBATEA ONLY (NO LIMIT FOR BPCTCA SET)
LEVEL A: LIME PRECIPITATION AND SECONDARY SETTLING
VIII-76
DRAFT
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DRAFT
Operating-Cost Assumptions for Level A;
Lime - 275 metric tons (302 short tons)/year
Operating personnel - 3 hr/day
Power - 32 kW (43 hp)
Capital Investment;
Facilities
Lagoon $ 8,000
Contingency and contractor's fee 1,040
Total facility cost $ 9,040
Land 615
Equipment
Lime precipitation unit 49,000
Piping 3.700
Equipment subtotal 52,700
Contingency and contractor's fee 6.850
Total equipment cost 59,550
Total Capital Investment $ 69,205
Annual Cost;
Amortization
Facility $ 920
Equipment 8>875
Total amortization ? 9,795
Operation and Maintenance (O&M)
Land 60
Operating personnel 9,450
Facility repair and maintenance 240
Equipment repair and maintenance 2,635
Materials 10,570
Taxes ^
Insurance ^90
Total OiiM costs 23,660
Electricity A >32°
Total Annual Cost $ 37,775
VI1I-77
DRAFT
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DRAFT
Ferroalloy Mine/Mills Annually Processing Less Than 5,000
Metric Tons (5,500 Short Tons) Ore By_ Methods Other Than
Ore Leaching
There are 60 mills in this subcategory. All are located in the
western U.S. The annual amount of ore milled ranges from 0 to
5,000 metric tons (0 to 5,500 short tons). The daily wastewater
flow ranges from 0 to 1,872 cubic meters (0 to 500,000 gallons).
Mills in this subcategory are small and are considered to
operate 100 days a year. The mine associated with each mill is
assumed to discharge 350 days and to require treatment year-round.
A hypothetical mill with an annual milling capacity of 5,000
metric tons (5,500 short tons) and with a daily wastewater flow
of 164 cubic meters (43,200 gallons) is representative for this
subcategory.
Two levels of technology are considered. The costs of achieving
these levels are shown in Table VIII-17.
Wastewater Treatment/Control
Level A_r_ Lime Neutralization and Use of_ Tailing Pond
The facilities necessary to achieve this level include a tailing
pond and a distribution system. Also required are a lime
neutralization system, piping, and pumps. The addition of 0.9
kg of hydrated lime per 3.785 cubic meters (2 lb/1000 gallons)
of wastewater is recommended for adequate neutralization.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Tailing pond - dike height of 2 m (7 ft)
top width of 3 m (LO ft)
perimeter of 1,062 m (3,483 ft)
Land - 6.25 hectares (8.7 IIL-IVM)
VIII-78
DRAFT
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DRAFT
TABLE VII-17. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-LOAD
CHARACTERISTICS FOR TYPICAL MINE/MILL (HYPOTHETICAL)-
FERROALLOY/LIMITED
Ferroalloy Mine/Mills Annually Processing Less than 5,000 Metric
SUBCATEGORY: Tons (5,512 Short Tons) Ore by Methods Other Than Ore Leaching
PLANT SIZE: 5.000
PLANT ACE:N/A YEARS
METRIC TONS i 5 , 512 SHORT TONS! PER YEAR OF ore mil
PLANT LOCATION: N/A
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTSISI/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
100.5
10.3
14.0
0.3
24.6
4.92
B
103.'J
10.8
15.3
0.6
26.7
5.34
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
i
WASTE LOAD (kg/1000 metric tons of product (ore milled) *>
RAW
(UN-
TREATED)
1 x 106
AFTER TREATMENT TO LEVEL
A
90
B
75
c
D
E
"ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.907
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: LIME NEUTRALIZATION AND USE OF TAILING POND
LEVEL B- LEVEL A PLUS FLOCCULATION
VII1-79
DRAFT
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DRAFT
Distribution system - 1,062 meters (3,483 feet) of pipe
measuring 5 cm (2 in.) in diameter
2 cyclones @$630 each
Lime neutralization system - 1 mix tank with capacity of
1,700 liters (449 gallons) to make up
1-day supply of line slurry
Piping - Flow @ 1 m (3.3 ft)/sec through pipe measuring 5 cm
(2 in.) x 100 meters (328 ft)
Pumps - one slurry type with capacity of 114 1 (30 gal)/min
Operating-Cost Assumptions for Level A;
Lime - 13.6 metric tons (IS short tons)/year
Distribution system maintenance @ 30% of distribution
system capital cost
Operating personnel - 1 hr/day
Power - 2.25 kW (3 hp)
Capital Investment:
Facilities
Tailing pond $ 57,350
Contingency and contractor's fee 7,455
Total facility cost § 64,805
Land 10,940
Equipment
Distribution system
Lime neutralization unit
Piping
Pumps
Equipment .subtotal
Contingency and contractor's IVi:
Total equ I pmcnt t-ust -''. '.'
Total Capital Investment $ J 'V *>
VIIJ-80
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Facility $ 6,600
Equipment 3,695
Total amortization $ 10,295
Operation and Maintenance (O&M)
Land 1,095
Operating personnel 3,150
Facility repair and maintenance 1,720
Equipment repair and maintenance 1,100
Materials 525
Distribution system maintenance 5,155
Taxes 275
Insurance 1,005
Total O&M costs 14,025
Electricity 300
Total Annual Cost $ 24,620
Level B; Level A_ plus Flocculation
In addition to level-A treatment, flocculation would be necessary.
However, this would be needed only 100 days a year.
A full-day supply of flocculant, in a 0.1 percent solution that
is prepared daily, is fed to the wastewater stream at a rate of
5 mg/1. Total cost of level-B treatment is shown in Table VIII-17.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Componen. ts and AauumpL Juiis for l.cvu 1 R_[_
Flocculation - mix tank with c-,ipm-Lty of 830 I (224 ga] )
metering pump
VIII-81
DRAFT
-------
DRAFT
Operating-Cost Assumptions for Level Bj_
Flocculant - 81.6 kg (180 lb)/year
Operating personnel - 1 hr/day for 100 days
Power - 2.25 kW (3 hp)
Capital Investment;
Equipment
Flocculation unit $ 3,000
Contingency and contractor's fee 390
Total equipment cost $ 3,390
Total Capital Investment $ 3.390
Annual Cost;
Amortization $ 505
Operation and Maintenance (O&M)
Operating personnel $ 900
Equipment repair and maintenance 150
Materials 180
Insurance 35
Total O&M costs 1,265
Electricity 300
Total Annual Cost $ 2,070
VIII-82
DRAFT
-------
DRAFT
Ferroalloy Mills Annually Processing More Than 5_,OOP Metric Tons
(5.500 Short Tons) Ore By_ Physical Methods
There are two mills in this subcategory, both of which are
located in the western U.S. The annual amount of ore milled
ranges from 7,200 to 1,800,000 metric tons (7,925 to 1,990,000
short tons). The daily wastewater flow ranges from 30 to
17,425 cubic meters (7,925 to 4,603,700 gallons).
A hypothetical mill was chosen to represent this subcategory.
The average annual milling capacity is 525,000 metric tons
(577,500 short tons), with a daily discharge of 4,920 cubic
meters (1,300,000 gallons).
Three alternative levels of technology are considered. The total
costs of implementing these levels are shown in Table VIII-18.
Wastewater Treatment/Control
Level A; Lime Precipitation
Level-A treatment consists of lime precipitation and settling.
The necessary settling ponds are currently available; therefore,
no cost estimates for these facilities have been made. The
addition of 1.36 kg of hydrated lime per 3.785 cubic meters
(3 lb/1000 gallons) of water would be necessary to raise the pH
sufficiently for precipitation of metals.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Lime precipitation system
Operating-Cost Assumptions for Level Aj_
Lime - 618 metric tons (682 short tons)/year
Operating personnel - 3 hr/day
Power - 37 kW (50 hp)
VI II-8:i
DRAFT
-------
DRAFT
TABLE VIII-18. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MILL (HYPOTHETICAL)-
FERROALLOY/PHYSICAL
Ferroalloy Mills Annually Processing More Than 5,000 Metric
8UBCATEOORY; Tons f5.512 Short Tons') Ore by Physical Methods
PLANT SIZE; 525 r OOP
PLANT APE; N/AYEARS
.METRIC TON*, ( 577.500 SHORT TONS) PER YEAR OF ore
PLANT LOCATION: N/A
milled
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS ($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
70.0
10.4
37.1
5.0
52.5
0.10
B
64.2
9.6
3.5
1.0
14.1
0.027
c
134.2
20.0
40.6
6.0
66.6
0.127
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
As
Cd
Cn
Mn
Mo
Pb
Zn
Fluoride
WASTE LOAD (kg/1000 metric torn of product [ore milled] t)
RAW
(UN-
TREATED)
1 x 106
9
3
0.2
0.3
50
1
7
15
30
AFTER TREATMENT TO LEVEL
A
20
3
0.4
0.05
0.05
1
0.1
0.15
10
B
0
0
0
o
0
0
0
0
0
0
c
6
0.9
0.1
0 02
0.01
0 2
0 ?
0.03
0.05
3
D
-
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWK B> 0.90/
tTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2 0
LEVEL A: LIME PRECIPITATION
LEVEL B: TOTAL RECYCLE (ZERO DISCHARGE)
LEVEL C: LEVEL A PLUS (LEVEL B WITHOUT ZERO DISCHARGE)
VIIT-84
DRAFT
-------
DRAFT
Capital Investment:
Lime precipitation unit $ 62,000
Contingency and contractor's fee 8,060
Total Capital Investment $ 70,060
Annual Cost;
Amortization $ 10,440
Operation and Maintenance (O&M)
Operating personnel $ 9,450
Equipment repair and maintenance 3,100
Materials 23,870
Insurance 7QQ
Total O&M costs $ 37,120
Electricity 5,020
Total Annual Cost $ 52,580
Level B_r_ Total Recycle (Zero Discharge)
Mills in this subcategory recycle approximately 60 percent
of their process water. The remaining 40 percent (1,968 cubic
meters, equivalent to 520,000 gallons, per day) is discharged
Level-B technology requires additional pumps and piping to
attain total recycle.
The capital and operating costs and assumptions for attaining
this level are shown below.
VLII-85
DRAFT
-------
DRAFT
Capital -Co at Components and Assumptions for Level B:
Piping - Flow @ 2 meters (6.6 feet)/second through pipe
measuring 12 cm (5 in.) x 1,750 meters
(5,740 feet)
Pumps - water pumps rated at 1,968 1 (361 gal)/min
Opera t ing-Coa t Assumptions for Level Z\_
Power - 7.5 kW (10 hp)
Capital Investment :
Piping $ 52,500
Pumps 4.300
Equipment subtotal 56,800
Contingency and contractor's fee 7,385
Total Capital Investment $ 64 . 185
Annual Cost;
Amortization $ 9,565
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 2,840
Insurance 640
Total O&M costs 3,480
Electricity
VIII-86
DRAFT
Total Annual Cost $ 14,045
-------
DRAFT
Level C; Level A plus Level JB
Level-C technology is applicable in areas where there
is excess water. The total cost of attaining this level
is the sum of the costs of attaining levels A and B.
These costs are shown in Table VIII-18.
Ferroalloy Mills Annually Processing More Than 5.000 Metric
Tons (5.500 Short Tons) Ore By Flotation
There are four mills in this subcategory, all of which are
located in the western U.S. The range of ore milled is 7,200
to 15,480,000 metric tons (7,925 to 17,030,000 short tons)
annually. The daily mill wastewater ranges from 30 to 94,600
cubic meters (7,925 to 25,000,000 gallons).
A hypothetical mill with an annual milling capacity of 5,600,000
metric tons (6,160,000 short tons) and with a daily wastewater
flow of 22,710 cubic meters (6,000,000 gallons) is representative
for this subcategory. Four levels of technology are considered.
The total costs of achieving these levels are shown in Table VIII-19.
Wastewater Treatment/Control
Level Aj^ Lime Precipitation and Discharge
The settling ponds necessary for adequate precipitation and settling
are considered to be already installed. The addition of 1.36
kg of pebbled lime per 3.785 liters (3.0 lb/1000 gallons) of water
is necessary for precipitation.
The capital and operating costs and assumptions for attaining
this level are shown below.
VIII-87
DRAFT
-------
DRAFT
TABLE VIII-19. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MILL (HYPOTHETICAD-
FERROALLOY/FLOTATION
Ferroalloy Mills Annually Processing More Than 5,000 Metric
8UBCATEOORY; Tons f5.512 Short Tons') Ore by Flotation
PLANT SIZE; 5 ,.600. OOP
PLANT AGE; N/AVEAR8
_METRIC TONS (6.160.000 SHORT TONS) PER YEAR OF ore nil led
PLANT LOCATION: N/A
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS ($)/METRIC TON OF PRODUCT*
COSTS ($10001 TO ATTAIN LEVEL
A
126.6
18.9
104.5
10.7
134.1
0.023
B
113.0
16.8
6.1
12.3
35.2
0.006
c
252.1
36.1
70.5
20.6
127.2
0.022
D
269.7
39.7
53.1
13.3
106.1
0.02
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
COD
Cyanide
As
Cd
Cu
Mn
Mo
Ph
Zn
Fluoride
WASTE LOAD (kg/1000 metric font of product [ore milled] T)
RAW
(UN-
TREATED)
1 x 106
8
200
0.95
2
0.4
42
100
52
4
50
23
AFTER TREATMENT TO LEVEL
A
40
6
100
0.11
0.8
0.1
0.1
2.0
_
0.2
<)..$
20
B
0
0
0
0
0
0
0
0
0
0
0
i)
c
10
1.5
13
0.01
0.2
0.03
0.02
0.4
0.3
O.OS
0.08
5
D
10
1.5
13
0.01
0.2
0.03
0.02
0.4
0.3
0.05
0.08
5
E
"ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED), MULTIPLY COSTS SHOWN BY O.B07
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: LIME PRECIPITATION AND DISCHARGE
LEVEL B: TOTAL RECYCLE
LEVEL C: LEVEL B PLUS FERRIC SULFATE ADDITION. FLOCCULATION. SETTLING. LIME NEUTRALIZATION.
SECONDARY SETTLING. AND AERATION
LEVEL D: LEVEL B PLUS AERATION. SETTLING. AND ION EXCHANGE
VIII-8B
DRAFT
-------
DRAFT
Capital-Cost Components and Assumptions for Level A;
Lime precipitation unit
Operating-Cost Assumptions for Level Aj_
Operating personnel - 3 hr/day x 360 days/year
Lime - pebbled, quantity of 2,857 metric tons (3,150
short tons)/year
Power - 75 kW (100 hp)
Capital Investment:
Equipment
Lime precipitation unit $ 112,000
Contingency and contractor's fee 14,560
Total equipment cost 126,560
Total Capital Investment $ 126.560
Annual Cost:
Amortization $ 18,860
Operation and Maintenance (O&M)
Operating personnel $ 9,450
Equipment repair and maintenance 5,600
Materials 88,200
Insurance 1,265
Total O&M costs 104,515
Electricity 10.700
Total Annual Cost $ 134,075
VIII-89
DRAFT
-------
DRAFT
Level B^ Total Recycle
To achieve total recycle, additional piping and pumps would
be necessary. The implementation of a total-recycle system
does not necessarily imply no discharge. The problem of
excess water due to rainfall still exists. The capital and
operating costs and assumptions for attaining this level are
shown below.
Capital-Cost Components and Assumptions for Level B;
Pumps - water pumps rated at 15,770 1 (4,163 gal)/min
Piping - Flow @ 2 meters (6.6 feet)/sec through pipe
measuring 42 cm (16.5 in.) x 1000 meters (3,280 feet)
Op era t ing-Co s t Assumptions for Level B;
Power - 89 kW (120 hp)
Capital Investment;
Equipment
Piping $ 21,000
Pumps 79.000
Equipment subtotal 100,000
Contingency and contractor's fee 13,000
Total Capital Investment $ 113,000
Annual Cost;
Amortization $ 16,840
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 5,000
Insurance 1.130
Total O&M costs 6,130
Electricity 12.25U
Total Annual Cost $ 35,220
VIII-90
DRAFT
-------
DRAFT
Level C; Level £ plus Ferric Sulfate Addition, Flocculation,
Settling. Lime Neutralization. Secondary Settling,
and Aeration
Level-C technology may be applied in areas of excess water.
It is assumed that 25 percent of the mill wastewater is bled
and discharged—a dally total of 5,677 cubic meters (1,500,000
gallons). The treatment recommended for mills In this sub-
category is the addition of 75 mg/1 of ferric sulfate and
5 mg/1 of flocculant to the wastewater stream. Acid is also
added to lower the pH to 4.5; however, no cost is shown for
this item, as the cost is negligible. The wastewater is then
contained for one day in a settling pond. Prior to discharge,
the wastewater is neutralized with lime (0.45 kg/3.785 cubic
meters, equivalent to 1 lb/1,000 gallons) and contained in
an aerated pond. Aeration is needed to lower COD and to
convert cyanide to cyanate. The capital and operating costs
and assumptions for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level C_:
2 Settling ponds - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 8,516 cubic meters
(2,250,000 gal)
Land - 1.06 hectares (2.6 acres)
Ferric sulfate addition - 2 mix tanks with capacity of
14.2 cubic meters (3,750
gallons)
1 metering pump
Flocculation system
Lime neutralization system
Aerator - 18 kW (24 hp)
Piping - Flow @ 2 meters (6.6 feet)/sec through pipe
measuring 21 cm (8.3 in.) x 200 meters
(656 feet)
VIII-91
DRAFT
-------
DRAFT
Operating-Coat Assumptions for Level C_:
Operating personnel - 6 hr/day
Materials - lime @ 236 metric tons (260 short tons)/year
ferric sulfate @ 149 metric tons (163 short
tons)/year
flocculant @ 9.9 metric tons (10.9 short
tons)/year
Power - 60 kW (81 hp)
Capital Investment;
Facilities
Lagoons
Contingency and contractor's fee
Total facility cost
Land
Equipment
Ferric sulfate system
Flocculation system
Lime neutralization unit
Piping
Aeration equipment
Equipment subtotal
Contingency and contractor's fee
Total equipment cost
Total Capital Investment
Annual Cost;
Amortization
Facility
Equipment
$ 22,000
2.860
12,550
14,900
55,000
9,000
8.000
99,450
12.930
$ 2,530
16.750
$ 24,860
1,860
112.380
$139.100
Total amortization
$ 19,280
VIII-92
DRAFT
-------
DRAFT
Operation and Maintenance (O&M)
Land $ 185
Operating personnel 18,900
Facility repair and maintenance 660
Equipment repair and maintenance 4,975
Materials 38,235
Taxes 45
Insurance 1,390
Total O&M costs $ 64,390
Electricity 8.270
Total Annual Cost $ 91,940
Level D_: Level JJ plus Aeration, Settling, and Ion Exchange
Level-D treatment is an alternative to level-C treatment.
Level-D technology may be applied in areas of excess water.
It is assumed that 10 percent of the mill wastewater is dis-
charged (a total of 2,271 cubic meters, equivalent to 600,000
gallons). This level of treatment includes an aeration pond
and an ion-exchange unit.
The excess wastewater is contained for one day in an aeration
pond to lower COD from 100 mg/1 to 20 mg/1 and to convert
cyanide to cyanate. The wastewater is then passed on to
an ion-exchange unit for further treatment before discharge.
The amount of ion-exchange resin actually needed would depend
upon the characteristics of the wastewater. For the purposes
of this report, it is assumed that 5.5 cubic meters (7.2 cubic
yards) of resin would be adequate.
The capital and operating costs and assumptions for attain-
ing this level are shown below.
Capital-Cost Components and Assumptions for Level D_:
Settling pond - dike height of 3 m (10 ft)
top width of 3 m (10 ft)
capacity of 3,400 cubic meters
(898,200 gallons)
VIII-93
DRAFT
-------
DRAFT
Land - 0.26 hectare (0.64 acre)
Aerator - 7.5 kW (10 hp)
Ion Exchanger - capacity of 5.5 cubic meters (7.1 cubic yards)
Piping - Flow @ 2 meters (6.6 feet)/sec through pipe
measuring 13 cm (5 in.) x 100 meters (328 feet)
Operating-Cost Assumptions for Level I):
Operating personnel - 10.8 hr/day
Resins - replacement every 3 years
Power - 7.5 kW (10 hp)
Capital Investment;
Facilities
Lagoon $ 6,200
Contingency and contractor's fee 805
Total facility cost $ 7,005
Land 455
Equipment
Aeration unit 3,400
Ion exchanger 125,000
Piping 3.200
Equipment subtotal 131,600
Contingency and contractor's fee 17,110
Total equipment cost 148,710
Total Capital Investment $156,170
Annual Cost;
Amortization
Facility $ 715
Equipment 22,165
Total amortization $ 22,880
VIII-94
DRAFT
-------
DRAFT
Operation and Maintenance (O&M)
Land
Operating personnel
Facility repair and maintenance
Equipment repair and maintenance
Materials
Taxes
Insurance
Total O&M costs
Electricity
Total Annual Cost
$ 46,985
1.020
$ 70,885
Ferroalloy Mills Practicing Ore Leaching
There is only one ferroalloy mill in this subcategory, and
it is located in the southeastern U.S. The annual ore milled
is 410,400 metric tons (451,500 short tons), with a daily
wastewater discharge of 5,300 cubic meters (1,400,000 gallons)
There are four levels of technologies considered. The total
costs of achieving these levels are shown in Table VIII-20.
Wastewater Treatment/Control
Level A: Lime Precipitation
Adequate sedimentation ponds are currently installed. The
addition of 0.9 kg of hydrated lime per 3.785 cubic meters
(2 lb/1000 gallons) of wastewater is required for precipi-
tation. The capital and operating costs and assumptions
for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level A_:
Lime precipitation system
Operating-Cost Assumptions for Level A:
Operating personnel - 2 hr/day
VII1-95
DRAFT
-------
DRAFT
TABLE VIH-20. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MILL (HYPOTHETICAL)
FERROALLOY/LEACHING
8UBCATEQORY: Ferroalloy Mills Practicing Ore Leacliing
PLANT SIZE: 410,400
PLANT AGE:N/A YEARS
METRIC TON8( 451,500 SHORT TONS) PER YEAR OF
PLANT LOCATION; N/A
milled
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATINQ AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTSISI/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
72.3
10.8
27.2
5.0
43.0
0.10
B
78.3
11.7
27.5
5.0
44.2
0.11
c
176.7
26.2
94.4
13.0
133.6
0.33
D
238.0
34.6
100.9
25.6
161.1
0.39
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Mn
Mo
Pb
V
Zn
Fluoride
WASTE LOAD (kg/1000 metric tons of product [ore milled! ')
RAW
(UN-
TREATED)
1 x 106
190
2.800
5.000
4.5
0.4
3
2
100
2.5
0.6
4.7
1.2
48
AFTER TREATMENT TO LEVEL
A
90
14
230
—
2.3
0.23
0.23
4.5
0.45
_
0.68
45
B
90
14
230
_
2.3
0.23
0.23
0.14
3.2
2.3
0.45
4.5
0.68
4f,
C
90
14
230
23
2.3
0.23
0.23
0.14
3.2
2.3
0.45
4..r,
0.6K
45
D
90
14
230
23
2.3
0.23
0.23
0.14
.V2
2.1
0.45
4.5
0.68
45
E
"ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BV 0.907
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: LIME PRECIPITATION
LEVEL B: LEVEL A PLUS SULFUR DIOXIDE INJECTION
LEVEL C: LEVEL B PLUS WASTEWATER SEGREGATION. CAUSTIC SODA ADDITION, SETTLING POND. AND
LEVEL D: LEVEL C PLUS AERATION AND SECONDARY SETTLING
VIII-96
DRAFT
-------
DRAFT
Lime - 441 metric tons (485 short tons)/year
Power - 37 kW (50 hp)
Capital Investment:
Equipment
Lime precipitation system $ 64,000
Contingency and contractor's fee 8,320
Total Capital Investment $ 72.320
Annual Cost;
Amortization $ 10,780
Operation and Maintenance (O&M)
Operating personnel $ 6,300
Equipment repair and maintenance 3,200
Materials 16,980
Insurance 725
Total O&M costs 27,205
Electricity 5.000
Total Annual Cost $ 42,985
Level Z: Level A plus Sulfur Dioxide Injection
Sulfur dioxide injection is required for chromium reduction.
The sulfur dioxide injection system requires a holding tank,
an injector, and sulfur dioxide. Total system costs are
shown in Table VIII-20. The incremental capital and operat-
ing costs and assumptions for attaining this level are shown
below.
Capital-Cost Components and Assumptions for Level B^:
Sulfur dioxide injection system -
1 holding tank with retention time of 5
minutes and capacity of 18,400 liters
(4,860 gallons)
ejector
VIII-97
DRAFT
-------
DRAFT
Operating-Cost Assumptions for Level JJ:
Sulfur dioxide - amount needed ia low, and its cost is
negligible
Capital Investment;
Equipment
Ejector $
Sulfur dioxide injection unit
Equipment subtotal
Contingency and contractor's fee
Total Capital Investment $ 5^,970
Annual Cost:
Amortization $ 890
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 220
Insurance 60
Total O&M costs $ 280
Total Annual Cost $ 1,170
Level £: Level IJ plus Wastewater Segregation. Caustic Soda
Addition, Settling Pond. and Air Stripper
Level-C technology suggests that 10 percent of the wastewater
(530 cubic meters, equivalent to 140,000 gallons) be segre-
gated from the rest of the mill wastewater. The reason is
that this water is contaminated with large amounts of ammonia.
To remove the ammonia, the wastewater must first be treated
with caustic soda to raise the pH to 10 or 11.5. Retention
for one day in a settling pond is necessary to remove sus-
pended solids. The wastewater is then sent to an air stripper,
which will remove 90 to 95 percent of the ammonia.
The costs for ammonia stripping have been provided by sur-
veyed operations. The capital and operating costs and assump-
tions for attaining this level arc shown below.
VIII-98
DRAFT
-------
DRAFT
Capital-Cost Components and Assumptions for Level C_:
Piping - Flow @ 1 meter (3.3 feet)/sec through pipe
measuring 9 cm (3.5 in.) x 1000 meters
(3,280 feet)
Pumps - slurry-type, rated at 370 1 (98 gal)/min
Caustic soda addition - plastic mix tank with
capacity of 11,4 cubic meters
(3 gallons) and cost of $3,000
(10-day supply of caustic)
Liquor feed pump @ $4,500, capacity of
42 liters (11 gallons)/hour
Instrumentation on mix tank @ $6,000 for
pH check/control
Ammonia stripper - packed column @ $33,000
fan @ $9,000
Settling pond - dike height of 2 meters (7 feet)
top width of 3 meters (10 feet)
capacity of 795 cubic meters (210,000
gallons)
Land - 0.11 hectare (0.27 acre)
Operating-Cost Assumptions for Level £:
Caustic soda - 175 metric tons (194 short tons)/year
@ $82/metric ton ($74.38 short ton)
Operating personnel - 1 man full-time
Power - 60 kW (80 hp)
VIII-99
DRAFT
-------
DRAFT
Capital Investment:
Facilities
Lagoon $ 2,000
Contingency and contractor's fee 260
Total facility cost " $ 2,260
Land 190
Equipment
Caustic soda addition unit 13,500
Ammonia stripper 42,000
Piping 25,000
Pumps 4.400
Equipment subtotal 84,900
Contingency and contractor's fee 11,040
Total equipment cost 95.940
Total Capital Investment $ 98.390
Annual Cost;
Amortization
Facility $ 230
Equipment 14,300
Total amortization $ 14,530
Operation and Maintenance (O&M)
Land 20
Operating personnel 47,250
Facility repair and maintenance 60
Equipment repair and maintenance 4,245
Materials 14,350
Taxes 5
Insurance 985
Total O&M costs 66,915
Electricity 8»°00
Total Annual Cost $ 89,445
VI11-100
DRAFT
-------
DRAFT
Level I): Level £ plus Aeration and Secondary Settling
Further treatment would include the merging of the waste
streams into an aerated pond. The purpose of aeration is
to lower COD. A one-day retention is recommended before
discharge. The capital and operating costs and assumptions
for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level £:
Settling pond - dike height of 3 meters (10 feet)
top width of 3 meters (10 feet)
capacity of 7,950 cubic meters
(2,100,000 gallons)
Land - 0.5 hectare (1.2 acres)
Aerator - 94 kW (126 hp)
Capital Investment;
Facilities
Lagoon
Contingency and contractor's fee
Total facility cost
$ 11,500
1.495
Land
Equipment
Aerator
Contingency and contractor's fee
Total equipment cost
Total Capital Investment
42,000
5.460
$ 12,995
875
47.460
$ 61.330
VIII-101
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Facility $ 1,325
Equipment 7,075
Total amortization $ 8,400
Operation and Maintenance (O&M)
Land 90
Facility repair and maintenance 345
Equipment repair and maintenance 5,460
Taxes 20
Insurance 615
Total O&M costs 6,530
Electricity 12,600
Total Annual Cost $ 27,530
VIII-102
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR MERCURY-ORE CATEGORY
Mercury-Ore Mines
The exact number of operating mercury mines is difficult to
determine at present. One open-pit mine is currently con-
sidered active; however, it does not have a discharge and is
closed seasonally.
Currently, existing market conditions have resulted in almost
no activity from underground mercury mines. It is expected
that, with a return to more favorable market conditions, some
underground mines will again become active.
In anticipation of a rise in the market price of mercury, a
hypothetical mine was chosen to represent this subcategory.
The representative mine has an annual ore production of
27,210 metric tons (30,000 short tons) with a daily wastewater
flow of 378.5 cubic meters (100,000 gallons).
One level of technology is considered. The total costs of
achieving this level are shown in Table VIII-21.
Wastewater Treatment/Control
Level A^ Lime Precipitation. Sulfide Precipitation, Settling.
and Discharge
The addition of 1.36 kg of hydrated lime per 3.785 cubic meters
(3.0 lb/1000 gallons) and 1 mg/1 of sodium sulfide to the
wastewater is recommended for adequate precipitation of metals.
Since this operation is small, a simplified method of liming
Is used. A 15-day supply of hydrated lime (2,040 kg, equiva-
lent to 4,488 Ib) is stored as a slurry (0.9 kg/3.785 1,
equivalent to 2 Ib/gallon) in a mixing tank. A portion
of the slurry is drawn off and mixed with the mine water in
another mixing tank for 15 minutes then is pumped into a
settling pond. This process is done on a continual basis.
VIIT-103
DRAFT
-------
DRAFT
TABLE VIII-21. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL)-
MERCURY
SUBCATEQQRY: Mercury-Ore Mines
PLANT SIZE: 27,210
PLANT AGE:N/A YEARS
.METRIC TONS (50,000
PLANT LOCATION: N/A
SHORT TONS) PER YEAR OF ore mined
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
29.6
4.2
9.7
1.1
15.0
0.55
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Hg
Ni
CONCENTRATION (mg/£) (ppm)
RAW
(UN-
TREATED)
25
0.001
0.2
AFTER TREATMENT TO LEVEL
A
20
0.0005
0.1
B
C
D
E
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0407
LEVEL A: LIME PRECIPITATION. SULFIDE PRECIPITATION. SETTLING. AND DISCHARGE
VIII-104
DRAFT
-------
DRAFT
The capital and operating costs and assumptions for attain-
ing this level are shown below.
Capital-Cost Components and Assumptions for Level A;
2 Ponds - dike height of 2 m (7 ft)
top width of 3 m (10 ft)
capacity of 570 cubic meters (150,600 gallons)
Land - 0.2 hectare (0.5 acre)
Lime precipitation unit -
slurry storage tank with a capacity of 8,580 liters
(2,265 gallons) and containing a 15-day supply
of slurry
mix tank with retention time of 15 minutes and
capacity of 3,975 liters (1,050 gallons),
based on flow of 265 liters (70 gallons)
per minute
pump with capacity of 265 liters (70 gallons) per
minute
Sulfide precipitation unit - drum with capacity of 208
cubic meters (55 gal)
Piping - Flow @ 2 m (6.6 feet)/sec through pipe measuring
5 cm (2 in.) x 1,100 meters (3,608 feet)
Operating-Cost Assumptions for Level A;
Lime - 47.5 metric tons (53 short tons)/year
Sodium sulfide - 132 kg (291 lb)/year
Operating personnel - 2 hr/day
Power - 8.2 kW (11 hp)
VIII-105
DRAFT
-------
DRAFT
Capital Investment;
Facilities
Lagoon $ 3,400
Contingency and contractor's fee 440
Total facility cost $ 3,840
Land 350
Equipment
Lime precipitation unit 6,950
Sulflde precipitation unit 100
Piping 15.400
Equipment subtotal 22,450
Contingency and contractor's fee 2,920
Total equipment cost 25.370
Total Capital Investment $ 29.560
Annual Cost;
Amortization
Facility $ 390
Equipment 3.780
Total amortization $ 4,170
Operation and Maintenance (O&M)
Land 35
Operating personnel 6,300
Facility repair and maintenance 100
Equipment repair and maintenance 1,120
Materials 1,885
Taxes 10
Insurance 295
Total O&M costs 9,745
Electricity Lilou
Total Annual Cost $ 15,015
VIII-106
DRAFT
-------
DRAFT
Mercury Mills Employing Flotation Process
There are no mills currently operating in this subcategory.
A mill utilizing a flotation process is due to open in 1975.
This mill was chosen to be representative for this subcate-
gory. It is expected to mill 159,000 metric tons (175,000
short tons) a year. Discharge of wastewater is expected to
be 7,570 cubic meters (2,000,000 gallons) dally.
The recommended level of treatment is zero discharge of waste-
water. Two alternatives for achieving zero discharge are
considered. They are total recycle, or Impoundment and
evaporation. The costs of implementing these alternatives
are shown in Table VIII-22.
Wastewater Treatment/Control
Level A_: Total Recycle (Zero Discharge)
The facilities required to achieve total recycle include a
rectangular pond of 40 hectares (100 acres) whose
length is equal to twice Its width. The pond would also
require one transverse dike to provide two separate ponds,
each having an area of 20 hectares (50 acres). The first
pond would be used for sedimentation of suspended solids.
The second pond would be used as a polishing pond. Water
in the polishing pond would be recycled back to the mill.
Diversion ditching along one length and one width is recom-
mended to avoid stress in the system due to seasonal runoff.
Additional equipment includes a tailing-disposal system and
decant pumps and pipes. The capital and operating costs and
assumptions for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level A_:
Pond - dike height of 2 m (7 ft)
top width of 3 m (10 ft)
capacity of 750,000 cubic meters
Land - 40 hectares (100 acres)
VIII-107
DRAFT
-------
DRAFT
TABLE VI11-22. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 9202
8UBCATEOQBY; Mercury Mills Employing Flotation Process
PLANT SIZE: 159.000
PLANT AGE: YEARS
(under construction in 1975)
METRIC TONS (175.000 SHORT TONS) PER YEAR OF ore milled
PLANT LOCATION; Nevada
«. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS (S)/METRIC TON OF PRODUCT*
COSTS (S1000) TO ATTAIN LEVEL
A
565.3
64.4
62.7
6.5
133.6
0.84
B
736.0
71.5
66.4
2.5
140.4
0.88
c
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
He
Ni
WASTE LOAD (kg/1000 metric tom of product (or* miltodl*)
RAW
(UN-
TREATED)
1.400
0.1
0.6
AFTER TREATMENT TO LEVEL
A
0
0
0
B
0
0
0
c
D
E
"ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.907
*TO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: TOTAL RECYCLE (ZERO DISCHARGE)
LEVEL B: IMPOUNDMENT AND EVAPORATION (ZERO DISCHARGE)
VIII-108
DRAFT
-------
DRAFT
Transverse dike - height of 461 meters (1,512 feet)
Diversion ditching - total of 1,405 meters (4,608 feet)
Distribution system - around one pond - pipe measuring
34 cm (13.4 in.) x 1,844 m
(6,048 ft)
Piping - mill to pond - flow @ 1 m (3.3 ft)/sec through
pipe measuring 34 cm (13.4 in.) x 1000 meters
(3,280 feet)
pond to mill - flow @ 2 m (6.6 feet)/sec through
pipe measuring 25 cm (9.8 in.) x 1000 meters
(3,280 feet)
Pumps - mill to pond - slurry type, capacity of 5,260 1
(1,389 gal)/minute
pond to mill - water type, capacity of 5,260 1
(1,389 gal)/minute
Operating-Cost Assumptions for Level A_:
Power - 48 kW (65 hp)
Capital Investment;
Facilities
Diversion ditching $ 2,320
Lagoon 149,760
Transverse dike 24.900
Facility subtotal 176,980
Contingency and contractor's fee 23,010
Total facility cost $ 199,990
Land 70,000
Equipment
Distribution system 119,860
Piping 116,000
Pumps 25.500
Equipment subtotal 261,360
Contlngenclty and contractor's fee 33.975
Total equipment cost 295,335
Total Capital Investment $ 565.325
VIII-109
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Facility $ 20,370
Equipment 44,015
Total amortization $ 64,385
Operation and Maintenance (O&M)
Land 7,000
Facility repair and maintenance 5,310
Equipment repair and maintenance 7,075
Distribution system maintenance 35,960
Taxes 1,750
Insurance 5.650
Total O&M costs 62,745
Electricity 6.500
Total Annual Cost $133,630
Level B; Impoundment and Evaporation (Zero Discharge)
The facilities required for level-B treatment are essentially
the same as those required for level-A treatment. However, a
larger pond area is required. An 80-hectare (200-acre) rectang-
ular pond with three transverse dikes to provide four separate
ponds of 20 hectares (50 acres) each is required for impound-
ment and evaporation.
The equipment required includes a tailing-disposal system (the
same as that for level A), pumps, and pipes. The capital and
operating costs and assumptions for attaining this level are
shown below.
Capital-Cost Components and Assumptions for Level B:
Pond - dike height of 2 meters (7 ft)
top width of 3 meters (10 ft)
capacity of 1,500,000 cubic meters (396,260,000 gal)
Land - 80 hectares (200 acres)
VIII-110
DRAFT
-------
DRAFT
Transverse dikes - 3, each 650 meters (2,132 feet) in length
Diversion ditching - around one length and one width, 1,970
meters (6,462 feet) in length
Distribution system - piping around one 20-hectare (50-acre)
pond; diameter of 34 cm (13.4 in.)
and length of 1,844 m (6,048 ft)
Piping - mill to pond flow @ 1 m (3.3 ft)/sec through pipe
measuring 34 cm (13.4 in.) x 1000 meters
(3,280 feet)
Pumps - mill to pond slurry type, capacity of 5,260 1
(1,390,000 gal)/min
Operating-Cost Assumptions for Level Bj_
Power - 19 kW (25 hp)
Capital Investment;
Facilities
Diversion ditching $ 3,250
Lagoon 211,200
Transverse dike 105.300
Facility subtotal 319,750
Contingency and contractor's fee 41,570
Total facility cost $ 361,320
Land 140,000
Equipment
Distribution system 126,750
Piping 65,000
Pumps 16.000
Equipment subtotal 207,750
Contingency and contractor's fee 27,010
Total equipment cost 234,760
Total Capital Investment $ 736.080
VIII-111
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Facility
Equipment
Total amortization
Operation and Maintenance (O&M)
Land
Facility repair and maintenace
Equipment repair and maintenance
Distribution system maintenance
Taxes
Insurance
Electricity
Total O&M costs
Total Annual Cost
$ 36,800
34,745
14,000
9,590
4,050
38,025
3,500
7,275
$ 71,545
66,440
2.500
$ 140,485
Mercury Mills Employing Gravity Separation
There is only one mill in this subcategory. The discharge of
wastewater is 1,665 cubic meters (436,000 gallons) a day during
wet seasons. The mill process water is recycled. Annual ore
milled is 27,000 metric tons (30,000 short tons).
One level of technology is considered. The total costs of
Implementing this level are shown in Table VIII-23.
Wastewater Treatment/Control
Level Ajr_ Diversion Ditching (Zero Discharge)
Diversion ditching along one length and one width of the present
tailing pond is recommended to avoid stress in the system due to
seasonal runoff. The capital costs and assumptions for attain-
ing this level are shown below.
VIII-112
DRAFT
-------
DRAFT
TABLE VIII-23. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 9201
SUBCATEOORY: Mercury Mills Employing Gravity Separation
PLANT SIZE; 27,000 METRIC TONS < 30. OOP SHORT TONS) PER YEAR OF Ore milled
PLANT APE; 4 YEARS PLANT LOCATION; California
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDINQ ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
0.4
0.045
0.010
0 055
0.002
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
HP
Mi
WASTE LOAD (kg/1000 metric tons of product (ore milled! T)
RAW
(UN-
TREATED)
10
0.013
n.o?
AFTER TREATMENT TO LEVEL
A
0
0
0
B
C
D
E
*ORi MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.907
tTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: DIVERSION DITCHING (ZERO DISCHARGE)
VIII-113
DRAFT
-------
DRAFT
Capital-Cost Components and Assumptions for Level A:
Diversion ditching - 225 meters (738 feet) @ $1.65/meter
($0.50/foot)
Capital Investment;
Facilities
Diversion ditching $ 370
Facility subtotal 370
Contingency and contractor's fee 50
Total facility cost $
Total Capital Investment §
Annual Cost;
§ 45
Amortization *
Operation and Maintenance (O&M)
Facility repair and maintenance $ 10
Total O&M costs —iP-
Total Annual Cost $ 55
VIII-1U
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR URANIUM-ORE CATEGORY
Uranium Mines
There are 175 known uranium mines in the U.S. The annual
amount of ore mined ranges from 1,800 to 504,000 metric
tons (1,980 to 554,500 short tons). The daily wastewater
flow ranges from 0 to 5,000 cubic meters (0 to 1,321,000
gallons).
A hypothetical mine with an annual ore production of 280,000
metric tons (308,000 short tons) and with a daily water flow
rate of 1,900 cubic meters (500,000 gallons) was chosen as
representative.
Several levels of technology have been considered. The
total costs of implementing these levels are shown in Table
VIII-24.
Wastewater Treatment/Control
Level A_: Flocculation
The necessary settling and polishing ponds are already
installed at the typical uranium mining operation. The
addition of 5 mg/1 of flocculant is required for settling
of suspended solids. The capital and operating costs and
assumptions for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level Aj
Flocculation -
1 mix tank with capacity of 1,900 liters
(500 gallons)
2 mix tanks with capacity of 9,500 liters
(2,500 gallons)
2 positive-displacement pumps
VIII-115
DRAFT
-------
DRAFT
TABLE VIII-24. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL) -
URANIUM (Sheet 1 of 2)
8UBCATEQORY: Uranium Mines
PLANT SIZE: 280»
PLANT AOEiN/A YEARS
_METHIC TONS I 308.000 SHORT TONS) PER YEAR OF ore mined
PLANT LOCATION: N/A
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER!
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTSISI/METRIC TON OF PRODUCT*
COSTS IS1000I TO ATTAIN LEVEL
A
16.8
2.5
11.4
1.3
25.2
0.09
B
86.8
12.9
15.2
1.5
29.6
0.11
c
228.1
33.9
(45.2)**
1.5
f9.81**
(0.035)*
D
240.5
35.8
(19.9)**
1.5
17.4
* 0.062
E
282. b
42.1
(2.0)**
3.5
41 f.
0.156
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
COD
TOC
As
Cd
Cr
Cu
Mo
Ni
Pb
V
Zn
Ra 226
Total Th
Total U
CONCENTRATION (mgtt) (ppm)
RAW
(UN-
TREATED)
530
750
45
2
0.05
2
0.5
1.2
0.5
0.2
2.1
0.4
3,200*
1
25
AFTER TREATMENT TO LEVEL
A
50
200
45
2
0 . 05
2
0.5
1.2
0.5
0.2
2.1
0.4
200 '
1
25
B
20
100
45
2
0.05
2
0.5
1.2
0.5
0.2
2.1
0.4
30*
1
25
c
20
100
45
2
0.05
2
0.5
1.2
0.5
0.2
2.1
0.4
30'
1
1
0
20
100
45
0.5
0.05
0.5
0.05
1.2
0.1
0.1
2.1
0.1
I1
1
1
E
20
100
45
0.5
0.05
0.5
0.05
1.2
0.1
0.1
2.1
0.1
l'
I. 1
1
*ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0*07
"TREATMENT RESULTS IN NET RETURN ON INVESTMENT. (REFER TO TEXTI
'VALUE IN PICOCURIES/&
LEVEL A- FLOCCULATION
LEVEL B. LEVEL A PLUS CLARIFICATION
LEVEL C LEVEL B PLUS ION EXCHANGE
LEVEL D: LEVEL C PLUS BARIUM CHLORIDE COPRECIPITATION
LEVEL E: LEVEL D PLUS LIME PRECIPITATION
VIII-116
DRAFT
-------
DRAFT
TABLE VIII-24. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL) -
URANIUM (Sheet 2 of 2)
suacATEQORv: Uranium Mines
PLANTS.ZE: 280.000
PLANT AOE!N/A YEARS
METRIC TONS I308.000
PLANT LOCATION! N/A
SHORT TONS) PER YEAR OF
"Uned
I. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTSISI/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
E
282.6
42.1
(2.0)**
3.5
43. (.
0.156
F
294.0
43.8
2.2
6.5
52.5
0.19
G
435.3
64.8
8.9
(j.5
80.2
0.280
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
COD
TOC
As
Cd
Cr
Cu
Mo
Ni
Pb
V
Zn
Ra 226
Total Th
Total U
CONCENTRATION (mglU (ppm)
RAW
(UN-
TREATED)
530
750
45
2
0.05
2
0.5
1.2
0.5
0.2
2.1
0.4
3,200T
1
25
AFTER TREATMENT TO LEVEL
E
20
100
45
0.5
0.05
0.5
0.05
1.2
0.1
0.1
2.1
0.1
1'
0.1
1
F
20
50
20
0.1
0.004
0.05
0.01
1.2
0.1
0.05
2.1
0.1
lf
0.1
1
G
20
50
20
0.1
0.004
0.05
0.01
0.5
0.1
0.05
1
0.1
1»
0.1
I
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0307
•TREATMENT RESULTS IN NET RETURN ON INVESTMENT (REFER TO TEXT)
'VALUE IN PICOCURIES/&
LEVEL F: LI-VtL h PLUS SutFiOe FRECiriTATiON AND AERATION
LEVEL G. LEVEL F PLUS ION EXCHANGE
VIII-117
DRAFT
-------
DRAFT
Operating-Cost Assumptions for Level AJ
Flocculant - 6,621 kg (7,300 lb)/year
Operating personnel - 1 hr/day
Power - 9.7 kW (13 hp)
Capital Investment;
Equipment
Flocculation system $ 14,900
Contingency and contractor's fee 1,940
Total Capital Investment $ 16.840
Annual Cost;
Amortization $ 2,510
Operation and Maintenance (O&M)
Operating personnel $ 3,150
Equipment repair and maintenance 745
Materials 7,300
Insurance 170
Total O&M costs 11,365
Electricity 11.300
Total Annual Cost $ 25,175
Level £: Level A plus Clarification
Level-B technology includes level-A technology plus clarifi-
cation. A one-hour retention time in the clarification unit
is required. The clarifier required has a capacity of 80
cubic meters (20,850 gallons). The capital and operating
costs and assumptions for attaining this level are shown
below.
VIII-118
DRAFT
-------
DRAFT
Capital-Coat Components and Assumptions for Level JB:
Clarifier - capacity of 80 cubic meters (20,850 gallons)
Oper at ing-Co at Assumptions for Level B^:
Power - 1.5 kW (2 hp)
Capital Investment;
Equipment
Clarifier $ 62,000
Contingency and contractor's fee 8,060
Total Capital Investment $ 70.060
Annual Cost;
Amortization $ 10,440
Operation and Maintenance (O&M)
Equipment repair and maintenance $ 3,100
Insurance 700
Total O&M costs 3,800
Electricity 200
Total Annual Cost $ 14,440
Level C_: Level £ plus Ion Exchange
The amount of resin needed is dependent upon the character-
istics of the wastewater. For this report, the amount of
resin chosen was based on actual operations.
A recovery of 13.6 kg (30 Ib) of U30J8 is made daily in the
ion-exchange unit.
The capital and operating costs and assumptions for attain-
ing this level are shown below.
VIII-119
DRAFT
-------
DRAFT
Capital-Cost Components and Assumptions for Level C_:
Ion exchanger - -.capacity of 5.6 cubic meters (7.3 cubic yards)
Operating-Coat Assumptions for Level C_:
Operating personnel - 3.5 hr/day
Materials - change resins every 3 years
Product recovery - 13.6 kg (30 lb)/day of U308. @ $17.60/kg
($7.99/lb)
Capital Investment;
Equipment
Ion exchanger $ 125,000
Contingency and contractor's fee 16.250
Total Capital Investment $ 141.250
Annual Cost;
Amortization $ 20,975
Operation and Maintenance (O&M)
Operating personnel $ 11,025
Equipment repair and maintenance 6,250
Materials 4,670
Insurance 1,410
Total O&M costs 23.355
Total Annual Cost 44,330
Less Product Recovery 83,775
Net Annual Recovery $ 39,445
Level D^: Level £ plus Barium Chloride Coprecipitation
Level-D technology, compared with that of level C, requires
the addition of flocculant and barium chloride for the
VIII-120
DRAFT
-------
DRAFT
precipitation of radium. The costs for this system are based
on actual operations. The costs for barium chloride copreci-
pitation are shown below. Total costs for level D are shown
in Table VIII-24.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level D^:
Barium chloride coprecipitation system
Operating-Cost Assumptions for Level D^:
Flocculant -6.4 metric tons (7 short tons)/year
Barium chloride - 5.4 metric tons (6 short tons)/year
@ $805/metric ton ($730/short ton)
Operating personnel - 2 hr/day
Capital Investment;
Equipment
Barium chloride coprecipitation system $ 11,000
Contingency and contractor's fee 1.430
Total Capital Investment $ 12,430
Annual Cost;
Amortization $ 1,850
Operation and Maintenance (O&M)
Operating personnel $ 6,300
Equipment repair and maintenance 550
Materials 18,345
Insurance 125
Total O&M costs $ 25.320
Total Annual Cost $ 27,170
VIII-121
DRAFT
-------
DRAFT
Level JJ: Level D^ plus Lime Precipitation
The required settling ponds are currently available for
precipitation. The addition of 0.9 kg of hydra ted lime per
3.785 cubic meters (2 lb/1000 gal) of wastewater is consid-
ered sufficient for precipitation of heavy metals. The
total costs for implementing level-E technology are shown
in Table VIII-24.
The incremental capital and operating costs and assumptions
for the lime precipitation necessary to attain this level
are shown below.
Capital-Cost Components and Assumptions for Level IS:
Lime precipitation system
Operating-Cost Assumptions for Level E_:
Lime - 160 metric tons (175 short tons)/year
Operating personnel - 3 hr/day
Power - 14.9 kW (20 hp)
Capital Investment;
Equipment
Lime precipitation system $ 37,250
Contingency and contractor's fee 4,845
Total Capital Investment $ 42.095
Annual Cost;
Amortization $ 6,275
Operation and Maintenance (O&M)
Operating personnel $ 9,450
Equipment repair and maintenance 1,865
Materials 6,125
Insurance 420
Total O&M costs $ 17.f«oO
VIII-122
DRAFT
-------
DRAFT
Electricity $ 2.000
Total Annual Cost $ 26,135
Level F_: Level E_ plus Sulfide Precipitation and Aeration
To achieve level F, the addition of 3 mg/1 of sodium sulfide
and aeration to lower COD levels would be necessary. The
total costs for implementing level-F technology are shown
in Table VIII-24.
The incremental capital and operating costs and assumptions
for attaining this level via sulfide precipitation and aera-
tion are shown below.
Capital-Cost Components and Assumptions for Level F_:
Sulfide precipitation system
Aeration - 30 kg (66 Ib) of oxygen/hour
Operating-Cost Assumptions for Level T_:
Sodium sulfide - 1,985 kg (4,375 lb)/year
Power - 22.4 kW (30 hp)
Operating personnel - 1 hr/day
Capital Investment:
Equipment
Sulfide precipitation unit $ 100
Aeration equipment 10.000
Equipment subtotal 10,100
Contingency and contractor's fee 1.315
Total Capital Investment $ 11.415
VIII-123
DRAFT
-------
DRAFT
Annual Cost;
Amortization $ 1,700
Operation and Maintenance (O&M)
Operating personnel $ 3,150
Equipment repair and maintenance 505
Materials 440
Insurance 115
Total O&M costs 4,210
Electricity 3.000
Total Annual Cost $ 8,910
Level G_: Level F_ plus Ion Exchange
For further removal and recovery of molybdenum and vanadium,
another ion-exchange unit would be necessary. Approximately
the same amount of Mo and V are recovered as uranium. The
incremental costs for this system are the same as for level
C. However, the values of the recovered Mo and V differ.
The incremental capital and operating costs and assumptions
for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level G_:
Ion exchanger - capacity of 5.6 cubic meters (7.3 cubic
yards)
Operating-Cost Assumptions for Level G_:
Operating personnel - 3.5 hr/day
Material - change resins every 3 years
Product recovery - 13.6 kg (30 lb)/day of Mo and V
@ $3.50/kg ($1.59/lb)
VIII-124
DRAFT
-------
DRAFT
Capital Investment:
Equipment
Ion exchanger $ 125,000
Contingency and contractor's fee 16.250
Total Capital Investment $ 141.250
Annual Cost;
Amortization $ 20,975
Operation and Maintenance (O&M)
Operating personnel $ 11,025
Equipment repair and maintenance 6,250
Materials 4,670
Insurance 1,410
Total O&M costs $ 23.355
Total annual cost 44,330
Less product recovery 16.660
Total Annual Cost $ 27,670
Uranium Mills Using Acid or Combined Acid/Alkaline Leaching
There are 16 mills in this subcategory. The annual amount
of ore milled ranges from 161,280 to 2,295,000 metric tons
(177,400 to 2,524,500 short tons). The daily wastewater flow
ranges from 865 to 10,945 cubic meters (228,500 to 2,900,000
gallons). There are only two operations in this subcategory
that are discharging. All others are at zero discharge.
An existing mill with an annual milling capacity of 648,000
metric tons (714,000 short tons) and a daily discharge of
3,260 cubic meters (861,300 gallons) was chosen as represen-
tative for this subcategory.
One level of technology is considered. The costs are shown
in Table VIII-25.
VIII-125
DRAFT
-------
DRAFT
TABLE VI11-25. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 9405
BUBCATEQORY! Uranium Mills Using Acid or Combined Arid/Alkaline LeachinP
PLANT SIZE; 648. OOQ METRIC TONS ( 714 f OOP SHORT TONS) PER YEAH OF ore milled
PLANT APE; 60 YEARS PLANT LOCATION; Colorado
•. COSTS OP TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTSISt/METRIC TON OF PRODUCT*
COSTS ($10001 TO ATTAIN LEVEL
A
1.101.8
128.8
68.7
10.2
207.7
0.32
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
COD
TOC
Al
Cu
Mn
Pb
Cr
Mo
V
Ra 226
WASTE LOAD (kg/1000 metric tons of product lore milled] f)
RAW **
(UN-
TREATED)
998,000
30
30
1,000
2
100
2
8
13
120
5001
AFTER TREATMENT TO LEVEL
A
0
0
0
0
0
0
0
0
0
0
* 0
B
C
D
E
*ORB MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN HY O.*»07
tTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN liY 7 0
••HYPOTHETICAL • BASED ON INDUSTRY AVERAGE
ttVALUB IN MICROCURIE8/1000 TONS
LEVEL A: IMPOUNDMENT AND EVAPORATION (ZERO DISCHARGE)
VIII-126
DRAFT
-------
DRAFT
Wastewater Treatment/Control
Level A; Impoundment and Evaporation (Zero Discharge)
To stop the present discharge of wastewater, a 93-hectare
(230-acre) evaporation pond would be necessary. Since land
is not available in the nearby area, a site which is approxi-
mately 16 kilometers (10 miles) from the mill was chosen as
the site of the pond. Because of the great distance that the
wastewater must be pumped, five pumping stages are necessary.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Pond - dike height of 2 m (7 ft)
top width of 3 m (10 ft)
capacity of 1,600,000 cubic meters (422,677,000 gallons)
Land - 93 hectares (230 acres)
Pumps - 5-stage water pumps with capacity of 2,264 liters
(598,100 gallons)/minute
Piping - Flow @ 2 meters (6.6 ft)/sec through pipe with
diameter of 16 cm (6.3 in.)
Operating-Cost Assumptions for Level A;
Power - 75 kW (100 hp)
Capital Investment;
Facilities
Lagoon $ 210,000
Contingency and contractor's fee 27,300
Total facility cost $ 237,300
Land 162,750
VIII-127
DRAFT
-------
DRAFT
Equipment
Piping ; ^
Pumps _j!!2jL.QP.P_
Equipment subtotal 6*21,000
Contingency and contractor's fee 80.730
Total equipment cost ~ $ 701,730
Total Capital Investment $1,101,780
Annual Cost;
Amortization
Facility $ 24,170
Equipment 104.580
Total amortization $ 128,750
Operation and Maintenance (O&M)
Land 16,275
Facility repair and maintenance 6,300
Equipment repair and maintenance 31,050
Taxes 4,070
Insurance 11,020
Total O&M costs 68,715
Electricity 10.210
Total Annual Cost $ 207,675
Uranium Mills Using Alkaline Leaching
There are three mills in this subcategury. The junuul amount
of ore milled ranges from 143,640 to 1,11>0,0()0 nu'Lru (mis
(158,000 to 1,265,000 short tons). The daJJy vMstuwutur flow
ranges from 865 to 6,340 cubic meters (228,500 LCI l,t>75,000
gallons).
VIII-128
DRAFT
-------
Of the three mills in this subcategory, only one is currently
discharging. All others are at zero discharge. The mill
currently discharging mills 432,000 metric tons (475,500
short tons) of ore annually and discharges 605 cubic meters
(160,000 gallons) of wastewater daily.
One level of technology is considered. The costs are shown in
Table VIII-26.
Wastewater Tr eatment/Contro1
Level Aj_ Impoundment and Evaporation (Zero Discharge)
To control the present discharge of wastewater, a 30-hectare
(74-acre) evaporation pond would be necessary. Land is not
readily available at the milling site; therefore, the waste-
water must be pumped 8 kilometers (5 miles). Two pumping stages
are necessary. The capital and operating costs and assumptions
for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level A;
Pond - dike height of 2 m (7 ft)
top width of 3 m (10 ft)
capacity of 500,000 cubic meters (132,087,000 gallons)
Land - 30 hectares (74 acres)
Piping - Flow @ 2 meters (6.6 ft)/sec through pipe with
diameter of 7 cm (2.75 in.)
Pumps - 2-stage water pumps with capacity of 420 liters
(111 gal)/minute
Operating-Cost Assumptions for Level A:
Power - 3.7 kW (5 hp)
VIII-129
DRAFT
-------
DRAFT
TABLE VII1-26. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 9403
Uranium Mills Using Alkaline Leaching
PLANT SIMi A32rnQQ
PLANT APE; 18 YEARS
METRIC TONS ( 475.500 SHORT TONS) PER YEAH OF QTR mill
PLANT LOCATION; Utah
e. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS/METRIC TON Or rHUUUCT
COSTS (11000) TO ATTAIN LEVEL
A
421.7
46.8
24.0
0.5
71.3
0.165
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
roo
Al
As
Cu
Mn
DK
WASTE LOAD (kg/1000 metric tons of product (ore milled] )
RAW
(UN-
TREATED)
998,000
100
67
5
4
140
3
140
AFTER TREATMENT TO LEVEL
A
0
0
0
0
0
0
0
0
B
=1=^=:
c
D
E
S^S^^BS
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHCKT..M BY 0.
tTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN r, - .1.0
LEVEL A: IMPOUNDMENT AND EVAPORATION (ZERO DISCHARGE)
VIII-130
DRAFT
-------
Capital Investment;
Facilities
Lagoon $ 154,000
Contingency and contractor's fee 20.020
Total facility cost $ 174,020
Land 52,500
Equipment
Piping 168,000
Pumps 4.700
Equipment subtotal 172,700
Contingency and contractor's fee 22.450
Total equipment cost 195,150
Total Capital Investment $ 421.670
Annual Cost;
Amortization
Facility $ 17,725
Equipment 29.085
Total amortization $ 46,810
Operation and Maintenance (O&M)
Land 5,250
Facility repair and maintenance 4,620
Equipment repair and maintenance 8,635
Taxes 1,315
Insurance 4.215
Total O&M costs 24,035
Electricity 510
Total Annual Cost $ 71,355
VIII-131
DRAFT
-------
DRAFT
WASTEWATER-TREATMENT COSTS FOR METAL ORES. NOT ELSEWHERE
CLASSIFIED
Antimony Mines
There is only one mine in this subcategory. To date, it
has no discharge; however, this mine was started in 1970, and
a discharge may occur as it becomes more extensively developed.
A hypothetical discharge of 378.5 cubic meters (100,000 gallons)
of wastewater daily is assumed for this operation. The
annual ore production is 10,300 metric tons (11,365 short
tons).
Two levels of technology are considered. The total cost
of each level is shown in Table VIII-27.
Wastewater Treatment/Control
Level A: Lime Precipitation and Settling
A simplified method of lime precipitation is considered.
The addition of 1.36 kg of hydrated lime per 3.785 cubic
meters (3 lb/1000 gallons) of wastewater is the recommended
dosage. A 15-day supply of lime slurry is drawn off as
needed, mixed with the raw wastewater for 15 minutes in a
mix tank, and discharged to a settling pond for a one-day
retention time. A secondary pond is needed for further
settling before discharge.
The capital and operating costs and assumptions for attain-
ing level A are shown below.
Capital-Cost Components and Assumptions for Level Aj
2 Ponds - dike height of 2 meters (7 feet)
- top width of 3 meters (10 feet)
- capacity of 570 cubic meters (150,600 gallons)
VIII-132
DRAFT
-------
DRAFT
TABLE VI11-27. WATER EFFLUENT TREATMENT COSTS AND RESULTING WASTE-
LOAD CHARACTERISTICS FOR TYPICAL MINE (HYPOTHETICAL)-
ANTIMONY
Antimony Mines
8UBCATEGORY
PLANT SIZE; 10 »300
PLANT AOE;N/A YEARS
.METRIC TONS ( 11,365
PLANT LOCATION: N/A
SHORT TONSI PER YEAR OF ore mined
a. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS(S)/METRIC TON OF PRODUCT*
COSTS ($1000) TO ATTAIN LEVEL
A
29.9
4.2
12.9
1.1
18.2
1.77
B
30.0
4.2
16.1
1.1
21.4
2.08
c
0
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
As
Fe
Sb
Zn
CONCENTRATION (mg/£) (ppm)
RAW
(UN-
TREATED)
25
0.4
1.5
1.0
0.2
AFTER TREATMENT TO LEVEL
A
20
0.07
1.0
1.0
0.1
B
20
0.07
1.0
0.5
0.1
c
o
E
ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0.907
LEVEL A: LIME PRECIPITATION AND SETTLING
LEVEL B: LEVEL A PLUS SULFIDE PRECIPITATION
VIII-133
DRAFT
-------
DRAFT
Lime precipitation unit -
one mix tank with capacity of 8,515 liters
(2,245 gallons)
one mix tank with capacity of 4,165 liters
(1,102 gallons)
Pump - capacity of 0.26 cubic meter (69 gallons) per
minute
Piping - mine to pond - Flow @ 2 meters (6.6 feet)/
second through pipe measuring 5 cm (2 in.)
x 1000 meters (3,280 ft)
pond A to pond B - Flow @ 1 meter (3.3 feet)/
second through pipe measuring 7 cm (2.75 in.)
x 100 meters (328 feet)
Land - 0.21 hectare (0.5 acre)
Operating-Cost Assumptions for Level A_:
Lime - 47.25 metric tons (52.5 short tons)/year
Operating personnel - 3 hr/day
Power - 8.2 kW (11 hp)
Capital Investment;
Facilities
Lagoons $ 3,200
Contingency and contractor's fee 415
Total facility cost $ 3,615
Land 350
Equipment
Lime precipitation unit 6,950
Piping 16.000
Equipment subtotal 22,950
Contingency and contractor's fee 2,985
Total equipment cost
Total Capital Investment
VIII-134
DRAFT
-------
DRAFT
Sodium sulflde - 132 kg (292 lb)/year
Capital Investment;
Equipment
Sulfide precipitation unit $ 100
Contingency and contractor's fee 13
Total Capital Investment $ 113
Annual Cost;
Amortization $ 15
Operation and Maintenance (O&M)
Operating personnel
Equipment repair and maintenance
Materials
Total O&M costs
Total Annual Cost
Titanium Mines
There Is one mine in this subcategory. It produces 1,180,000
metric tons (1,300,000 short tons) of ore annually. The
daily mine discharge is 2,650 cubic meters (700,000 gallons)
of wastewater. One level of technology Is considered for
this subcategory. The cost of implementing this level is
shown In Table VIII-28.
Wastewater Treatment/Contro1
Level Aj Lime Neutralization and Settling
The addition of 0.9 kg of hydrated lime per 3.785 cubic
meters (2 lb/1000 gallons) of wastewater is recommended
for neutralization. The treated effluent is retained for
one day in a settling pond before discharge.
VI11-135
DRAFT
-------
DRAFT
TABLE VI11-28. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR, MINE 9905
SUBCATBOQRY; Titanium Mines
PLANT SIZE; 1.180.000 METRIC TONS (1 f 300 ,000 SHORT TONS) PER YEAR Q|K)re mined
PLANT AQE:30 YEARS PLANT LOCATION; New York
•. COSTS OP TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS($)/METRIC TON OF PRODUCT*
COSTS (91000) TO ATTAIN LEVEL
A
94.3
13.6
23.0
3.0
39.6
0.034
B
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
Oil and Grease
Fluoride
Fe
CONCENTRATION (mg/JU torn)
RAW
(UN-
TREATED)
25
3.0
1.95
1.5
AFTER TREATMENT TO LEVEL
A
20
1.0
1.0
1.0
B
C
D
E
"ORE MINED. TO OBTAIN COSTS/SHORT TON OF PRODUCT. MULTIPLY COSTS SHOWN BY 0*07
LEVEL A: LIME NEUTRALIZATION AND SETTLING
VIII-136
DRAFT
-------
DRAFT
The capital and operating costs and assumptions for attain-
ing this level are shown below.
Capital-Cost Components and Assumptions for Level A_:
Lime precipitation unit
Piping - Flow at 2 m (6.6 feet)/sec through pipe measur-
ing 13 cm (5.1 in.) x 1000 meters (3,280 feet)
Pond - dike height of 3 meters (10 ft)
top width of 3 meters (10 ft)
capacity of 4,000 cubic meters (1,057,000 gallons)
Land - 0.3 hectare (0.75 acre)
Operating-Cost Assumptions for Level A_:
Lime - 222 metric tons (245 short tons)/year
Operating personnel - 3 hr/day
Power - 22.4 kW (30 hp)
Capital Investment;
Facilities
Lagoon $ 7,000
Contingency and contractor's fee 910
Total facility cost $ 7,910
Land 525
Equipment
Lime neutralization unit 43,000
Piping 33.000
Equipment subtotal 76,000
Contingency and contractor's fee 9.880
Total equipment cost 85,880
Total Capital Investment $ 94.315
VIII-137
DRAFT
-------
DRAFT
Annual Cost;
Amortization
Facility $ 805
Equipment 12.800
Total amortization $ 13,605
Operation and Maintenance (O&M)
Land
Operating personnel
Facility repair and maintenance
Equipment repair and maintenance
Materials
Taxes ;
Insurance
Total O&M costs 23,045
Electricity 3.000
Total Annual Cost $ 39,650
Titanium Mills Employing Electrostatic and/or Magnetic
Separation with Gravity and/or Flotation Process
There is only one mill in this subcategory. It mills 1,179,100
metric tons (1,300,000 short tons) annually and has a daily
water discharge of 35,770 cubic meters (9,450,000 gallons).
This mill recycles its process water; however, there is a
seasonal discharge from the tailing-pond system. The dis-
charge is approximately 757 cubic meters (200,000 gallons)
a day for two months of the year.
Two levels of technology are considered. The total costs of
implementing these levels are shown in Table VIII-29.
VIII-138
DRAFT
-------
DRAFT
TABLE VI11-29. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MILL 9905
Titanium Mills Employing Electrostatic .and/or Magnetic
Separation with Gravity and/or Flotation Process
PLANT SIZE: 1.180.000
PLANT APE; 30 YEARS
_METRIC TONS ( 1 . 300. OOP SHORT TONS) PER YEAR OF
PLANT LOCATION; New York __ .
milled
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
BBBB=B-s3B=B=B=====s===s==s=B=SB===!BaB;ss=a
COSTS ($)/METRIC TON OF PRODUCT*
COSTS (S1000) TO ATTAIN LEVEL
1.9
0.20
0.07
0.27
=====
0.0002
B
12.1
1.2
0.4
1.6
0=3===
0.0013
c
D
=====
E
===========
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
^^^S_55BS-5-fi-53^-H-^-^-^-5^^^^3H-535^-5-53H-5-55-B
Fluoride
Cr
Ni
WASTE LOAD (ko/1000 metric tom of product [ore milled] t)
RAW
(UN-
TREATED)
=========
210.000
260
5
10
60
4.000
AFTER TREATMENT TO LEVEL
A
=======0=
16
0.8
0.08
0.08
0.12
0.8
B
x======
0
0
0
0
0
0
c
D
-_-_-_-_-_-_-__-•
E
_-__•_•••-•--•
•••••••«••••••
"ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.907
tTO OBTAIN LOAD AS LB/1000 SHORT TONS OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: DIVERSION DITCHING
LEVEL B: LEVEL A PLUS HOLDING POND (ZERO DISCHARGE)
VIII-139
DRAFT
-------
DRAFT
Wastewater Treatment/Control
Level A: Diversion Ditching
Diversion ditching around one length and one width of the
tailing pond should help to reduce stress in the system due
to seasonal runoff. The exact length and width of the tail-
Ing pond are not known. Therefore, a hypothetical length and
a hypothetical width are assumed.
The capital and operating costs for attaining this level are
shown below and in Table VIII-29.
Capital-Cost Components and Assumptions for Level A_:
Diversion ditching - 1000 meters (3,280 feet)
Capital Investment;
Facilities
Diversion ditching $ 1,650
Contingency and contractor's fee 215
Total Capital Investment $ 1,865
Annual Cost:
Amortization $ 19°
Operation and Maintenance (O&M)
Facility repair and maintenance $ 50
Insurance 20_
Total O&M costs __7Q
Total Annual Cost $ 260
Level IJ: Level A plua Holding Pond (Zero Discharge)
In addition to diversion ditching, a holding pond for the
excess water may be necessary. This pond is located such
that any runoff collected by the diversion ditching would
flow into it and be stored for at least five days.
VIII-140
DRAFT
-------
DRAFT
Water from the holding pond could be discharged after the
suspended solids have settled. The Incremental costs for
the holding pond are shown below. The total costs for
level-B treatment are shown In Table VIII-29.
Capital-Cost Components and Assumptions for Level B_:
Pond - dike height of 3 meters (10 ft)
top width of 3 meters (10 ft)
capacity of 5,678 cubic meters (1,500,000 gallons)
Capital Investment;
Facilities
Lagoon $ 9,000
Contingency and contractor's fee 1.170
Total Capital Investment $ 10,170
Annual Cost;
Amortization $ 1,035
Operation and Maintenance (O&M)
Facility repair and maintenance $ 270
Insurance 100
Total O&M costs 370
Total Annual Cost $ 1,405
Platinum Mine/Mills Employing Dredging
There is one known platinum mine/mill complex. The daily
discharge of wastewater is 32,702 cubic meters (8,640,000
gallons). Annual ore production is 2,267,500 metric tons
(2,500,000 short tons).
Two alternative levels of treatment are considered. The
total costs of implementing these levels are shown in Table
VIII-30.
VTII-141
DRAFT
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DRAFT
TABLE VIII-30. WATER EFFLUENT TREATMENT COSTS AND RESULTING
WASTE-LOAD CHARACTERISTICS FOR MINE/MILL 9904
8UBCATEOORY: Platinum Mine/Mills Employing Dredging
PLANT SIZE; 2,267,500 METRIC TONS (2,500, OOP SHORT TONS) PER YEAR OF material handl ed
PLANT AQE: ^40 YEARS PLANT LOCATION; Alaska
•. COSTS OF TREATMENT TO ATTAIN SPECIFIED LEVELS
COST CATEGORY
TOTAL INVESTED CAPITAL
ANNUAL CAPITAL RECOVERY
ANNUAL OPERATING AND MAINTENANCE
COSTS (EXCLUDING ENERGY AND POWER)
ANNUAL ENERGY AND POWER COSTS
TOTAL ANNUAL COSTS
COSTS (SI/METRIC TON OF PRODUCT*
COSTS (S1000) TO ATTAIN LEVEL
A
18.0
&7
35.6
1.1
39f4
0.017
B
16.8
2.5
73.5
1.3
77.3
0.034
C
D
E
b. RESULTING WASTE-LOAD CHARACTERISTICS
PARAMETER
TSS
WASTE LOAD (kg/1000 metric tons of product (ore milled] *)
RAW
(UN-
TREATED)
0.65
AFTER TREATMENT TO LEVEL
A
0.43
B
0.36
c
D
E
ORE MILLED. TO OBTAIN COSTS/SHORT TON OF PRODUCT (ORE MILLED). MULTIPLY COSTS SHOWN BY 0.807
*TO OBTAIN LOAD AS LB/8HORT TON OF PRODUCT (ORE MILLED). MULTIPLY VALUES SHOWN BY 2.0
LEVEL A: COAGULATION WITH ALUM
LEVEL B: FLOCCULATION
VIII-142
DRAFT
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DRAFT
Wastewater Treatment/Control
Level A_: Coagulation with Alum
It is assumed that the addition of 25 mg/1 of alum is suffi-
cient for coagulation. The necessary settling ponds have
already been constructed.
The alum feed system consists of two mixing tanks, each
having a capacity of 16.5 cubic meters (4,359 gallons), and
two positive-displacement pumps for adding the alum solution.
The alum solution is mixed and fed to the wastewater stream
at a 1-percent solution. The capital and operating costs
and assumptions for attaining this level are shown below.
Capital-Cost Components and Assumptions for Level A_:
two mix tanks, each with capacity of 16.5 cubic meters
(4,359 gallons)
two positive-displacement pumps
Operating-Cost Assumptions for Level A:
Alum - 285 metric tons (315 short tons)/year
Operating personnel - 5 mixes/day @ 1 hr/mix
Power - 8.2 kW (11 hp)
Capital Investment;
Equipment
Alum feed system $ 15,900
Contingency and contractor's fee 2,070
Total Capital Investment $ 17.970
Annual Cost;
Amortization $ 2,680
VIII-143
DRAFT
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DRAFT
Operation and Maintenance (O&M)
Operating personnel $ 15,750
Equipment repair and maintenance 795
Materials 18,900
Insurance 180
Total O&M costs $ 35,625
Electricity
Total Annual Cost $ 39,405
Level B_: Flocculation
The flocculant feed system is the same as that previously
described. However, for this operation, the recommended
dosage of flocculant is 2 mg/1.
Level-B costs are shown in Table VIII-30. This level is not
an addition to level -A treatment, but an alternative for it.
The capital and operating costs and assumptions for attaining
this level are shown below.
Capital-Cost Components and Assumptions for Level B_:
Flocculant feed system
Operating-Cost Assumptions for Level B_:
Flocculant - 23 metric tons (25.2 short tons) /year
Operating personnel - 7 mixes/day @ 1 hr/mix
Power - 9.7 kW (13 hp)
Capital Investment :
Flocculant feed system $ 14,900
Contingency and contractor's fee 1,940
Total Capital Investment $ 16.840
VIII-144
DRAFT
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DRAFT
Annual Cost;
Amortization
Operation and Maintenance (O&M)
Operating personnel
Equipment repair and maintenance
Materials
Insurance
Total O&M costs
$ 2,510
$ 22,050
845
50,400
170
Electricity
Total Annual Cost
73,465
1,300
$ 77,275
VIII-145
DRAFT
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DRAFT
SECTION IX
BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY
AVAILABLE, GUIDELINES AND LIMITATIONS
INTRODUCTION
The effluent limitations which must be achieved by July 1,
1977 are based on the degree of effluent reduction attainable
through the application of the best practicable control
technology currently available. For the ore mining and
dressing industry, this level of technology is based on
the average of the best existing performance by facilities of
various sizes, ages, and processes within each of the industry's
subcategorie's. In Section IV, the ore mining and dressing
industry has been divided into ten major categories. Several
of these major categories have been further subcategorized,
and, for reasons explained in Section IV, each subcategory
will be treated separately for the recommendation of effluent
limitation guidelines and standards of performance.
Best practicable control technology currently available
emphasizes treatment facilities at the end of a manufact-
uring process but also includes the control technology within
the process itself when It is considered to be normal practice
within an industry. Examples of waste management techniques
which are considered normal practice within these industries
are:
(a) manufacturing process controls;
(b) recycle and alternative uses of water; and
(c) recovery and/or reuse of some wastewater constituents.
Consideration was also given to:
(a) the total cost of application of technology in
relation to the effluent reduction benefits to be
achieved from such application;
(b) the size and age of equipment and facilities involved;
(c) the process employed;
(d) the engineering aspects of the application of various
types of control techniques;
(e) process changes; and
(f) nonwater-quality environmental impact (including
energy requirements).
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
IX-1
DRAFT
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DRAFT
The following is a discussion of the best practicable con-
trol technology currently available for each of the subcate-
gories, and the proposed limitations on the pollutants in
their effluents.
GENERAL WATER GUIDELINES
Process Water
Process water is defined as any water contacting the ore,
processing chemicals, intermediate products, byproducts,
or products of a process, Including contact cooling water.
All process water effluents are limited to the pH range of
6.0 to 9.0 unless otherwise specified.
Cooling Water
In the ore mining and dressing Industry, cooling and process
waters are sometimes mixed prior to treatment and discharge.
In other situations, cooling water is discharged separately.
Based on the application of best practicable control technology
currently available, the recommendations for the discharge of
such cooling water are as follows:
An allowed discharge of all non-contact cooling water pro-
vided that the following conditions are met:
(a) Thermal pollution be in accordance with standards
to be set by EPA policies. Excessive thermal rise
in once-through non-contact cooling water in the
ore mining and dressing industry has not been a
significant problem.
(b) All non-contact cooling waters be monitored to
detect leaks of pollutants from the process.
Provisions should be made for treatment to the
standards established for process wastewater
discharges prior to release in the event of such
leaks.
(c) No untreated process waters be added to the cooling
waters prior to discharge.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-2
DRAFT
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DRAFT
The above non-contact cooling water recommendations should
be considered as interim, since this type of water plus
blowdowns from water treatment, boilers, and cooling towers
will be regulated by EPA at a later date as a separate
category.
Storm-Water Runoff
Storm-water runoff may present pollution-control problems
for certain subcategorles. This is true where large stock-
piles of ore to be processed or where waste materials are
piled in areas where the topography may result in runoff to
a nearby stream. Other particular problems include the
size and location of tailing impoundments. Where such impound-
ments occur, their design, construction, and maintenance should
be sufficient to contain the precipitation from the 25-year,
24-hour rainfall event as established by the U.S. National
Weather Service for the area in which such impoundment is
Icoated. That volume of process wastewater which is equivalent
to the volume of precipitation that falls within the impound-
ment in excess of that attributable to the 25-year, 24-hour
rainfall event may be discharged, when such an event occurs.
To preclude unfavorable water-balance conditions resulting
from precipitation and runoff in connection with tailing
impoundments, diversion ditching to prevent natural drainage
or runoff from mingling with process wastewater should be
constructed.
N.?J!£Ji J-18E ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-3
DRAFT
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DRAFT
BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY
AVAILABLE BY ORE CATEGORY AND SUBCATEGORY
Category; Iron Ores
Subcategory; Iron-Ore Mines
This subcategory includes mines operated to obtain iron ore,
regardless of the type of ore or its mode of occurrence. The
limitations proposed here apply to the discharge and treatment
of mine waters. Mine water which is used as process water or
which is conveyed to a mill treatment facility is covered by
limitations proposed for the milling subcategories.
Identification of_ BPCTCA. Best practicable control technology
currently available (BPCTCA) for the control of wastewater from
the mining of iron ore is settling ponds with coagulation/
flocculation systems. At selected locations, it may be possible
to employ settling ponds alone to meet the effluent limitations
specified herein. For acid mine discharge, lime-neutralization
technology is well-understood and is generally applied in other
mining industries.
To implement this technology for use at facilities not already
employing the recommended treatment techniques, settling
impoundments with dispersal systems available for delivery of
flocculating agents will need to be constructed.
Rationale for Selection. At least five iron-ore mines are
known to be currently employing settling impoundments for
treatment of mine wastewater. Suspended-solid removal is
enhanced by coagulation/flocculation systems, as demonstrated
at one mill tailing-impoundment system.
Levels of_ Effluent Reduction Attainable. The levels of
effluent parameters in wastewaters attainable, using the
above technology, are summarized in Table IX-1.
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARC
SUBJECT TOCMANOE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY tPA
IX-4
DRAFT
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DRAFT
TABLE IX-1. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-IRON-ORE MINES
PARAMETER
PH
TSS
Oil and Grease
Dissolved Fe
Total Mn
CONCENTRATION (mg/£) IN EFFLUENT
30-day average
6» to 9«
20
10
1.0
3.0
24-hour maximum
8» to 9«
30
10
2.0
6.0
• Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-5
DRAFT
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DRAFT
Subcategory; •Iron-Ore Mills Employing Physical and/or
Chemical Separation
This subcategory contains iron-ore milling operations that
employ chemical reagents and/or physical methods to beneficiate
iron ore. Mine waters used in milling processes, or mine
waters discharged to mill treatment facilities, are subject
to the limitations proposed below.
Identification of_ BPCTCA. Best practicable control technology
currently available for the control of wastewater from the
milling of iron ore in this subcategory is the use of tailing
ponds with coagulation/flocculatlon systems.
Rationale for Selection. Every known iron-ore beneficiation
facility in this subcategory currently employs tailing-pond-
impoundment treatment facilities. The use and efficiency of
flocculating agents have been demonstrated at one milling
tailing-impoundment system.
Levels of_ Effluent Reduction Attainable. The levels of
effluent reduction attainable through the use of the above
technology are summarized in Table IX-2.
Subcategory; Iron-Ore Mills Employing Magnetic and
Physical Separation (Mesabi Range)
This subcategory includes milling operations operated in
the Mesabi Range of Minnesota.
Identification £f BPCTCA. The best practicable control
technology currently available for the control of wastewater
from this subcategory is no discharge of wastewater.
Rationale for Selection. To implement this technology, no
additional technology is needed, because all mills operating
in this subcategory are currently attaining zero discharge
by the use of large tailing ponds for effective settling of
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-6
DRAFT
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DRAFT
TABLE IX-2. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-IRON-ORE MILLS EMPLOYING
PHYSICAL METHODS AND/OR CHEMICAL REAGENTS
PARAMETER
PH
TSS
Oil and Grease
Dissolved Fe
Total Mn
CONCENTRATION (mg/fc)
IN EFFLUENT
30-day average
6* to 9"
20
10
1.0
2.0
24-hour maximum
6" to 9*
30
10
2.0
4.0
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
—
26.2
13.2
1.32
2.70
lb/1000
short tons
—
52.4
26.4
2.64
5.40
24-hour maximum
kg/1000
metric tons
—
39
13.2
2.64
5.4
lb/1000
short tons
—
78
26.4
5.28
10.8
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-7
DRAFT
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DRAFT
suspended solids prior to reuse and recycle of water back to
the mill for processing. The use of clarifiers and thickeners
to reduce the volume of water discharged to the tailing pond,
and to supply water for recycle back to the milling opera-
tion, can reduce costs incurred in pumping, as well as pipe
size and energy requirements, for implementation of this
technology.
Levels £f_ Effluent Reduction Attainable. Zero discharge of
pollutants can be attained by use of the above technology.
Category; Copper Ores
Subcategory: Copper-Ore Mines
This subcategory includes operations obtaining copper ore
from open-pit, underground, and overburden or ore stripping
operations.
Identification of_ BPCICA. The best practicable control tech-
nology currently available for the discharge of wastewater from
the mining of copper ores is the use of lime precipitation and
settling, with pH adjustment prior to discharge, if necessary.
Implementation of this technology can be enhanced by reduc-
tion or elimination of discharge through the application of
one or more of several techniques: (1) Reuse' of water in
other operations, such as leaching or milling; (2) Control
of mine-water drainage by modification of mining techniques,
construction of diversion structures, or ditching; and (3)
Use of solar radiation to evaporate excess water.
Rationale for Selection. Six primary copper mines discharge
mine water to surface waters. Three of these operations
treat the water by lime precipitation and settling before
its discharge.
Levels of_ Effluent Reduction Attainable. The levels of
effluent parameters in wastewaters attainable using the
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
,IX-8
DRAFT
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DRAFT
above technology are presented in Table IX-3.
Subcategory; Copper Mines Employing Hydrometallurgical
Processes
This subcategory includes mines or operations employing
dump, heap, or in-situ leach processes for the extraction of
copper from ores or ore waste materials.
Identification of_ BPCTCA. The best practicable control tech-
nology currently available in this subcategory is no discharge
of process or mine wastewater.
To achieve this limitation, reuse, recycle, and consumption
of water by evaporation may be employed, resulting in no
discharge of water:
Leach Solution Within the Dump/Ore Bed; Dams, ditches,
and collection ponds are needed to enable the acid-leach
solution to be recovered and fully contained.
Barren Leach Solution; Barren, or used, acid solutions
should be retained in holding ponds and recycled to the
waste ore body for reuse.
Leach Solution Bleed; The use of concrete holding ponds
for precipitation and settling of dissolved solids prior
to evaporation or recycling of water is necessary to
achieve no discharge of these solutions.
Rationale for Selection. All operations surveyed currently
practice recycle or zero discharge of process water.
Levels of_ Effluent Reduction Attainable. Zero discharge is
attainable for solutions resulting from the operations of
this subcategory.
Subcategory: Copper Mills Employing Vat-Leaching Process
This subcategory includes those operations employing the
vat-leach method of copper extraction from ores.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-9
DRAFT
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DRAFT
TABLE IX-3. PARAMETERS SELECTED AND EFFLUENT
LIMITATIONS RECOMMENDED FOR
BPCTCA-COPPER MINES
PARAMETERS
PH
TSS
Oil and Grease
As
Cd
Cu
Fe
H9
Pb
Zn
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6« to 9*
20
1
0.05
0.005
0.05
0.1
0.0002
0.02
0.05
24-hour maximum
6" to 9»
30
2
0.1
0.01
0.1
0.2
0.0004
0.04
0.1
•Value in pH units
WB=rT SJ5ff»M«IIS!!lIlTIVE REC°«MENDATIONS BASEO WON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-10
DRAFT
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DRAFT
Identification of BPCTCA. The best practicable control
technology currently available for this subcategory is no
discharge of process wastewater.
To achieve this limitation, reuse, recycle, and consumption
of process water by evaporation may be implemented. The
total containment of vat-leach solutions in tanks or vats,
with total recycle to the process, is necessary to implement
the above control technology.
Rationale for Selection. Zero discharge of vat-leach barren
solution is currently practiced at all facilities. Of the
four operations examined, three recycle all solutions, and
one reuses the acidic process water in the production of
acid from smelter gases containing sulfur dioxide.
Levels of Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable through the use of the above
control technology.
Subcategory; Copper Mills Employing Froth Flotation (Areas
Where Net Evaporation Equals or Exceeds 76.2 Centimeters
(30 Inches) per Year)
This subcategory indues copper milling operations which
employ the froth-flotation process and are geographically
located in areas of the U.S. which have a net evaporation-
minus-precipitation value equal to, or greater than, 76.2 cm
(30 in.) per year, as defined by the National Weather Service.
Identification £f_ BPCTCA. The best practicable control
technology currently available in this subcategory is no
discharge of process waters through the use of total recy-
cling. In addition to recycle, the reuse and/or consumption
of water by evaporation may be necessary to attain this
effluent limitation.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT ANO ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-11
DRAFT
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DRAFT
Rationale for Selection. All but one of the facilities
surveyed in this subcategory are currently attaining no dis-
charge by employing techniques such as recycle and reuse of
tailing-pond water. One facility surveyed accomplishes
recycle or evaporation of 80 percent of its process water by
means of recycling. As employed in this subcategory, the
practice of recycle is highly desirable for the conservation
of process water.
Levels of Effluent Reduction Attainable. Zero discharge of
process wastewater can be attained by the use of the technology
specified above.
Subcategory: Copper Mills Employing Froth Flotation (Areas
Where Net Evaporation is Less Than 76.2 Centimeters (30 Inches)
Per Year)
This subcategory Includes those copper milling operations which
employ the froth-flotation process and are geographically
located in areas of the U.S. which have a net evaporation-minus-
evaporation value less than 76.2 cm (30 in.) per year, as
defined by the National Weather Service.
Identification of_ BPCTCA. The best practicable control and
treatment technology currently practiced within this subcategory
is lime precipitation and settling, coupled with at least
partial recycle of process wastewater.
Rationale for Selection. Within this subcategory, there are
five major copper mills currently practicing recycle of zero
to 90 percent of the process-water volume. Two of these opera-
tions treat their process wastewater with additional lime prior
to settling in a tailing impoundment.
Levels of Effluent Reduction Attainable. The levels of
concentration and waste loading attainable by implementation
of the technology recommended above are presented in Table IX-4.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-12
DRAFT
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DRAFT
TABLE IX-4. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS RECOMMENDED
FOR BPCTCA-COPPER MILLS USING FROTH FLOTATION (NET
EVAPORATION LESS THAN 76.2 CM [30 IN.] PER YEAR)
PARAMETER
pH
TSS
Oil and Grease
CN
Cd
Cu
Fe
Hg
Pb
Zn
CONCENTRATION (mg/i)
IN EFFLUENT
30-day average
6* to 9*
10
1
0.01
0.005
0.05
0.1
0.001
0.05
0.05
24-hour maximum
6' to 9*
20
2
0.02
0.01
0.1
0.2
0.002
0.1
0.1
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
_
35
3.5
0.035
0.02
0.2
0.35
0.00035
0.2
0.2
lb/1000
short tons
_
70
7
0.07
0.04
0.4
0.7
0.0007
0.4
0.4
24-hour maximum
kg/1000
metric tons
_
70
7
0.07
0.04
0.4
0.7
0.0007
0.4
0.4
lb/1000
short tons
_
140
14
0.15
0.08
0.8
1.5
0.0015
0.8
0.8
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
IX-13
DRAFT
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DRAFT
Category; Lead and Zinc Ores
Subcategory; Lead and/or Zinc Mines Having No Solubilization
Potential
This subcategory Includes mines operated for the recovery
of lead and/or zinc ores where conditions for extensive
solubilization of metals do not exist.
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory is a
settling- or sedimentation-pond system. To attain the levels
of treatment specified here, a ten-day retention time in a
primary sedimentation pond will be required, with discharge
to a secondary settling (or "polishing") pond having a two-
day retention time. Reuse of mine water as mill process water,
or treatment in a mill treatment system, is an acceptable
alternative, provided that the levels specified for the mill
subcategory can be met, and provided that unfavorable water
balances are not caused for the mill process-water treatment
system. Treatment by aeration may be necessary at selected
locations to meet recommended levels of COD.
Rationale for Selection. Currently, one-half of all operations
in this subcategory practice the above treatment and control
technology.
Levels of Effluent Reduction Attainable. The levels of
effluent reduction attainable in this subcategory through the
application of the above technology are presented in Table
1X-5.
Subcategory! Lead and/or Zinc Mines Having Solubilization
Potential
This subcategory includes mines operated for the recovery of
lead and/or zinc ores where conditions for extensive solubili-
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-14
DRAFT
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DRAFT
TABLE IX-5. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-LEAD AND ZINC MINES
HAVING NO SOLUBILIZATION POTENTIAL
PARAMETER
PH
TSS
COD
Oil and Grease
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
CONCENTRATION (mg/ft)
IN EFFLUENT
30-day average
6" to 9"
20
50
10
0.01
0.1
0.05
1.0
0.0001
0.5
0.1
0.1
24-hour maximum
6» to 9»
30
100
10
0.02
0.2
0.1
2.0
0.0002
1.0
0.2
0.2
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-15
DRAFT
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DRAFT
zatlon of heavy metals are present, resulting in high
concentrations and waste loading for these parameters.
Extensive solubilization conditions are defined as total
heavy-metal concentration in untreated mine wastewater equal
to or exceeding 2 mg/1 for the sum of the concentrations
of Cd, Cr, Cu, Fb, and Zn.
Identification of BPCTCA. The best practicable control
technology currently available for this subcategory is
the use of lime precipitation in combination with a settling
or sedimentation pond with a retention time of ten days. An
alternative technology which may be employed is the use of
high-density sludge neutralization with a clarifier (8-hour
retention time).
Rationale for Selection. The levels proposed for this sub-
category are based on application of this technology at one
zinc/copper mining operation, as well as on extensive applica-
tion of this treatment at lead/zinc/copper mines in Canada,
both at full-scale operations and in pilot-evaluation
facilities (References 64, 69, and 70).
Levels of_ Effluent Reduction Attainable. The levels of
effluent reduction attainable in this subcategory through the
application of the above technology are presented in Table
IX-6.
Subcategory; Lead and/or Zinc Mills
This subcategory includes all mills operated for the recovery
of lead and/or zinc concentrates. All current operations in
this subcategory employ the process of froth flotation for the
beneficiation of ores.
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory is a
settling- or sedimentation-pond system with a primary tailing
pond (10-day retention) and a secondary-settling or "polishing"
pond (two-day retention).
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN .THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-16
DRAFT
-------
DRAFT
TABLE IX-6. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-LEAD AND ZINC MINES
HAVING SOLUBILIZATION POTENTIAL
PARAMETER
PH
TSS
COD
Oil and Grease
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
CONCENTRATION (mg/2.)
IN EFFLUENT
30-day average
6" to 9»
20
50
10
0.05
0.2
0.05
1.0
0.0005
0.5
0.2
0.5
24-hour maximum
6»to9»'
30
100
10
0.1
0.5
0.1
2.0
0.001
1.0
0.3
1.0
•Value in pH units
* J«5.1EA^IIE«TATIVE "COMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX- 17
DRAFT
-------
DRAFT
Rationale for Selection. Currently, approximately 20 percent
(at least six of the operations surveyed) have implemented
the above technology. Many operations actually currently
exceed the ten-day retention time specified in the above
technology.
Levels of_ Effluent Reduction Attainable. The levels of
effluent reduction attainable by application of the above
technology are presented in Table IX-7.
Category; Gold Ores
Subcategory: Gold Mines (Alone)
This subcategory Includes mines operated for the recovery of
gold ores by open-pit or underground methods. Discharge of
mine wastewater into mill waste-treatment systems, or reuse
of mine water in the milling process, is acceptable provided
that effluent limitations for the mill subcategory are met,
and provided that unfavorable water balances affecting mill
waste-treatment systems do not result.
Identification o_f_ BPCTCA. The best practicable control
technology currently available for the discharge of wastewater
resulting from the mining of gold ores is the use of chemical-
precipitation methods in conjunction with settling-pond
removal of suspended solids and precipitates. The use of lime-
precipitation or sulfide-precipitation methods to effect
removal of heavy-metal constituents are equally acceptable
methods of implementing this technology. Settling of suspended
solids may be performed either in settling impoundments or by
the use of mechanical clarification equipment to meet the
levels of effluent reduction specified here.
Rationale for Selection. Treatment of mine wastewater as
currently practiced by these operations varies from non-existent
to the use of settling impoundments. Because the level of
treatment which results is uniformly inadequate, the well-
demonstrated technology of chemical precipitation is specified
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA
1X-18
DRAFT
-------
DRAFT
TABLE IX-7. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-LEAD AND/OR ZINC MILLS
PARAMETER
PH
TSS
COD
Oil and Grease
Cyanide
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
CONCENTRATION (mg/H)
IN EFFLUENT
30-day average
6" to 9'
10
44
4
0.01
0.005
0.015
0.02
0.15
0.0001
0.1
0.03
0.05
24-hour maximum
6* to 9"
20
100
10
0.02
0.01
0.03
0.05
0.30
0.0002
0.2
0.1
0.1
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
-
15
130
13
0.04
0.01
0.05
0.03
0.25
0.0004
0.5
0.07
0.1
lb/1000
short tons
-
30
260
26
0.08
0.02
0.1
0.06
0.5
0.0008
1.0
0.14
0.2
24-hour maximum
kg/1000
metric tons
-
30
260
26
0.08
0.02
0.1
0.06
0.5
0.0008
1.0
0.14
0.2
lb/1000
short tons
-
60
520
52
0.16
0.04
0.2
0.12
1.0
0.0016
2.0
0.28
0.4
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-19
DRAFT
-------
DRAFT
because of its demonstrated use and efficiency of treatment
attained in other categories of the ore mining and dressing
industry.
Levels of Effluent Reduction Attainable. The levels of effluent
reduction attainable through the use of the above technology
are presented in Table IX-8.
Subcategory; Mills or Mine/Mills Employing Cyanidation
This subcategory includes operations obtaining gold by the
cyanldation process of extraction from gold ores.
Identification of_ BPCTCA. The best practicable control
technology currently available in this subcategory is no
discharge of process wastewater.
Implementation of this control technology may be achieved in
either of two ways: impoundment or complete recycle of process
wastewater. At some locations, destruction of cyanide by
alkaline chlorination may be necessary if the presence of
cyanide in recycled water adversely affects the process.
Rationale for Selection. Of the two mills currently
employing cyanidation processing, one operation has achieved
zero discharge by impoundment and recycle of process wastewater.
An important engineering aspect of a zero-discharge system is
the design of the water-management system. A recycle system
generally involves discharge of mill process water to a tailing
pond for settling of solids and subsequent decantation and
pumping of clarified pond water back to the mill. Inputs to
this system may result from precipitation and runoff. It is
standard engineering practice in the industry to design tailing
ponds to accommodate periodic excesses of water. In addition,
relatively simple systems to divert runoff away from the
tailing pond can be (and are) employed. Diversion ditching is
normally used to convey water away from the tailing-pond area.
A measure of control over the quality of the reclaim water is
normally maintained by the use of d two-celled pond system.
Tailings are discharged to the first pond for settling; then,
the decant from this pond Is collected Ln the second pond,
which serves as ;i surge pond Ln tin? recycle system.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASEDI UPON> INFORMATIONI IN JH'S REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-20
DRAFT
-------
DRAFT
TABLE IX-8. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-GOLD MINES
PARAMETER
pH
TSS
Cu
Fe
H8
Zn
CONCENTRATION (mg/°)
IN EFFLUENT
30-day average
6» to 9«
20
0.05
1.0
0.0005
0.15
24-hour maximum
6" to 9"
30
0.1
2.0
0.001
0.25
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-21
DRAFT
-------
DRAFT
Level of Effluent Reduction Attainable. Zero discharge of
pollutants is attainable by implementation of the above control
technology.
Subcategory; Mills or Mine/Mills Employing Amalgamation
This subcategory includes mill or mine/mill complexes
extracting gold by use of the amalgamation process.
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory is chemical
precipitation in conjunction with sedimentation or tailing
impoundment, with in-process recycle of the mercury reagent
in the amalgamation process. Adjustment of the pH of waste-
waters may be necessary.
Chemical precipitation may be accomplished by the use of lime-
or sulfide—precipitation systems.
Rationale for Selection. Currently, there is one operating
facility employing the amalgamation process for gold extraction.
To effect removal of heavy meta]s, the use of chemical preci-
pitation methods in conjunction witli tailing impoundment is
well-documented and has been demonstrated in the ore mining
and dressing industry at other locations.
Levels of Effluent Reduction Attainable. The levels of
effluent reduction attainable for this subcategory by use of
the above technology are presented in Table IX-9.
Subcategory; Mills or Mine/Mills Beneficiating Gold Ores
(Froth-Flo tat ion Process)
This subcategory includes mills or mine/mill complexes oper-
ated for the beneficiation of gold ores by froth flotation.
The single operation employing tlus mothod also practices
cyanldation of tailings from Uu> flotation circuit by agita-
tion/cyanidation.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS UAStO UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVkU AND FURTHER INTERNAL REVIEW BY EPA.
DRAFT
-------
DRAFT
TABLE IX-9. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-GOLD MINES USING
AMALGAMATION PROCESS
PARAMETER
pH
TSS
Cu
Fe
Hg
Zn
CONCENTRATION (mg/fcl
IN EFFLUENT
30-day average
6« to 9"
20
0.05
1.0
0.0005
0.15
24-hour maximum
6* to 9*
30
0.1
2.0
0.001
0.25
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
—
120
0.29
5.8
0.003
0.87
lb/1000
short tons
—
240
0.58
11.6
0.006
1.45
24-hour maximum
kg/1000
metric tons
—
170
0.58
11.6
0.006
1.74
lb/1000
short tons
—
340
1.16
23.2
0.012
2.9
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-2 3
DRAFT
-------
DRAFT
Identification of BPCTCA. The best practicable control tech-
nology currently available in this subcategory is the use of
tailing impoundments in conjunction with diversion ditching
and partial recycle of process water to reduce discharge
volume. If cyanide is present in wastewater, alkaline chlori-
nation for cyanide destruction of discharge waters is required.
Rationale for Selection. The single operating facility in
this subcategory currently practices impoundment during approxi-
mately nine to ten months of the year. Reduction of discharge
volume on a seasonal basis is possible by diversion of runoff
to prevent accumulation of excess water in tailing impoundments
or by recycle of tailing decant water In conjunction with
alkaline chlorination to remove cyanide (which would inter-
fere with the flotation of the gold-bearing ore).
Levels of_ Effluent Reduction Attainable. The levels of efflu-
ent reduction attainable for this subcategory by use of the
above technology are presented in Table IX-10.
Subcategory; Mills or Mine/Mills Employing Gravity-Separation
Methods
This subcategory includes mills or mine/mills beneficiating
gold ore by gravity-separation methods. This subcategory
also includes placer or dredge mining or concentrating opera-
tions, as well as hydraulic-mining operations, regardless of
size.
Identification of_ BPCTCA. The best practicable control tech-
nology currently available for this subcategory is the use
of settling or tailing Impoundments for settling of suspended
solids. An alternative technology which may be employed is
the pumping of wastewater from dredging operations back to
a tailing-disposal area for filtration through sands and
gravels. At some operations, it may be necessary to employ
flocculating agents to enhance settling of suspended solids
to meet the effluent limitations specified herein.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-24
DRAFT
-------
DRAFT
TABLE IX-10. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-GOLD MILLS USING
FLOTATION PROCESS
PARAMETER
PH
TSS
Cyanide
Cu
Fe
Hg
Zn
CONCENTRATION (mg/J,)
IN EFFLUENT
30-day average
6" to 9«
20
0.01
O.OS
1.0
0.0005
0.15
24-hour maximum
6« to 9*
30
0.02
0.1
2.0
0.001
0.25
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
_
6
0.001
0.01
0.3
0.0001
0.045
lb/1000
thort tons
_
12
0.002
0.02
0.6
0.0002
0.09
24-hour maximum
kg/1000
metric tons
_
8.9
0.002
0.02
0.6
0.0002
0.074
lb/1000
short tons
_
17.8
0.004
0.04
1.2
0.0004
0.15
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-25
DRAFT
-------
DRAFT
Rationale for Selection. The practice specified is the best
technology now utilized at several operations recovering
gold by gravity-separation methods. The prevailing practice
in this industry subcategory is direct discharge of waste-
water.
Levels of_ Effluent Reduction Attainable. The levels of efflu-
ent reduction attainable employing the above technology are
given in Table IX-11.
Subcategory; Mill Operations Where Gold JIB Recovered as_
Byproduct <>f_ Base-Metal Milling Operation
This subcategory includes facilities operated primarily to
obtain concentrates of base metals (usually lead, zinc, or
copper). Gold may be obtained from the base-metal concen-
trates at a_ refinery or a_ smelter.
Identification of BPCTCA. No separate technology or limita-
tions are recommended for this subcategory. Instead, the
limitations and technology for each applicable base-metal
subcategory are recommended, because the characteristics of
the primary ore and processes employed dominate the waste-
water parameters.
Category; Silver Ores
Subcategory; Silver Mines (Alone)
This subcategory includes facilities which are operated for
the mining of silver ores by open-pit or underground methods.
Discharge of mine waters into mill treatment systems, or for
reuse as process water, is covered in the applicable limita-
tion guidelines for milling subcategories.
Identification o£ BPCTCA. The best practicable control tech-
nology currently available for silver-mine discharges is use
of chemical-precipitation methods for heavy-metal removal in
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-26
DRAFT
-------
DRAFT
TABLE IX-11. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-GOLD MINES OR MILLS
USING GRAVITY-SEPARATION METHODS
PARAMETER
PH
TSS
CONCENTRATION (mg/fc)
IN EFFLUENT
30-day average
6" to 9'
30
24-hour maximum
6- to 9"
40
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
36
lb/1000
short tons
72
24-hour maximum
kg/1000
metric tons
48
lb/1000
short tons
96
•Value in pH unit*
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
IX-27
DRAFT
-------
DRAFT
conjunction with the use of settling pond(s) for suspended-
solid removal. Lime- or sulfide-precipitation methods may
be used for chemical-precipitation treatment, while an alter-
native suspended-solid treatment is the use of mechanical
clarifiers for suspended-solid removal. At selected locations,
pH adjustment of discharge waters may be necessary.
Rationale for Selection. Current treatment practices in the
silver mining industry range from no treatment to use of
settling ponds where discharge to mill treatment systems or
use in a mill process is not practiced. Treatment practices
are considered to be uniformly inadequate for the removal of
pollutants present in silver-mine wastewater. Therefore,
chemical-treatment methods which have been demonstrated to
be effective in other segments of the ore mining and dressing
industry have been adopted in addition to use of settling
ponds.
Levels of_ Effluent Reduction Attainable. The levels of efflu-
ent reduction attainable through the use of the above tech-
nology are presented in Table IX-12.
Subcategory: Mills or_ Mine/Mi11s Employing Froth Flotation
for Beneficiation o_f Silver Ores
This subcategory includes those milling operations employing
the forth-flotation process for extraction of silver concen-
trates from silver ores.
Identification of_ BPCTCA. The best practicable control tech-
nology currently available is zero discharge by use of total
recycle of process water or total impoundment.
Rationale for Selection. Currently, two silver mills are
recycling their process water. One mill reclaims all of
its tailing-pond decant, while the second presently reclaims
60 percent of its tailing-pond decant. Recycle of all process
water is currently technically achievable at the second mill;
however, it is currently less expensive for the operator to
use fresh water Instead of recycled process water from an
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-2 8
DRAFT
-------
DRAFT
TABLE IX-12. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-SILVER MINES (ALONE)
PARAMETER
pH
TSS
As
Cu
Fa
Pb
Sb
Zn
CONCENTRATION (mg/X,)
IN EFFLUENT
30-day average
6» to 9»
20
0.07
0.05
1.0
0.1
0.5
0.15
24-hour maximum
6« to 9"
30
0.15
0.1
2.0
0.2
0.8
0.25
•Value in pH units
tinnce. TMBRE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
•UUUT TOCMANOES!!££UPO«?COMMENT?RECE7vED AND FURTHER INTERNAL REVIEW BY EPA.
IX-29
DRAFT
-------
DRAFT
Impoundment. The technical feasibility of achieving no dis-
charge Is discussed in detail in Section VII.
Levels of Effluent Reduction Attainable. Implementation of
the above technology will result in no discharge of pollu-
tants to surface waters.
Subcategory! Mills or Mine/Mills Using Cyanidation for
Recovery o_f Silver
This subcategory includes those milling operations employing
the cyanidation process for recovery of silver from silver
ores. The recovery of silver by this method is usually done
in connection with gold recovery.
Identification of_ BPCTCA. The best practicable control tech-
nology currently available for this subcategory is attainment
of zero discharge by use of recycle or total impoundment of
process water.
To Implement this technology, recycling in the process reagent
circuits may be necessary to achieve economy in reagent use
and avoid high concentrations of cyanide in recycled process
water. Diversion of runoff may be necessary in some geographic
locations to reduce the volume of water which must be impounded.
Rationale for Selection. Currently, no treatment technology
is being practiced at the one known discharging milling
establishment in this subcategory. However, the attainment
of zero discharge at a cyanidation mill in the gold category
has been well-documented and demonstrated to be effective
for use in similar operations involving the cyanidation pro-
cess at silver mills. In addition, comparison of percentage
recovery for a mill employing cyanidation for gold/silver
recovery with no treatment to that of a gold mill practicing
total recycle Indicates that no loss of recovery is necessary
with recycling of process water.
Levels £f Effluent Reduction Attainable. Zero discharge of
pollutants to surface waters will result with employment of
the above technology.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-30
DRAFT
-------
DRAFT
Subcategory; Mills or_ Mine/Mills Concentrating Silver Ores
by_ Amalgamation
This subcategory includes milling operations engaged in the
recovery of silver by use of amalgamation of silver ores.
This process is often employed for the extraction of both
gold and silver from ores.
Identification £f_ BPCTCA. The best practicable control tech-
nology currently available is lime precipitation for
metal removal in conjunction with the use of settling impound-
ments. To achieve reduction of mercury concentrations in
process wastewater, in-process recycling within the mercury
reagent circuit should be used. The adjustment of pH of dis-
charge waters may be necessary at selected operations to
achieve pH limitations.
Rationale for Selection. At present, there is one operation
utilizing amalgamation for the recovery of silver. This
operation currently employs two sedimentation ponds, but
metal removal by this method is inadequate. The use of
chemical-precipitation methods has been well-demonstrated
in the ore mining and dressing industry to be effective in
reduction of heavy metal pollutant concentrations.
Levels of_ Effluent Reduction Attainable. The levels of
pollutant concentrations attainable by use of the above
methods are presented in Table IX-13.
Subcategory; Mills Recovering Silver bv_ Gravity-Separation
Methods
This subcategory includes those operations operated for the
recovery of silver by gravity-separation methods. Silver
is recovered in minor amounts as part of gold placer opera-
tions.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-31
DRAFT
-------
DRAFT
TABLE IX-13. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-SILVER MILLS USING
AMALGAMATION PROCESS
PARAMETER
PH
TSS
Cu
iF« '
>
Zn
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6* to 9*
20
O.OS
1.0
0.0005
0.15
24-hour maximum
6»to9*
30
0.1
2.0
0.001
0.25
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric toni
_
116
0.29
5.8
0.003
0.87
lb/1000
thort tons
_
232
0.58
11.6
0.006
1.45
24-hour maximum
kg/1000
metric toni
—
175
0.58
11.6
0.006
1.45
lb/1000
short torn
_
350 "
1.16
23.2
0.012
2.90
•Value in pH unit*
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
IX-32
DRAFT
-------
DRAFT
Identification of_ BPCTCA. The best practicable control tech-
nology currently available for this subcategory is the use
of settling or tailing impoundments for settling of suspended
solids. An alternative technology which may be employed is
the pumping of wastewater from dredging operations back to
a tailing-disposal area for filtration through sands and
gravels. At some operations, it may be necessary to enhance
the settling of suspended solids to meet the effluent limi-
tations specified here.
Rationale for Selection. The use of settling impoundments
such as dredge ponds or tailing impoundments is the best
technology now utilized in connection with gravity methods
of extraction of silver in the dredges or placer mining
industry today.
Levels of_ Effluent Reduction Attainable. The levels of
effluent reduction attainable employing the above technology
are given in Table IX-14.
Subcategory; Mill Operations Where Silver is^ Recovered a£
Byproduct of_ Base-Metal Milling Operation
This subcategory Includes facilities operated primarily to
obtain concentrates of base metals (usually, lead, zinc,
or copper). Silver may be obtained from the base-metal con-
centrates ait a_ refinery or_ a_ smelter.
Identification £f_ BPCTCA. No separate technology or limita-
tions are recommended for this subcategory. Instead, limi-
tations and technology for each applicable base-metal sub-
category are recommended, because the characteristics of the
primary ore and processes employed dominate the wastewater
parameters.
NOTICi: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-3 3
DRAFT
-------
DRAFT
TABLE IX-14. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-SILVER MILLS USING
GRAVITY SEPARATION
PARAMETER
PH
TSS
CONCENTRATION (rng/H)
IN EFFLUENT
30-day average
6" to 9"
30
24-hour maximum
6" to 9'
40
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
0.432
lb/1000
short tons
0.864
24-hour maximum
kg/1000
metric tons
0.576
lb/1000
short tons
1.152
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND -Kb
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-34
DRAFT
-------
DRAFT
Category; Bauxite Ores
This category includes establishments engaged in the mining of
bauxite ores. No beneficiation of these ores is currently
practiced, with the exception of crushing, grinding, and
washing activities at the two currently operating sites. No
subcategories were identified in this category.
Identification of_ BPCTCA. The best practicable control
technology currently available for the removal of pollutants
present in mine drainage in the bauxite mining Industry is use
of lime precipitation and settling.
Rationale for Selection. The two currently operated facilities
are both practicing lime neutralization and/or precipitation
on most mine effluents at the present time. The efficiency of
this method of treatment has been well-demonstrated in these
operations on both full- and pilot-scale bases. In the case of
alkaline ground-water drainage, aeration of wastewater may be
necessary to convert iron to a form more amenable to lime
precipitation in order to meet the limitations specified here.
Levels of_ Effluent Reduction Attainable. The concentration
levels attainable through implementation of BPCTCA are pre-
sented in Table IX-15.
Category: Ferroalloy Ores
Subcategory; Mines Extracting Ores of Ferroalloy Metals
This subcategory includes mines operated to obtain ferroalloy
metals and which discharge to surface waters of the U.S.,
regardless of the particular ferroalloy metal involved. The
ferroalloy-metal ores covered here include chromium, cobalt,
columbium/tantalum, manganese, molybdenum, nickel, tungsten,
and vanadium (recovered alone). Vanadium is also recovered
as a byproduct of uranium mining and milling operations under
NRC licensing regulations as part of radioactive-ore extraction.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-3 5
DRAFT
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DRAFT
TABLE IX-15. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-BAUXITE MINES
(ACID OR ALKALINE MINE DRAINAGE)
PARAMETER
PH
TSS
Al
Fe
Mn
Ni
Zn
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6" to 9»'
20
0.35
0.30
1.5
0.10
0.05
24-hour maximum
6» to 9"
30
0.70
0.60
3.0
0.20
0.10
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-36
DRAFT
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DRAFT
Identification pjE BPCTCA. The best practicable control
technology currently available for this subcategory is the
use of lime precipitation in conjunction with a settling
pond. The use of a mechanical clari-flocculator or
equivalent equipment is an acceptable alternative for suspended-
solid removal.
For use of this technolgoy, liming prior to removal of suspended
solids is desirable. Treatment requirements can also be
minimized by the use of diversion structures to prevent surface
waters from entering the mine, good housekeeping practices to
Isolate wastewater from oil and grease spills associated with
machinery, and reduction of contact with broken ore and rock
dust.
Rationale for Selection. Sedimentation or settling impound-
ments are widely used in the ore mining and dressing Industry
for suspended-solid removal. The use of lime for pH adjustment
and precipitation of metals is both an effective practice and
a standard, longstanding practice at many milling establishments.
Because metal removal by settling methods alone is Inadequate
at most ferroalloy-ore mines, relatively simple lime-precipi-
tation methods are recommended for use. Engineering difficulties
may be encountered where large mine flows coincide with limited
land availability, but the employment of mechanical clarifying/
flocculating devices is an acceptable alternative. At one
ferroalloy mining site, a mechanical device for settling
suspended solids was used, and levels of less than 15 mg/1 of
suspended solids were attained. Adjustment of pH to the range
of 8.5 to 9, with removal of solid precipitates, should enable
attainment of the effluent levels recommended here.
Levels of_ Effluent Reduction Attainable. The levels of
effluent reduction attainable and the parameters selected for
control for this subcategory are presented in Table IX-16.
Note that no limitations for molybdenum and vanadium are recom-
mended for BPCTCA, because these metals are not effectively
removed by currently available treatment. Discharge concentra-
tions of these metals will be minimized by sound practice (as
discussed above), and by avoiding leaching of ores exposed
for long periods to oxidizing conditions.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-37
DRAFT
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DRAFT
TABLE IX-16. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-FERROALLOY-ORE MINES
PARAMETER
PH
TSS
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
V
Zn
Fluoride
CONCENTRATION (mg/fc)
IN EFFLUENT
30-day average
6Mo9»
20
3
0.4
0.05
0.05
1.0
t
0.10
t
0.15
10
24-hour maximum
6»to9"
30
5
0.7
0.07
0.07
1.6
t
0.15
t
0.2
20
•Value in pH units
tNo limitations proposed for BPCTCA
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-3 8
DRAFT
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DRAFT
Subcategory; Mills QIC Mine/Mills Processing Less Than 5.000
Metric Tons (5.512 Short Tons) Per Year of Ferroalloy Ores By_
Methods Other Than Ore Leaching
This subcategory includes those operations processing less
than 5,000 metric tons (5,512 short tons) of ore per year by
methods other than ore leaching. Operations in this subcategory
are confined primarily to intermittent operation, and benefi-
ciation of the ores is frequently performed by gravity methods.
Tungsten-ore mines/mills are the prime components of this
subcategory.
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory is the use
of settling or tailing ponds in conjunction with pH control.
Rationale for Selection. Operations in this subcategory are,
in general, intermittent; economically marginal; and of a low
level of technical sophistication. Present practice at these
operations is predominantly direct discharge without treatment.
Data gathered here indicate that current practices in this
subcategory are uniformly inadequate. Therefore, the relatively
simple, well-demonstrated and well-documented technology of
tailing or settling impoundment with pH control is recommended.
The use of this technology will represent a major improvement
over present practice at most operations in this subcategory.
Mine water, where available, should be used for mill feed, and
the mine and mill waters should be treated together. Neutraliza-
tion and suspended-solid removal will result in some degree of
removal of dissolved metals, in addition to reduction of COD
and other waste components, by use of this technology, although
monitoring of these parameters is not recommended here.
Levels of Effluent Reduction Attainable. The parameters
selected and the recommended effluent levels attainable by use
of the above technology in this subcategory are presented in
Table IX-17.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-3 9
DRAFT
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DRAFT
TABLE IX-17. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS RECOMMENDED
FOR BPCTCA-FERROALLOY-ORE MILLS TREATING LESS THAN 5.000
METRIC TONS (5,512 SHORT TONS) PER YEAR
PARAMETER
PH
TSS
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6" to 9*
30
24-hour maximum
6" to 9*
50
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
90
lb/1000
short tons
180
24-hour maximum
kg/1000
metric tons
ISO
lb/1000
short tons
300
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND APE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-40
DRAFT
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DRAFT
Subcategory; Mills Processing More Than 5.000 Metric Tons
(5.512 Short Tons) of_ Ferroalloy Ores Per Year By Physical
Methods —
This subcategory includes mill or mine/mill facilities pro-
cessing more than 5,000 metric tons (5,512 short tons) of
ferroalloy ores per year by purely physical methods. These
methods include ore crushing, washing, jigging, heavy-media
and gravity separation, and magnetic and electrostatic
separation.
Identification of BPCTCA. The best practicable control tech-
nology currently available for this subcategory is the use
of process-water recycle practices in conjunction with tail-
ing Impoundment (including runoff diversion), lime precipita-
tion, flocculation, and secondary settling.
Total recycle of process water with zero discharge is a possible
viable alternative technology for many operations of this type.
Rationale for Selection. The recommended BPCTCA technology
has been in large-scale use within the ore mining and dress-
ing industry, and its successful implementation on waste
streams is expected to pose no significant technical problems.
Treatment to BPCTCA levels is achieved at the largest industry
representative of this subcategory, although natural alka-
linity and low soluble ore contents obviate the need for the
practice of lime precipitation at that site. Recycle of process
waters is currently practiced at many sites and is limited
technically only where wet scrubbers are used for air-pollution
control on ore-drying or ore-roasting installations. In such
operations, dissolved-solid buildup in the scrubber-water
circuit could lead to decreased effectiveness in scrubbing
and consequent Increased maintenance.
The extent to which a given site may approach zero discharge
in this subcategory is limited by climatic and topographic
factors affecting casual water influx to the system. Diver-
sion of runoff may reduce the volumes of Impounded and treated
water. Total recycle with ho process-water discharge reportedly
will be practiced upon reopening of a manganiferous-ore con-
centrator, which is expected to occur some time during 1975.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-41
DRAFT
-------
DRAFT
Levels of Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable by implementation of this technology are pre-
sented in Table IX-18.
Subcategory; Mills Processing More Than 5.000 Metric Tons
(5.512 Short Tons) of Ferroalloy Ores Per Year By_ Flotation
Methods
This subcategory includes mills processing more than 5,000
metric tons (5,512 short tons) of ferroalloy ores per year
by froth-flotation methods.
Identification of_ BPCTCA. The best practicable control tech-
nology currently available for this subcategory Includes the
use of primary settling or tailing ponds (with runoff diver-
sion) in conjunction with lime precipitation and secondary
settling. Flocculation may be necessary at selected locations
to meet suspended-solid limitations.
Lime precipitation will not be necessary at some sites,
because their flotation circuits are maintained at alkaline
pH. The use of flocculants may be occasionally necessary
to achieve suspended-solid limitations.
Rationale for Selection. The recommended treatment and con-
trol technology is currently in use within the ore mining
and dressing industry, and its successful implementation for
waste streams from mills in this subcategory is expected to
pose no significant technical problems. Because of alkaline
pH at flotation mills and the use of settling ponds with
adequate retention time, levels recommended here are currently
being achieved at sites within the subcategory.
Recycle of process water is not recommended as BPCTCA for
these operations, since nonsulflde-ore flotation operations
would require extensive process development work and process
modification. In addition, no successful operations are known
at present which employ total recycle for fatty-acid flotation
of scheellte.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-42
DRAFT
-------
DRAFT
TABLE IX-18. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-FERROALLOY-ORE MILLS
TREATING MORE THAN 5,000 METRIC TONS
(5,512 SHORT TONS) PER YEAR BY PHYSICAL PROCESSING
PARAMETER
PM
TSS
Oil and Grease
At
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
8" to 9"
20
3
0.4
O.OB
0.06
1.0
t
0.10
0.1
10
24-hour maximum
6' to 9"
30
6
0.8
0.1
0.1
1.5
t
0.2
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
_
20
3
0.4
O.OS
O.OS
1.0
_
0.10
0.1
10
lb/1000
short tons
— —
40
6
0.8
0.1
0.1
2.0
_
0.2
0.2
20
24-hour maximum
kg/1000
metric tons
_
40
10
0.8
0.1
0.1
3.0
_
0.2
0.4
20
lb/1000
short tons
_
80
20
1.6
0.2
0.2
6.0
_
0.4
0.8
40
•Value In pH units
fNo limitations proposed for BPCTCA
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-43
DRAFT
-------
DRAFT
Total recycle is a viable alternative technology for some
mills within the subcategory—particularly, since treatment
of smaller wastewater volumes may, in some cases, offer sub-
stantial economic advantages.
Levels £f Effluent Reduction Attainable. The parameters
selected and levels of effluent reduction attainable by
implementation of BPCTCA are presented in Table IX-19.
Levels of cyanide and COD can be controlled by control of
reagent usage, and by natural aeration and degradation during
delivery of tailings to impoundment and during retention in
settling ponds.
Subcategory;! Mills Processing Ferroalloy Ores By Leaching
Techniques —*- *•
This subcategory includes mills processing ferroalloy ores
by leaching techniques (whether acid or alkaline) and associa-
ted chemical-beneficiation techniques.
Identification of BPCTCA. The best practicable control tech-
nology currently available for this subcategory includes
tailing-pond impoundment for primary settling, in conjunction
with lime precipitation, flocculation, secondary settling,
and segregation of wastewater streams.
The segregation of highly contaminated leaching, solvent-
extraction, precipitation, and scrubber waste streams from
noncontact cooling water and uncontaminated waste streams
is currently practiced and is essential to effective removal
of metals from the wastewater. Aeration may be necessary to
reduce COD to acceptable levels.
Rationale for Selection. The recommended BPCTCA is currently
in use within the ore mining and dressing industry. Control
and treatment technology within the subcategory (except at one
site leaching only concentrates) is inadequate at present.
This results in the discharge of appreciable quantities of
heavy metals removable by lime precipitation, and in excessive
suspended-solid loads. Sin'ce effluent streams are currently
very high in sulfates (10,000 mg/1), application of lime
8UBJKT ™SfAM«IS,cI«T«IJIS2,MMENDATIONS BASED UPON '"FORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-44
DRAFT
-------
DRAFT
TABLE IX-19. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-FERROALLOY-ORE MILLS
USING FLOTATION PROCESS
PARAMETER
PH
TSS
Oil and Grease
COD
Cyanide
As
Cd
Cu
Mn
Mo
Pta
Zn
Fluoride
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6" to 9'
20
3
50
0.05
0.4
0.05
0.05
1.0
t
0.1
0.1
10
24-hour maximum
6* to 9«
30
5
100
0.1
0.8
0.1
0.1
1.5
t
0.2
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
_
40
6
100
0.1
0.8
0.1
0.1
2.0
t
0.2
0.2
20
lb/1000
thort toni
_
80
12
200
0.2
1.6
0.2
0.2
4.0
t
0.4
0.4
40
24-hour maximum
kg/1000
metric toni
_
60
10
200
0.2
1.6
0.2
0.2
3.0
t
0.4
0.4
40
lb/1000
thort torn
_
120
20
400
0.4
3.2
0.4
0.4
6.0
t
0.8
0.8
80
•Value in pH units
*No limitations proposed for BPCTCA
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-45
DRAFT
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DRAFT
precipitation will result In marginal decreases (estimated
to be 10 to IS percent) In total dissolved solids, as well
as In substantial removal of heavy metals.
Levels of_ Effluent Reduction Attainable. The parameters
selected for control and the effluent reduction attainable by
Implementation of BPCTCA are presented In Table IX-20.
The limitation of ammonia, Cr, Mo, and V Is not recommended
using the BPCTCA. Control technology at BPCTCA Is not avail-
able. Hexavalent-chromium removal requires chemical reduc-
tion, which will require development work before application
to mill waste streams. Only trivalent chromium will be
removed by lime precipitation.
The application of air stripping for ammonia removal is not
considered practical without segregation of high-ammonia
process waste streams, which may require process modifica-
tion. Total dissolved solids, although a major waste con-
stitutent, are not limited because practical control tech-
nology applicable to these operations is not currently
available. Proper management of the discharge to ensure
rapid mixing and dispersal can alleviate possible problems
of stratification and formation of pockets of saline water
in the receiving waters.
Category; Mercury Ores
Subcategory: Mercury Mines
This subcategory Includes all mines, whether open-pit or
underground, operated for the extraction of mercury ores.
Identification of BPCTCA. The best practicable control tech-
nology currently available is use of chemical precipitation
in conjunction with settling Impoundments.
Chemical-precipitation methods for heavy-metal removal may
include lime- or sulflde-preclpitation methods. Mechanical
clarlfiers are an acceptable alternative method for suspended-
solid removal. Adjustment of the pH Co acceptable levels may
be necessary at some locations prior to discharge.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-4 6
DRAFT
-------
DRAFT
TABLE IX-20. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-FERROALLOY-ORE MILLS
USING LEACHING PROCESS
PARAMETER
PH
TSS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fluoride
CONCENTRATION (mg/8.)
IN EFFLUENT
30-day average
6" to 9"
20
3
50
t
0.5
0.05
t
0.05
1.0
t
0.1
t
0.1
10
24-hour maximum
6" to 9"
30
5
100
t
0.7
0.1
t
0.1
1.5
t
0.2
t
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
88
14
230
t
2.3
0.23
t
0.23
4.5
t
0.45
t
0.45
45
lb/1000
short tons
176
28
460
t
4.6
0.46
t
0.46
9.0
t
0.9
t
0.9
90
24-hour maximum
kg/1000
metric tons
132
23.3
460
t
3.22
0.46
t
0.46
6.8
t
0.9
t
0.9
90
lb/1000
short tons
264
46.6
920
t
6.44
0.92
t
0.92
13.6
t
1.8
t
1.8
180
•Value in pH units
*No limitations proposed for BPCTCA
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-47
DRAFT
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DRAFT
Rationale for Selection. The use of settling Impoundments
has been demonstrated to be effective In removal of suspended
solids at a large number of locations. Chemical-precipitation
methods are necessary to reduce heavy-metal levels because
present treatment at most^ locations, If any Is used, Is Inade-
quate. The use of lime-precipitation methods with effective
pH control Is a demonstrated and effective means of reducing
heavy-metal concentrations. The technology selected for
control of the pollutant parameters named will also have the
additional benefit of reducing other heavy metals as well.
Levels of Effluent Reduction Attainable. The parameters
selected and the levels of effluent reduction attainable are
presented in Table IX-21.
Subcategory; Mercury Mills or_ Mine/Mills Employing Gravity-
Separation Methods
This subcategory Includes those mills processing mercury ores
by gravity-separation methods. At present, there is one known
operation employing this method.
Identification £f_ BPCTCA. The best practicable control tech-
nology currently available is zero discharge by recycle of
process water or total impoundment. This practice can be
enhanced by diversion of runoff.
Rationale for Selection. The only operation using these
methods is currently attaining zero discharge by impoundment
and recycle of process water back to the process after tail-
ing-pond treatment. A secondary pond is maintained to Impound
overflow should unusual conditions prevail, and to collect
any seepage through the tailing impoundment. This water,
if any, is pumped back to the primary tailing pond.
Levels oif Effluent Reduction Attainable. Zero discharge
of pollutants will result from Implementation of BPCTCA.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANOB BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
IX-4 8
DRAFT
-------
DRAFT
TABLE IX-21. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-MERCURY MINES
PARAMETER
PH
TSS
Hg
Ni
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6" to 9»
20
0.0005
0.1
24-hour maximum
8Mo9»
30
0.001
0.2
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-49
DRAFT
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DRAFT
Subcategory; Mercury Mills or Mine/Mills Using Flotation
Process
This category Includes those operations beneficiating mercury
ores by the froth-flotation process. A single operation is
represented by this subcategory, with operation scheduled to
begin in 1975. Production is expected to reach 20,000 flasks
(695 metric tons, equivalent to 765 short tons), and this
operation will be the largest single producer of mercury in
the U.S.
Identification o_f_ BPCTCA. The best practicable control tech-
nology currently available for this subcategory is zero dis-
charge by the use of total recycle and complete impoundment
of process wastewater.
Rationale for Selection. The only known facility in this
subcategory is expected to be operational in 1975 and is
designed to attain zero discharge by recycle and impoundment
of process water. If the treatment system to be used is not
found adequate to handle the total wastewater volume, provi-
sions have already been made for construction to double the
present impoundment volume and take advantage of evapora-
tive losses.
Levels £f_ Effluent Reduction Attainable. The level of efflu-
ent reduction attainable by Implementation of BPCTCA is zero
discharge of wastewater to surface waters of the U.S.
Subcategory; Mills Recovering Mercury as _a_ Byproduct of
Base- or Precious-Metal Concentrates
This subcategory includes operations where mercury is obtained
as a byproduct of base- or precious-metal concentrates. The
recovery of mercury takes place at a refinery or smelters.
Identification of BPTCA. No separate limitations or tech-
nology are proposed. The waste treatment technology and
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
IX-50
DRAFT
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DRAFT
effluent limitations for the appropriate subcategory of base-
or precious-metal mills are applicable to this subcategory.
Category; Uranium. Radium, and Vanadium Ores
This category Includes mines and mills operated for the
extraction and/or concentration of uranium, radium, and
vanadium ores. Vanadium byproduct production from uranium
ores Is covered under this category, while primary vanadium
production Is covered, for purposes of this report, under
Ferroalloy Ores.
Subcategory; Uranium Mines
This subcategory Includes all uranium mines, whether open-pit
or underground, from which discharge of wastewater results.
Identification of_ BPCTCA. The best practicable control tech-
nology currently available for this subcategory Is the use
of settling ponds In conjunction with lime precipitation,
Ion exchange (for uranium removal), barium chloride coprecl-
pltatlon (for radium removal), and secondary settling.
The use of settling ponds Is almost universal In this sub-
category; however, frequently, the ponds used are small and
have Inadequate retention time. Where space limitations do
not permit use of such ponds, mechanical clarifler-floccula-
tors are acceptable alternatives for settling of suspended
solids.
Rationale for Selection. Nearly every uranium mine with
wastewater discharge currently practices suspended-solid
removal by the use of settling ponds. Treatment, as practiced,
Is currently uniformly inadequate to achieve acceptable levels
of pollutant control.
Currently, in addition to settling ponds, the best treatment
employed at uranium mines Includes the use of ion exchange
for removal of uranium from mine water. This has the dual
benefit of effluent treatment plus recovery of uranium values.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJBCT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-51
DRAFT
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DRAFT
This treatment has been economically applied for value recovery
at concentrations as low as 2 mg/1 of uranium.
Treatment, as currently practiced, Is judged to be Inadequate
for removal of either heavy metals or radium concentrations
In mine wastewater. The effectiveness of barium chloride
copreclpitation has been demonstrated at two mills In this
Industry category and can effectively reduce radium concentra-
tions to 1 picocurie per liter. Lime precipitation Is In use
at facilities in the ore mining and dressing Industry and has
been demonstrated to be effective for heavy-metal removal.
Secondary settling ponds may be necessary for removal of preci-
pitated solids.
Levels p_f Effluent Reduction Attainable. The parameters
selected for control and levels of effluent reduction attainable
by use of BPCTCA for this subcategory are presented in Table
IX-22. No limitations are proposed for TOC, Mo, and V reductions
using BPCTCA.
Subcategory; Mills Processing Uranium Ores by Acid or
Combined Acid/Alkaline Leaching
This subcategory includes operations which are operated using
the acid-leach or combined acid/alkaline-leach processes.
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory is zero
discharge by the use of impoundment and evaporation.
Lime neutralization of impounded water may be necessary to
prevent potential problems should seepage into ground water
take place.
\
Rationale for Selection. Approximately 90 percent of the
mills in this subcategory Impound and evaporate wastewater.
The remaining 10 percent are located in areas with light
precipitation and high evaporation and could practice impound-
ment. There are currently no uranium or uranium/vanadium
byproduct operations in wet or humid climates. Raw wastewaters
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND AHE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-52
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TABLE IX-22. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-URANIUM MINES
PARAMETER
PH
TSS
COD
TOC
As
<*
Cr
Cu
Mo
Ni
Pb
V
Zn
Ra226
Total U
CONCENTRATION (mg/K.)
IN EFFLUENT
30-day average
6* to 9»
20
100
t
0.5
0.05
0.05
0.05
t
0.1
0.1
t
0.1
!••
1
24-hour maximum
6«to9»
30
200
t
1.0
0.1
0.1
0.1
t
0.2
0.2
t
0.2
3"
3
•Value in pH units
TNo limitations proposed for BPCTCA
**Value in pieoeurias par liter
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-53
DRAFT
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DRAFT
from mills using acid leaching remain acid at the process dis-
charge, retain various heavy metals, and generally are not
suitable for recycling without additional or specialized
' treatment. Wastewaters from the alkaline-leach process are
normally recycled In part.
Levels of Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable by implementation of the above
technology.
Subcategory: Mills Processing Uranium Ores by_ Alkaline
Leaching
This subcategory includes those operations which are operated
using the alkaline-leach process only for extraction of
uranium, radium, and vanadium ores.
Identification of BPCTCA. The best practicable control
technology currently available for this subcategory is zero
discharge by the use of impoundment and recycle of mill process
wastewater.
Implementation of this technology requires the use of impound-
ment (for evaporation) and separation of effluent from the
purification or sodium-removal stages. This separated effluent
should be impounded and evaporated. The separation of this
wastewater from the mill process water facilitates recycle
from the tailing Impoundment by preventing the buildup of
sodium and sulfate, which would adversely affect the use of
recycled water.
Rationale for Selection. Currently, zero discharge by use of
this technology is attained at two of three alkaline-leach
mills. The alkaline-leach process lends itself to recycle.
In some instances, additional evaporation area may be necessary
during years with a less favorable water balance. All current
operations In this subcategory are located in arid areas.
In addition to the separate treatment of purification wastewater,
a fraction of wastewater from the recycle pond might have to be
bled off periodically to control the buildup of sodium and sulfate
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-54
DRAFT
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DRAFT
Ions in the recycle loop. In current practice, some bleeding
from the recycle pond occurs by seepage into the subsurface.
Levels of_ Effluent Reduction Attainable. Zero discharge is
attainable by implementation of the above technology.
Metal Ores. Not Elsewhere Classified
This group of metal-ore operations Includes mining and milling
of ores of antimony, beryllium, platinum, tin, titanium,
rate-earth metals, and zirconium.
Category; Antimony Ores
Subcategorv: Antimony-Ore Mines Alone
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory is chemical
precipitation (i.e., lime and sulfide precipitation) in
conjunction with removal of suspended solids by the use of
settling impoundments.
To implement the above technology, mechanical clarification
devices (e.g., clarifiers, clari-flocculators, etc.) may also
be used. Adjustment of pH by neutralizing agents may be
necessary at selected locations prior to discharge. Secondary
settling ponds may be necessary1 for removal of precipitated
solids.
Rationale for Selection. Chemical precipitation for removal
of heavy metals by lime addition is well-documented and has
been well-demonstrated in the ore mining and dressing industry.
Sulfide precipitation is the only effective and demonstrated,
economical method for removal of mercury and antimony to low
levels. The use of settling Impoundments is an almost-universal
treatment method for removal of suspended solids. Present
NOTICE: THESE ABE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
IX-55
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treatment methods in use in this subcategory consist of
settling alone. Heavy-metal discharges resulting from the
use of this treatment alone Indicate its uniform inadequacy.
Level of Effluent Reduction Attainable. The parameters
selected for control and effluent reduction attainable by use
of the above technology in this subcategory are presented in
Table IX-23.
Subcategory; Antimony Mills Using Flotation Process
Identification .of BPCTCA. . The best practicable control
technology currently available for this subcategory is zero
discharge by impoundment and/or recycle of process waste-
water.
To achieve zero discharge by recycling, additional, secondary
settling of process water may be necessary to reduce slime
content. Adequate impoundment area is necessary to achieve
zero discharge by impoundment.
Rationale for Selection. The only flotation mill operating
for primary-product recovery of antimony is currently achieving
zero discharge by impoundment. Recycle of process water, with
additional settling treatment for suspended-solid removal
should not present any technical difficulty.
Levels of? Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable by implementation of this
technology. ,
Subcategory; Mills Obtaining Antimony As £ Byproduct of_
Base- or Precious-Metal Milling Operation
This subcategory Includes operations where antimony is
recovered from a concentrate at a smelter or refinery (antimony
extraction plant).
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARC
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-5 6
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TABLE IX-23. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-ANTIMONY MINES
PARAMETER
pH
TSS
As
Fe
Sb
Zn
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6»to9«
20
0.07
1.0
0.5
0.1
24-hour maximum
6» to 9»
30
0.15
2.0
0.8
0.2
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-57
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Identifleation of BPCTCA. No separate limitations are proposed
for this subcategory. Limitations developed for the subcategory
of the primary metal recovered are recommended for this sub-
category.
Category; Beryllium Ores
Subcategory; Beryllium Mines
Identification of_ BPCTCA. The best practicable control
technology currently available is zero discharge by impoundment
of mine wastewater.
Rationale for Selection. The single operating mine in this
subcategory is achieving zero discharge by impoundment.
Levels of_ Effluent Reduction Attainable. Zero discharge of
mine wastewater is attainable by implementation of this tech-
nology .
Subcategory; Beryllium Mills
Ident i f icat ion of BPCTCA. The best practicable control
technology currently available is the total impoundment of
process wastewater.
Rationale for Selection. The above technology is currently
practiced at the single beryllium mill now operating.
Levels of_ Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable by implementation of this
technology.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND AhE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-58
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Category; Platinum Ores
This category represents facilities operated for the mining and
concentration of platinum ores by gravity-separation methods.
Most platinum in the U.S. is obtained as a byproduct of
smelting and refining of base or precious metals. A single
operating facility currently obtains platinum concentrates by
dredging and gravity separation for concentration of platinum
and a small amount (3 to A percent of concentrates) of byproduct
gold.
Identification of_ BPCTCA. The best practicable control
technology currently available is the use of settling ponds
for control of suspended-solid levels.
An alternative to implementation of this technology is the
pumping of wastewater back over tailings for sand and gravel
filtration, but a settling impoundment of some type will be
required for primary settling before discharge.
Rationale for Selection. The single operating facility of
this type currently employs settling ponds and filtration
through sands prior to discharge. Therefore, no additional
costs will be incurred.
Levels of_ Effluent Reduction Attainable. The parameters
chosen for control and the levels of effluent reduction attain-
able for this category are presented in Table IX-24.
Category: Rare-Earth Ores
Subcategory; Mines Operated For Obtaining Primary or
Byproduct Rare-Earth Ores
This subcategory is represented by one rare-earth mine,
which currently has no discharge of mine water.
Identification o_f BPCTCA. The best practicable control
technology currently available for this subcategory is zero
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-59
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TABLE IX-24. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-PLATINUM MILLS
PARAMETER
pH
TSS
Total Fe
CONCENTRATION (mg/fc)
IN EFFLUENT
30-day average
6» to 9«
30
1.0
24-hour maximum
6' to 9"
40
2.0
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
0.43
0.014
lb/1000
short tons
0.86
0.029
24-hour maximum
kg/1000
metric tons
0.58
0.029
lb/1000
short tons
1.16
0.058
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-60
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discharge by impoundment and/or reuse of mine water as process
water In a mill.
Rationale for Selection. Currently, no rare-earth-ore mines
exist which discharge wastewater. An operation located In the
arid region of the U.S. might practice total impoundment should
mine wastewater be encountered.
Levels of_ Effluent Reduction Attainable. Zero discharge of
pollutants can be attained should mine wastewater result.
Subcategory; Rare-Earth-Ore Mills Using Flotation and/or
Leaching Process
This subcategory includes a single operation extracting rare-
earth metals from rare-earth ores by means of a flotation and
leaching process.
Identification of_ BPCTCA. The best practicable control
technology currently available for this subcategory Is zero
discharge by separation of waste streams, followed by Impound-
ment and evaporation of leaching-process wastewater and recycle
of flotation-process water from a sedimentation Impoundment.
Rationale for Selection. The single operating facility in this
subcategory is currently practicing BPCTCA.
Levels of_ Effluent Reduction Attainable. Zero discharge of
process-water effluent is attainable by this technology.
Subcategory; Mills or Mine/Mills Obtaining Rare-Earth
Minerals By Gravity Methods
The rare-earth mineral monazite is currently recovered as
a byproduct of placer operations for titanium minerals.
BPCTCA for this subcategory is covered under the appropriate
titanium-ore subcategory. No separate or additional limit-
ations are proposed.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-61
DRAFT
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Category; Tin Ores
Currently, tin is primarily recovered at one location in the
U.S. as a byproduct of molybdenum mining and milling. A small
amount of tin is also produced at dredging operations for gold
as a byproduct of placer mining In Alaska. Recommended BPCTCA
and the levels of effluent reduction attainable are covered under
the appropriate ferroalloy-ore or gold-ore subcategory.
Although tin is recovered by placer and gravity methods as well
as by magnetic and electrostatic separation or extraction, no
major deposits are currently exploited in the U.S.
Cagegory; Titanium Ores
Subcategory; Mines Obtaining Titanium Ore by_ Lode Mining
Currently in the U.S., there is one operation mining a
titanium-ore deposit by open-pit methods.
Identification of BPCTCA. The best practicable control
technology currently available is neutralization in
conjunction with the use of a settling pond for suspended-
solid removal.
Rationale for Selection. Current practice in the single
operating facility is Impoundment and discharge of mine waste-
water. Retention time for this small settling pond is short,
and treatment for suspended solids in the discharge water is
inadequate. Expansion of the settling pond to allow Increased
retention time is necessary. Neutralization of mine waters
is necessary to maintain pH values at levels which will prevent
solubillzation of heavy metals.
Levels of Effluent Reduction Attainable. The parameters
selected and levels of effluent reduction attainable by use of
the above technology are presented in Table IX-25.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-62
DRAFT
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TABLE IX-25. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-TITANIUM MINES
PARAMETER
PH
TSS
Oil and Grease
Fe
Fluoride
CONCENTRATION (mg/&)
IN EFFLUENT
30-day average
6* to 9*
20
1.0
1.0
1.0
24-hour maximum
6» to 9»
30
1.6
2.0
2.0
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-63
DRAFT
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Subcategory; Titanium Mills or Mine/Mills Using Electrostatic
and/or Magnetic plus Gravity and/or Flotation Methods
This subcategory is currently represented by one milling
operation, which concentrates ilmenite from an ilmenite/
magnetite ore.
Identification of BPCTCA. The best practicable technology
currently available is the use of tailing ponds with lime
precipitation in conjunction with diversion ditching and recycle
of process water.
Rationale for Selection. Currently, the one operating mill
in this subcategory is practicing Impoundment and recycle
during approximately ten months of the year. Diversion
ditching should reduce the volume of water to be treated.
Lime precipitation is well-documented and has been well-
demonstrated in other segments of the ore mining and dressing
industry, and its use is necessary to reduce heavy-metal
concentrations in discharge water.
Levels of Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable by use of the above technology are presented in
Table IX-26.
Subcategory; Titanium-Ore Mills Using Physical Milling
Methods in Conjunction With Dredge Mining
This subcategory includes operations engaged in the dredge
mining of placer deposits of sands containing rutile, ilmenite,
and leucoxene. Monazite, zircon, and other heavy minerals are
also obtained as byproducts from these operations. Milling
techniques employed in this subcategory include the use of
wet gravity methods in conjunction with electrostatic and/or
magnetic methods.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANQE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-64
DRAFT
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TABLE IX-26. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-TITANIUM MILLS
PARAMETER
pH
TSS
Cr
Fe
Ni
Zn
Fluoride
CONCENTRATION (mg/a)
IN EFFLUENT
30-day average
6» to 9'
20
0.1
1.0
0.1
0.1
1.0
24-hour maximum
6» to 9*
30
0.2
2.0
0.2
0.2
2.0
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric toni
—
160
0.8
8
0.8
0.8
8
lb/1000
short tons
-
320
1.6
16
1.6
1.6
16
24-hour maximum
kg/1000
metric tons
-
240
1.6
16
1.6
1.6
16
lb/1000
short tons
-
480
3.2
32
3.2
3.2
32
•Value in pH units
IX-65
DRAFT
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Identification of BPCTCA. The best practicable control
technology currently available for this category is settling
impoundment with maintenance of a pH of 3.5, secondary settling,
and neutralization.
Current practice of this technology normally involves the use
of three sedimentation ponds. The first pond is maintained at
acid pH (3.5) for control of organic matter. Secondary settling
is practiced at the second .pond, with a third "polishing pond"
being used for final clarification and neutralization by lime
addition.
Rationale for Selection. Three operations are currently
practicing this technology, and it has been demonstrated effective
for reduction of TOC and COD resulting from humic materials
present in the process wastewater. Suspended-solid levels are
maintained at low values due to the use of three settling ponds.
Levels of_ Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable by use of the above technology are presented in
Table IX-27.
Category; Zirconium Ores
Zircon is produced as a byproduct of titanium placer operations.
Mining and milling methods are inseparable from those used in
titanium dredge mining and wet milling. As a result, no separate
technology or limitations are proposed for zirconium ores.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-66
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TABLE IX-27. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BPCTCA-TITANIUM DREDGE MINE
WITH WET SEPARATION MILL
PARAMETER
pH
TSS
COD
TOC
Oil and Graaie
Fa
CONCENTRATION (mg/fc)
IN EFFLUENT
30-day average
6f to 9f
10
15
8
1.0
1.0-
24-hour maximum
6* to 91
20
18
11
1.5
2.0
WASTE LOAD per unit concentrate produced*
30-day average
kg/
metric ton
_
0.43
0.65
0.34
0.043
0.043
Ib/
short ton
_
0.86
1.30
0.68
0.086
0.086
24-hour maximum
kg/
metric ton
_
0.86
0.77
0.47
0.064
0.086
Ib/
short ton
_
1.72
1.55
0.95
0.128
0.172
•Ore-production figures unavailable
f Value in pH units
NOTICE; THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
IX-67
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SECTION X
BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE,
GUIDELINES AND LIMITATIONS
INTRODUCTION
The effluent limitations which must be achieved by July 1,
1983 are based on the degree of effluent reduction attainable
through the application of the best available technology
economically achievable (BATEA). For the ore mining and
dressing industry, this level of technology was based on the
very best control and treatment technology employed by a
specific point source within each of the industry's sub-
categories, or which is readily transferable from one industry
process to another. In Section IV, the ore mining and dressing
industry has been divided into ten major categories. Several
of those major categories have been further subcategorized, and,
for reasons explained in Section IV, each subcategory will be
treated separately for the recommendation of effluent limitations
guidelines and standards of performance.
The following factors were taken into consideration in deter-
mining the best available technology economically achievable:
(1) age of equipment and facilities involved;
(2) process employed;
(3) engineering aspects of the application of various
types of control techniques;
(4) process changes;
(5) cost of achieving the effluent reduction resulting
from application of BATEA; and
(6) nonwater-quality environmental impact (including
energy requirements).
In contrast to the best practicable control technology currently
available, best available technology economically achievable
assesses the availability in all cases of in-process controls
as well as control or additional treatment techniques employed
at the end of a production process. In-process control
options available which were considered in establishing these
control and treatment technologies include:
NOTICB: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-l
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(1) alternative water uses
(2) water conservation
(3) waste-stream segregation
(4) water reuse
(5) cascading water uses
(6) byproduct recovery
(7) reuse of wastewater constituents
(8) waste treatment
(9) good housekeeping
(10) Preventive maintenance
(11) quality control (raw material, product, and effluent)
(12) monitoring and alarm systems.
Those plant processes and control technologies which, at the
pilot plant, semi-works, or other level, have demonstrated
both technological performances and economic viability at a
level sufficient to reasonably justify investing in such
facilities were also considered in assessing the best avail-
able technology economically achievable. Although economic
factors are considered in this development, the costs for
this level of control are intended to be for the top-of-the-
line of current technology subject to limitations Imposed by
economic and engineering feasibility. However, this technology
may necessitate some industrially sponsored development work
prior to its application.
Based upon the information contained in Sections III through
IX of this report, the following determinations were made on
the degree of effluent reduction attainable with the appli-
cation of the best available technology economically achievable
in the various categories and subcategories of the ore mining
and dressing industry.
GENERAL WATER GUIDELINES
Process Water
Process water is defined as any water contacting the ore,
processing chemicals, intermediate products, byproducts, or
products of a process, including contact cooling water. All
process-water effluents are limited to the pH range of 6.0
to 9.0 unless otherwise specified.
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE M8ED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-2
DRAFT
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DRAFT
Cooling Water
In the ore mining and dressing industry, cooling and process
waters are sometimes mixed prior to treatment and discharge.
In other situations, cooling water is discharged separately.
Based on the application of best available technology econo-
mically achievable, the recommendations for the discharge of
such cooling water are:
An allowed discharge of all non-contact cooling waters
provided that these conditions are met:
(1) Thermal pollution be in accordance with standards
to be set by EPA polcies. Excessive thermal rise
in once-through, non-contact cooling water in the
ore mining and dressing Industry has not been a
significant problem.
(2) All non-contact cooling waters be monitored to
detect leaks of pollutants from the process.
Provisions should be made for treatment to the
standards established for the process-wastewater
discharges prior to release in the event of such
leaks.
(3) No untreated process waters be added to the cooling
waters prior to discharge.
The above non-contact cooling-water recommendations should be
considered as interim, since this type of water plus blowdown
for water treatment, boilers, and cooling towers will be
regulated by EPA at a later date as a separate category.
Storm-Water Runoff
Storm-water runoff may present pollution-control problems for
certain subcategorles. This is true where large stockpiles
of ore to be processed or where waste materials are piled in
areas where the topography may result in runoff to a nearby
stream. Other particular problems include the size and location
of tailing Impoundments. Where such impoundments occur, their
design, construction, and maintenance should be sufficient to
contain the precipitation from the 25-year, 24-hour rainfall
event as established by the U.S. National Weather Service for
the area in which such impoundment is located. That volume of
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-3
DRAFT
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DRAFT
process wastewater which is equivalent to the volume of preci-
pitation that falls within the impoundment ill excess of that
attributable to the 25-year, 24-hour rainfall event may be
discharged, when such an event occurs.
To preclude unfavorable water-balance conditons resulting
from precipitation and runoff in connnection with tailing
impoundments, diversion ditching to prevent natural drainage
or runoff from mingling with process wastewater should be
constructed.
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPOF, i AND ARE
SUBJECT TOCHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-4
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BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE. BY ORE
CATEGORY AND SUBCATEGORY
Category; Iron Ores
Subcategory; Iron-Ore Mines
Identification of_ BATEA. The best available technology econo-
mically achievable for the wastewater resulting from the mining
of iron ore is the use of settling ponds with coagulation/
flocculation systems in conjunction with chemical precipitation
by lime to a pH of 8.5 to 9.
To implement the above technology, secondary settling may be
required for removal of precipitated solids.
Rationale for Selection. The use of lime neutralization and
precipitation has been well-demonstrated in the ore mining and
dressing industry, as well as in the coal mining industry,
where it is used for control of acid mine drainage and for
precipitation of metals. Application of this technology in the
bauxite mining industry has been well-documented, both on a
full-scale basis and on a pilot scale.
Levels of_ Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable by the use of this technology are presented in
Table X-l.
Subcategory; Iron-Ore Mills Employing Physical and/or
Chemical Separation
Identification of BATEA. The best available technology
economically achievable for the treatment of wastewater
resulting from milling processes used in this Subcategory is
the use of tailing impoundments with coagulation/flocculation
systems in conjunction with chemical precipitation by lime
addition to a pH of 8.5 to 9.
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANQE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-5
DRAFT
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DRAFT
TABLE X-1. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-IRON-ORE MINES
PARAMETER
PH
TSS
Oil and Grease
Dissolved Fa
Total Mn
CONCENTRATION (mg/H)
30-day average
6* to 9*
20
10
0.5
2.0
daily maximum
6" to 9*
30
10
1.0
4.0
•Value in pH units
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPOR1 *ND ARL
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-6
DRAFT
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DRAFT
To implement the above technology, secondary settling ponds
may be required for removal of precipitated solids. Treatment
requirements can be substantially reduced by partial recycling
of process water, a practice which has widespread use in this
subcategory.
Rationale for Selection. The use pf lime neutralization and
precipitation has been well-demonstrated in the ore mining and
dressing industry, as well as in the coal mining and bauxite
mining industries, where it has been used extensively for
control of acid mine drainage and heavy-metal removal.
Levels of Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable by application of BATEA are presented in Table X-2.
Subcategory; Iron-Ore Mills Employing Magnetic and Physical
Separation (Mesabi Range)
Identification of BATEA. The best available technology
economically achievable for this subcategory is zero discharge
of process wastewater. (Same as BPCTCA.)
Category; Copper Ores
Subcategory; Copper-Ore Mines
Identification of_ BATEA. The best available technology
economically achievable for this subcategory is the use of
lime precipitation and settling, with pH adjustment prior to
discharge, if necessary. (Same as BPCTCA.)
Subcategory; Copper-Ore Mines Employing Hydrometallurgical
Processes
Identification of_ BATEA. The best available technology
economically achievable is zero discharge of process or mine
wastewater. (Same as BPCTCA.)
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-7
DRAFT
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DRAFT
TABLE X-2. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATE A-I RON-ORE MILLS EMPLOYING
PHYSICAL METHODS AND/OR CHEMICAL REAGENTS
PARAMETER
PH
TSS
Oil and Grease
Diuolved Fe
Total Mn
CONCENTRATION (mg/A)
IN EFFLUENT
30-day average
6Mo9»
20
10
0.6
2.0
24-hour maximum
6Mo9»
30
10
1.0
4.0
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
-
26.2
13.2
0.66
2.70
lb/1000
short tons
—
52.4
26.4
1.32
5.40
24-hour maximum
kg/1000
metric tons
—
39
13.2
1.32
5.4
lb/1000
short tons
—
78
26.4
2.64
10.8
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-8
DRAFT
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DRAFT
Subcategory; Copper Mills Employing Vat-Leaching Process
-o
Identification of BATEA. The best available technology
economically achievable is zero discharge of process waste-
water. (Same as BPCTCA.)
Subcategory! Copper Mills Employing Froth Flotation (Areas
Where Net Evaporation Equals or Exceeds 76.2 centimeters
(30 inches) per Year)
Identification of_ BATEA. The best available technology
economically achievable is zero discharge«of process waters
through use of impoundment and total recycle. (Same as BPCTCA.)
Subcategory: Copper Mills Employing Froth Flotation (Areas
Where Net Evaporation is Less Than 76.2 centimeters (30 inches)
per Year)
5
Identification of BATEA. The best available technology
economically achievable for this Subcategory is zero discharge
of process wastewater through the reuse, recycle, and evapora-
tion of all process waters.
Rationale for Selection. The procedures which can be employed
at flotation mills in this Subcategory for recycling are presently
being demonstrated in the copper milling industry. Because of
the location of mills of this Subcategory in areas receiving
greater rainfall than mills in the previous Subcategory, it may
be necessary to employ one or more methods of control and
treatment:
Diversion of_ Runoff; Water which may enter tailing
impoundments by runoff or drainage must be prevented
from entering by the use of diversion ditching and
conveyance around the Impoundment.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-9
DRAFT
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DRAFT
Segregation of Wastewater; Water conveyed to a mill
treatment system from mine pumpout may result in excess
water and, thus, a discharge. 'Where this occurs, separate
treatment of mine water will be necessary to reduce the
amount of water to be impounded and to improve the water
balance for a recycle system. Evaporation ponds for a
portion of wastewater may be employed seasonally to reduce
wastewater volume.
Recycle of Process Water; Process water should be recycled
from impoundments. Makeup water can be added, when
necessary, to maintain the needed volume of process water.
Tailings-Pond Seepager Seepage, where it occurs, should
be diverted to a ditch and pumped back into the tailing
pond.
Current operations in this subcategory employ partial
recycle of process water. Application of methods for
reduction of wastewater flow, and recycle of process
water, will enable the zero-discharge limitation to be
met.
Levels of_ Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable by the implementation of
this technology.
Category; Lead and Zinc Ores
Subcategory; Lead and/or Zinc Mines Having No Solubilization
Potential
Identification of_ BATEA. The best available technology
economically achievable for this subcategory is the same as
BPCTCA.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND AHfc
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-10
DRAFT
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DRAFT
There is currently no evidence that the quality of effluent
from mines in this subcategory can be significantly improved
by resort to treatments such as ion exchange, reverse
osmosis, or other advanced processes. These processes still
generally produce concentrated waste streams requiring
further treatment and/or disposal, and it appears that no
significant benefit would be obtained. Therefore, no change
from BPCTCA is recommended for this subcategory.
Subcategory; Lead and/or Zinc Mines Having Solubilization
Potential
Identificat ion of_ BATEA. The best available technology
economically achievable for this subcategory is use of settling
Impoundment(s) with lime precipitation.
Secondary settling ponds may be needed at some facilities for
removal of precipitated solids.
Rationale for Selection. The success of several operations
which would be classed in this subcategory in attaining effluent
quality which approaches that in the lead/zinc-mine subcategory
without solubilization problems suggests that, with proper pH
control and/or chemical (lime) precipitation, improved effluent
quality can be obtained by all mines in this subcategory.
The limits recommended here are based on the results reported
for a Canadian mining operation, as well as on information
from an exemplary mine discussed in Section VII.
Levels of_ Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable are presented in Table X-3.
Subcategory; Lead and/or Zinc Mills
Identification of_ BATEA. The best available technology
economically achievable is zero discharge through total
recycle and impoundment of process water.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-ll
DRAFT
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DRAFT
TABLE X-3. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-LEAD AND ZINC MINES
HAVING SOLUBILIZATION POTENTIAL
PARAMETER
PH
TSS
COD
Oil and Create
Cd
Cr
Cu
Fe
Hg
Mn
Pb
Zn
CONCENTRATION (mg/8,)
IN EFFLUENT
30-day average
6" to 9*
20
50
10
0.01
0.1
0.05
1.0
0.0005
0.5
0.1
0.1
24-hour maximum
6« to 9*
30
100
10
0.02
0.2
0.1
2.0
0.001
1.0
0.2
0.2
*Value in pH units
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPOR T AND ARE
SUBJECT TOCHANQE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-12
DRAFT
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DRAFT
To implement this technology, prevention of runoff entry to
tailing impoundments by diversion may be necessary. Segrega-
tion and treatment of mine water separately from process water
may be necessary at some locations because of an excess water
balance adversely affecting the ability to impound.
Rationale for Selection. The fact that several lead/zinc
and copper sulfide ore mills do operate in a total-recycle
mode suggests that zero discharge is an attainable mode of
operation for all such mills. The technological feasibility
of recycle at lead/zinc/copper (sulfide-mineral) mills has
been demonstrated and, with adequate development work, should
be applicable to all mill operations. In some cases, engin-
eering modifications—and, perhaps alternative modes of
solids disposal and retention—would appear to provide feasible
solutions to water-balance problems. For example, dewatering
of tailings in a clarifier with recirculation of the overflow
may be necessary where precipitation presently creates
difficulty for total recycle and impoundment.
Levels p_f_ Effluent Reduction Attainable. Zero discharge of
effluent will result from implementation of BATEA.
Category; Gold Ores
Subcategory: Gold Mines (Alone)
Identification of_ BATEA. The best available technology
economically achievable is zero discharge of process waste-
water by impoundment and/or recycle. (Same as BPCTCA.)
Subcategory; Gold Mines or_ Mine/Mills Employing Amalgamation
Identification o_f_ BATEA. The best available technology
economically achievable is zero discharge of process water
by a process change to cyanidation extraction, settling-
pond treatment, and recycle of decant water.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-13
DRAFT
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DRAFT
To Implement this technology, a higher degree of control over
the quality of the reclaimed water can be maintained If the
tailing-pond decant Is collected In a secondary or polishing
pond prior to recycle back to the mill circuit.
Rationale for Selection. The BATEA identified for this
subcategory has demonstrated application and reliability in
the gold milling Industry. Total recycle of tailing-pond
decant is currently practiced by one mill. Total-recycle.
systems are also being employed in several other milling
subcategories. The change in process from amalgamation to
cyanldation will entail engineering modifications. The
feasibility of this process change is demonstrated by the
recent change of a gold mill from amalgamation to cyanidation.
Levels of Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable by Implementation of this
technology.
Subcategory; Mills or Mine/Mills Beneficiating Gold Ores
(Froth-Flotation Process)
Identification of_ BATEA. The best available technology
economically achievable for this subcategory is zero discharge
by impoundment, total recycle of process water, and diversion
ditching to reduce impounded water volume.
The recommended technology is essentially the same as BPCTCA
except that engineering modifications of the process-water
system and control of runoff volume are designed for total
recycle and Impoundment.
Rationale for Selection. The single operating facility in
this subcategory currently is achieving zero discharge, nine
to ten months of the year, by prevention of runoff entry into
tailing impoundments, Increased impoundment volume, and total
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANOE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-14
DRAFT
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DRAFT
recycle of process water. Optimization of the existing system
by minor modifications and engineering changes should enable
attainment of zero discharge.
Levels of Effluent Reduction Attainable. Zero discharge of
process wastewater is attainable by implementation of this
technology.
Subcategory; Gold Mills or Mine/Mills Employing Gravity Separation
Identification p_f_ BATEA. The best available technology
economically achievable is the use of settling or tailing
impoundments. (Same as BPCTCA.)
Subcategory; Mill Operations Where Gold ijs Recovered as
Byproduct £f_ Base-Metal Milling Operation
Identification o_f BATEA. No separate limitations are recom-
mended for this Subcategory. The BATEA for this Subcategory
is the same as BATEA for the primary metal recovered.
Category; Silver Ores
Subcategory; Silver Mines (Alone)
Identification £f BATEA. The best available technology
economically achievable is chemical (lime or sulfide)
precipitation with settling ponds. (Same as BPCTCA.)
Subcategory; Silver Mills £r_ Mine/Mills Employing Froth
Flotation
Identification of_ BATEA. The best available technology
economically achievable is zero discharge by use of total
recycle of process water and/or total impoundment. (Same
as BPCTCA.)
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-15
DRAFT
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DRAFT
Subcategory; Silver Mills or Mine/Mills Using Cyanidation
Identification of_ BATEA. The best available technology
economically achievable is attainment of zero discharge by
total recycle and/or total impoundment of process wastewater.
(Same as BPCTCA.)
Subcategory; Silver Mills or Mine/Mills Using Amalgamation
Identification of_ BATEA. The best available technology
economically achievable for this Subcategory is the attainment
of zero discharge by a process change to cyanidation and total
recycle and/or total impoundment of process wastewater.
In order to achieve total recycle, a higher degree of control
over the quality of the reclaim water can be maintained if
the tailing-pond decant is collected in a secondary settling,
or polishing, pond prior to recycle back to the mill circuit.
The secondary pond will serve as the surge pond in the recycle
system.
Rationale for Selection. The recommended technology has been
demonstrated as feasible in both the gold and silver milling
industries. Recycle systems are also being employed in the
copper, lead, and zinc milling industries. Process modification
from amalgamation to cyanidation has been technically accomplished
in the gold milling industry with no apparent loss of recovery
and with elimination of high mercury levels in the discharge.
Levels of Effluent Reduction Attainable. No discharge of
process wastewater is attainable by implementation of the above
technology.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPOH I AND >\flE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-16
DRAFT
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DRAFT
Subcategory; Mills Recovering Silver by_ Gravity Separation
Methods
Identification of BATEA. The best available technology
economically achievable is the use of settling impoundment.
(Same as BPCTCA.)
Snbcategory; Mill Operations Where Silver is^ Recovered £s.
Byproduct of_ Base-Metal Milling Operation
Identification of_ BATEA. No separate limitations are
recommended for this subcategory. The BATEA for this subcategory
is the same as BATEA for the primary metal recovered.
Category; Bauxite Ores
Identification of_ BATEA. The best available technology
economically achievable is the same as BPCTCA for this subcategory
(lime precipitation and settling) with optimized pH control and
operating efficiencies.
Rationale for Selection. The recommended treatment is currently
being operated at one bauxite operation with no technical
difficulties. Although relatively low flow conditions prevail,
a large-scale treatment plant is currently under construction
and is expected to be operational in mid-1975.
Levels of Effluent Reduction Attainable. The parameters
selected and effluent limitations attainable by implementation
of this technology are presented in Table X-4.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-17
DRAFT
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DRAFT
TABLE X-4. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR ALKALINE MINE DRAINAGE
BATEA-BAUXITE MINES (ACID OR ALKALINE MINE
DRAINAGE)
PARAMETER
PH
TSS
Al
Fe
Mn
Ni
Zn
CONCENTRATION (mg/S.)
IN EFFLUENT
30-day average
6» to 9»
20
0.3
0.30
0.5
0.05
0.05
24-hour maximum
6Mo9"
30
0.6
0.60
1.0
0.1
0.10
•Value in pH units
kin-Tire THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
MJBCT TO CHANGE MXD UPOlI^COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-18
DRAFT
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DRAFT
Category; Ferroalloy Ores
Subcategory; Mines Extracting Ores of Ferroalloy Metals
Identification of_ BATEA. The best available technology
economically achievable is the lime precipitation in conjunction
with a settling pond (BFCTCA) and the use of flocculants and
secondary settling.
In selected instances, the use of coprecipitatlon by
ferric sulfate, or ion exchange, for removal of molyb-
denum and vanadium may be necessary. An alternative method
for suspended-solid removal is the use of a mechanical
clari-flocculator.
Rationale for Selection. The use of chemical flocculants and
secondary settling is a common practice in the ore mining and
dressing industry and has been demonstrated effective. The
limitations on molybdenum and vanadium are met at existing
mines by the practice of sound water management within the
mine (preventing contact with finely divided ore). The removal
of molybdenum and vanadium by coprecipitation or ion exchange
is currently being practiced at a pilot plant and on the
laboratory scale.
Levels £f_ Effluent Reduction Attainable. The parameters
selected and levels of effluent reduction attainable are
presented in Table X-5.
Subcategory; Mills or_ Mine/Mills Processing Less Than 5.OOP
Metric Tons (5.512 Short Tons) per Year o_f Ferroalloy Ores By_
Methods Other Than Ore Leaching
Identification of_ BATEA. The best available technology
economically achievable is the addition of flocculation and
runoff diversion to BPCTCA (neutralization and settling).
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.
X-19
DRAFT
-------
DRAFT
TABLE X-S. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-FERROALLOY-ORE MINES
PARAMETER
pH
TSS
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
V
Zn
Fluoride
CONCENTRATION (mg/K.)
IN EFFLUENT
30-day average
6» to 9"
20
3
0.4
0.5
0.03
0.7
0.5
0.10
1.0
0.1
10
24-hour maximum
6"to9»
30
5
0.8
0.1
0.06
1.0
1.0
0.15
2.0
0.2
20
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-20
DRAFT
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DRAFT
Rationale for Selection. Flocculation and diversion are
widely known, demonstrated, and practiced within this
Industry.
Levels of Effluent Reduction Attainable. The parameters
selected for control and effluent reduction attainable are
presented in Table X-6.
Subcategory: Mills Processing More Than 5,000 Metric Tons
(5,512 Short Tons) £f_ Ferroalloy Ores per Year By_ Physical
Methods
Identifi cat ion £f BATEA. The best available technology
economically achievable is the addition of total process water
recycle to BPCTCA (partial recycle, lime precipitation, tailing
pond, flocculation, and secondary settling).
Rationale for Selection. There are no technical obstacles to
process-water recycle at these operations. Effective suspended-
solid removal precludes deleterious effects from circulating
slimes on recovery. At certain locations, total recycle with
zero discharge might be employed, eliminating the need for
lime precipitation.
Levels £f Effluent Reduction Attainable. The parameters
selected and effluent reduction attainable by Implementation
of this technology are presented in Table X-7.
Subcategory; Mills Processing More Than 5.000 Metric Tons
(5,512 Short Tons) o.f Ferroalloy Ores per Year By_ Flotation
Methods
Ident i ficat ion o£ BATEA. The best available technology
economically achievable is the addition of process-water
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-21
DRAFT
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DRAFT
TABLE X-6. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-FERROALLOY-ORE MILLS
TREATING LESS THAN 5,000 METRIC TONS (5,512
SHORT TONS) PER YEAR
PARAMETER
PH
TSS
CONCENTRATION (mg/4)
IN EFFLUENT
30-day average
6* to 9"
25
24-hour maximum
6' to 9"
35
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
75
lb/1000
short tons
150
24-hour maximum
kg/1000
metric tons
105
lb/1000
short tons
210
•Value in pH units
TABLE X-7. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-FERROALLOY-ORE MILLS
TREATING MORE THAN 5,000 METRIC TONS
(5,512 SHORT TONS) PER YEAR BY PHYSICAL PROCESSING
PARAMETER
PH
TSS
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
CONCENTRATION (mg/e)
IN EFFLUENT
30-day average
6»to9»
20
3
0.4
0.05
0.03
0.7
0.6
0.10
0.1
10
24-hour maximum
6» to 9"
30
5
0.8
0.1
0.06
1.0
1.0
0.2
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
_
6
0.9
0.1
0.02
0.01
0.2
0.15
0.03
0.03
3
lb/1000
short tons
_
12
1.8
0.2
0.04
0.02
0.4
0.30
0.06
0.06
6
24-hour maximum
kg/1000
metric tons
_
9
1.5
0.2
0.04
0.02
0.3
0.3
0.06
0.06
6
lb/1000
short tons
18
3.0
0.4
0.08
0.04
0.6
0.6
0.12
0.12
12
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPOR1 AND AHb
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-22
DRAFT
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DRAFT
recycle, oxidation (aeration, chlorination, or ozonation),
and coprecipitation or ion exchange.
Rationale for Selection. The use of recycle to reduce the
volume of water discharged, and the employment of treatment
processes aimed specifically at the removal of COD, cyanide,
and molybdenum, will effect substantial reduction in total
pollutant load discharged from operations in this subcategory.
Treatment technology is drawn from pilot-plant studies and
examples of waste treatment in other Industries, as well as
from other segments of the ore mining and milling industry.
In some cases, substantial process development and optimiza-
tion effort will be required for the successful application
of selected treatment technology in the ferroalloy-ore mining
and milling Industry.
As discussed in Section IX, recycle can be difficult to
apply successfully in flotation operations—particularly, in
fatty-acid floats. Nonetheless, the industry affords numerous
examples of operations successfully practicing a high degree
of water reuse. Although simple sulfide-float circuits are
found to be most compatible with recycle, examples of recycle
may be cited even in plants with complex fatty-acid flotation
circuits. Auxiliary techniques such as aeration may be required
to limit problems with recirculating reagent, and, since
some floats are found to be sensitive to Inorganic salts in
the water, a certain amount of bleed from some float circuits
is expected to be necessary. For some flotation circuits,
extensive development is expected to be required to achieve
stable operation with recycled water. Based on what has been
achieved in the industry to date, discharge of 25 percent or less
of process-water volume can be achieved. Zero discharge may
be attained by use of total recycle of process water and/or by
impoundment, at selected sites.
The oxidation of cyanide ion to cyanate (and, ultimately, carbon
dioxide and nitrate) and aeration for the reduction of COD
are standard treatment practices in a variety of other industries
which are applicable to flotation-mill effluents. Since raw
waste values of both cyanide and COD are relatively low, a
simple aeration or ozonation or chlorination treatment will
be effective. Such treatment must, of course, follow removal
of partlculates and oxldlzablc species, such as metal sulfides,
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-23
DRAFT
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DRAFT
from the waste stream. Data for existing operations indicate
that, for many sites, this treatment may be rendered unnecessary
by proper reagent control and oxidization incidental to
other treatment.
Two techniques for the removal of molybdenum from solution
which are currently in the pilot-plant stage hold promise for
large-scale application and provide the basis for 1983 effluent
limitations. Coprecipitation with ferric hydroxide by ferric
sulfate addition, and ion exchange, both have been shown to be
viable, although not presently optimized, techniques. A
considerable history of unintentional collection (and subsequent
rejection) of molybdenum in ion-exchange uranium-recovery
operations provides background for the application of that
technique. Coprecipitation has been studied extensively as
part of an examination of the potential pollutions associated
with molybdenum.
Levels o_f Effluent Reduction Attainable. The parameters
selected and effluent reduction attainable by implementation
of the above technology are presented in Table X-8.
Subcategory: Mills Processing Ferroalloy Ores By_ Leaching
Techniques
Identification p_f BATEA. The best available technology econo-
mically achievable is the addition of chromium reduction and
air stripping to BPCTCA (lime precipitation, primary and secondary
settling, flocculation, and wastewater segregation).
Rationale for Selection. Segregation of waste streams from
solvent-extractlon/precipitation circuits is currently practiced
at one site in the ferroalloy milling industry where concen-
trates are leached. This allows treatment of the segregated
waste stream for TDS removal by evaporation and crystallization,
and for removal of ammonia in an air stripper. Similar waste
segregation and ammonia removal is under development for a
plant in the ferroalloy subcategory practicing ore leaching.
The use of sulfur dioxide for reduction of hexavalent chromium
to trlvalent forms, with subsequent precipitation of the
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-24
DRAFT
-------
DRAFT
TABLE X-8. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-FERROALLOY-ORE MILLS
USING FLOTATION PROCESS
PARAMETER
PH
TSS
Oil and Grease
COD
Cyanide
At
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
CONCENTRATION Img/H)
IN EFFLUENT
30-day average
6* to 9*
20
3
25
0.02
0.4
0.05
0.03
0.7
0.5
0.1
0.1
10
24-hour maximum
6" to 9»
30
5
50
0.04
0.8
0.1
0.06
1.0
1.0
0.2
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric toni
10
1.5
13
0.01
0.2
0.03
0.02
0.4
0.3
0.05
0.05
5
lb/1000
short tons
20
3.0
25
0.02
0.4
0.05
0.03
0.7
0.5
0.1
0.1
10
24-hour maximum
kg/1000
metric tons
16
2.5
25
0.02
0.4
0.05
0.03
0.6
0.5
0.1
0.1
10
lb/1000
short tons
30
5.0
50
0.04
0.8
0.1
0.06
1.0
1.0
0.2
0.2
20
•Value In pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-25
DRAFT
-------
DRAFT
hydroxide, Is a standard waste-treatment practice in many
industries. Application to milling wastes will require process
optimization for lower Initial chromium concentrations but
does not present any insurmountable problems.
Other treatment techniques which may be used on thpsr waste
streams have been discussed under previous subcatcgnrles and
pose no special problems in treating leaching-mll J waste
water. (The copreclpitation and ion-exchange techniques
described for molybdenum may also be applied to vanadium.)
The feasibility of process-water recycle will be highly
variable, depending on the details of specific operations,
amount of soluble material in the ore, leaching reagents,
eluents, preclpitants, etc. Zero discharge may be achieved
at specific sites.
Levels ojE Effluent Reduction Attainable . The parameters
selected and effluent reduction attainable for this subcategory
are presented in Table X-9.
Category; Mercury Ores
Subcategory; Mercury Mines
Identification of BATEA. The best available technology
economically achievable is the use of chemical (Jime or sulfide)
precipitation and settling Impoundments. (Same as BPCTCA.)
Subcategory; Mercury Mills or Mine/Mills Employing Gravity
Separation
Identification £f_ BATEA. The best available technology
economically achievable is zero discharge by recycle of process
water and/or total impoundment. (Same as BPCTCA.)
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMA DON IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-26
DRAFT
-------
DRAFT
TABLE X-9. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-FERROALLOY-ORE MILLS
USING LEACHING PROCESS
PARAMETER
PH
TSS
Oil and Grease
COD
Ammonia
Ai
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fluoride
CONCENTRATION dug/*)
IN EFFLUENT
30-day average
6" to 9«
20
3
50
5
0.6
0.05
0.05
0.05
0.7
0.5
0.1
1.0
0.1
10
24-hour maximum
6» to 9'
30
5
100
10
0.7
0.1
0.1
0.1
1.0
1.0
0.2
2.0
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric toni
_
90
14
230
23
2.3
0.23
0.23
0.23
3.2
2.3
0.45
4.5
0.45
45
lb/1000
ihort tons
_
180
28
460
45
4.5
0.46
0.46
0.46
6.4
4.5
0.9
9
0.9
90
24-hour maximum
kg/1000
metric tons
_
135
23
460
45
3.2
0.46
0.46
0.46
4.5
4.5
0.9
9
0.9
90
lb/1000
short tons
-
270
45
920
90
6.4
0.92
0.92
0.92
9.0
9.0
1.8
18
1.8
180
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-27
DRAFT
-------
DRAFT
Subcategory; Mercury Mills or Mine/Mills Using Flotation
Process
Identification of BATEA. The best available technology
economically achievable is zero discharge by the use of total
recycle and/or total Impoundment of process wastewater.
Subcategory; Mills Recovering Mercury as_ £ Byproduct of Base-
or Precious-Metal Concentrates
Identification of_ BATEA. No separate limitations or
technology are proposed. The BATEA for this Subcategory is
the same as BATEA for the primary base or precious metal
recovered.
Category; Uranium. Radium, and Vanadium Ores
Subcategory; Uranium Mines
Identification £f BATEA. The best available technology
economically achievable is the use of BPCTCA technology in
conjunction with sulfide precipitation, ion exchange for Mo
and V removal, and aeration.
Rationale for Selection. The use of sulfide precipitation
for removal of heavy metals has been demonstrated in the chlor-
alkalai industry, as well as in numerous pilot- and bench-scale
experimental treatment systems. Relatively simple, inexpensive
systems are available for use in implementing this treatment.
Ion-exchange technology has been demonstrated in the uranium
industry as effective in extraction of uranium values from
mine or process water. Ion-exchange resins are available which
are specific for the ions involved. Aeration of wastewater
will assist in raising dissolved oxygen levels and in lowering
of COD and TOC.
BUBJICT TO ru^tnr I«I«T',n« "EC°MMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-28
DRAFT
-------
DRAFT
Levels of_ Effluent Reduction Attainable. The parameters
selected for control and the effluent reductions attainable by
implementation of this technology are presented in Table
X-10.
Subcategory; Mills Processing Uranium Ores by Acid or
Combined Acid/Alkaline Leaching
Identification £f BATEA. The best available technology
economically achievable is zero discharge by the use of
impoundment and evaporation. (Same as BPCTCA.)
Subcategory; Mills Processing Uranium Ores by Alkaline
Leaching
Identification pjf BATEA. The best available technology
economically achievable is zero discharge by the use of
impoundment and recycle of mill process wastewater.
Metal Ores. Not Elsewhere Classified
Category; Antimony Ores
Subcategory; Antimony-Ore Mines (Alone)
Identification of_ BATEA. The best available technology
economically achievable for this Subcategory is chemical
(lime and sulfide) precipitation in conjunction with settling
impoundments. (Same as BPCTCA.)
Subcategory; Antimony Mills Using Flotation Process
Iden tification of_ BATEA. The best available technology
economically achievable is zero discharge by impoundment
and/or recycle of process wastewater. (Same as BPCTCA.)
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-29
DRAFT
-------
DRAFT
TABLE X-10. PARAMETERS SELECTED AND EFFLUENT
LIMITATIONS RECOMMENDED FOR
BATEA-URANIUM MINES
PARAMETER
PH
TSS
COD
TOC
As
Cd
Cr
Cu
Mo
Ni
Pb
V
Zn
Ra226
Total U
CONCENTRATION (mg/Jl)
IN EFFLUENT
30-day average
6» to 9"
20
50
20
0.1
0.004
0.05
0.01
0.5
0.1
0.05
1
0.1
1'
1
24-hour maximum
6»to9«
30
100
100
0.2
0.008
0.1
0.02
1.0
0.2
0.1
2
0.2
3*
3
•Values in pH units
Values in picocuries per liter
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-30
DRAFT
-------
DRAFT
Subcategory: Mills Obtaining Antimony As a_ Byproduct of
Base- or_ Precious-Metal Milling Operation
Identification of BATEA. No separate limitations are
proposed for this subcategory. Limitations developed for the
subcategory of the primary metal recovered are recommended
for this subcategory.
Category; Beryllium Ores
Subcategory: Beryllium Mills
Identification of_ BATEA. The best available technology
economically achievable is zero discharge by total impoundment
of process wastewater. (Same as BPCTCA.)
Category: Platinum Ores
Identification of BATEA. The best available technology
economically achievable is the use of settling ponds in
conjunction with the use of a coagulating agent. The use of
flocculants such as polyelectrolytes is an acceptable alter-
native.
Rationale for Selection. The use of coagulants and
flocculants to reduce suspended-solid concentrations has
had demonstrated application in the iron-ore milling industry
and the water-treatment Industry.
Levels of Effluent Reduction Attainable. The parameters
selected for control and the levels of effluent reduction
attainable for this category are presented in Table X-ll.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-31
DRAFT
-------
DRAFT
TABLE X-11. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR BATEA-PLATINUM MILLS
PARAMETER
pH
TSS
Total Fe
CONCENTRATION img/U
IN EFFLUENT
30-day average
6» to 9"
25
1.0
24-hour maximum
6« to 9"
30
2.0
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
—
0.36
0.014
lb/1000
short toni
-
0.72
0.029
24-hour maximum
kg/1000
metric tons
-
0.43
0.029
lb/1000
short tons
-
0.86
0.058
•Value in pH units
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION 'NJHIS REPORT ANO ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-32
DRAFT
-------
DRAFT
Category; Rare-Earth Ores
Subcategory: Mines Operated For Obtaining Primary or
Byproduct Rare-Earth Ores
Identification £f_ BATEA. The best available technology
economically achievable is zero discharge by impoundment
and/or reuse of mine water as process water in a mill.
(Same as BPCTCA.)
Subcategory; Rare-Earth-Ore Mills Uging Flotation and/or
Leaching Process
Identification £f BATEA. The best available technology
economically achievable is zero discharge by separation of
waste streams, followed by Impoundment and evaporation of
leachlng-process wastewater and recycle of flotation-process
water from a sedimentation impoundment. (Same as BPCTCA.)
Subcategory; Mills 0£ Mine/Mills Obtaining Rare-Earth
Minerals By_ Gravity Methods
BATEA for this Subcategory is covered under the appropriate
titanium-ore Subcategory. No separate limitations are
proposed.
Gategory: Tin Ores
No separate limitations are proposed for this category.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
X-33
DRAFT
-------
DRAFT
Category; Titanium Ores
Subcategory; Mines Obtaining Titanium Ore By_ Lode Mining
Identification o_f BATEA. The best available technology
economically achievable is neutralization in conjunction
with a settling pond for suspended-solid removal. (Same as
BPCTCA.) Maintenance of an alkaline pH will prevent
solubillzation of heavy metals and reduce their concentration
in the discharge waters.
Subcategory; Titanium Mills or Mine/Mills Using Electrostatic
and/or Magnetic plus Gravity and/or Flotation Methods
Identification o£ BATEA. The best available technology
economically achievable is zero discharge by tailing-pond
treatment and total recycle of the tailing-pond decant.
Diversion ditching and careful control to prevent runoff from
entering the impoundment will be necessary to maintain the
proper water balance. In addition, a small secondary pond
may be necessary to collect excess water from the primary
pond during periods of high precipitation. This water may
either be allowed to evaporate or be used as process makeup
water during drier periods.
Rationale for Selection. The single mill currently operating
in this Subcategory recycles its process water following
tailing-pond treatment. A discharge from this impoundment
currently exists on a seasonal basis. Diversion ditching
and a secondary pond should enable attainment of zero dis-
charge .
Levels o_f Effluent Reduction Attainable. Zero discharge of
process water is attainable by implementation of the above
technology.
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-34
DRAFT
-------
DRAFT
Subcategory; Titanium-Ore Mills Using Physical Milling
Methods in_ Conjunction With Dredge Mining
Identification of_ BATEA. The best available technology
economically achievable is settling impoundment with
maintenance of a pH of 3.5, secondary settling, and neutrali-
zation prior to discharge. (Same as BPCTCA.)
Category: Zirconium Ores
No separate limitations are recommended. The mining and
milling of zirconium (zircon) are practiced as a part of
titanium dredge mining.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION'IN THISREPORTT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
X-35
DRAFT
-------
DRAFT
SECTION XI
NEW SOURCE PERFORMANCE STANDARDS
AND PRETREATMENT STANDARDS
INTRODUCTION
This level of technology is to be achieved by new sources.
The term "new source" is defined in the Act to mean "any
source, the construction of which is commenced after the
publication of proposed regulations prescribing a standard
of performance." This technology is evaluated by adding, to
the consideration underlying the identification of best
available technology economically achievable, a
determination of what higher levels of pollution control are
available through the use of improved production processes
and/or treatment techniques. Thus, in addition to
considering the best in-plant and end-of-process control
technology, new source performance standards are how the
level of effluent may be reduced by changing the production
process itself. Alternative processes, operating methods,
or other alternatives were considered. However, the end
result of the analysis identifies effluent standards which
reflect levels of control achievable through the use of
improved production processes (as well as control technology),
rather than prescribing a particular type of process or
technology which must be employed.
The following factors were considered with respect to
production processes which were analyzed in assessing the
best demonstrated control technology currently available
for new sources:
(a) type of process employed and process changes;
(b) operating methods;
(c) batch, as opposed to continuous, operations;
(d) use of alternative raw materials and mixes of
raw materials;
(e) use of dry, rather than wet, processes (Including
substitution of recoverable solvents from water); and
(f) recovery of pollutants as byproducts.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-1
DRAFT
-------
DRAFT
In addition to the effluent limitations covering discharges
directly into waterways, the constituents of the effluent
discharge from a plant within the Industrial category which
would Interfere with, pass through, or otherwise be
incompatible with a well designed and operated publicly
owned activated sludge or trickling filter wastewater
treatment plant were identified. A determination was made
whether the introduction of such pollutants into the
treatment plant should be completely prohibited.
GENERAL WATER GUIDELINES
The process-water, cooling-water, and storm-water runoff
guidelines for new sources are Identical to those based on
best available technology economically achievable.
NEW SOURCE STANDARDS BY ORE CATEGORY
Based upon the Information contained in Sections III through
X of this report, the following determinations were made on
the degree of effluent reduction attainable with the.
application of new source standards for the various categories
and subcategorles of the ore mining and dressing industry.
The industry categories and subcategories which follow are
required to achieve no discharge of process wastewater based
upon best available technology economically achievable or
best practicable control technology currently available.
Iron-Ore Mills - Magnetic/Physical Process (Mesabi Range)
Copper Mines - Hydrometallurgical Process
Copper Mills - Vat Leaching
Copper Mills - Froth Flotation (Net Evaporation greater
than or equal to 76.2 cm (30 inches))
Copper Mills - Froth Flotation (Net Evaporation Less than
76.2 cm (30 inches))
Lead and Zinc Mills
Gold Mills - Cyanidatlon Process
Gold Mills - Amalgamation Process
Gold Mills - Froth-Flotation Process
Silver Mills - Froth-Flotation Process
Silver Mills - Cyanidation Process
Silver Mills - Amalgamation Process
Mercury Mills - Gravity-Separation Process
Mercury Mills - Flotation Process
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-2
DRAFT
-------
DRAFT
Uranium (Ra, V) Mills - Acid or Combined Acid/Alkaline
Leach Process
Uranium (Ra, V) Mills - Alkaline Leach Process
Antimony Mills - Flotation Process
Beryllium Mines
Beryllium Mills
Rare-Earth Mines
Rare-Earth Mills
Titanium Mills - Electrostatic, Magnetic or Gravity
Processes or Flotation Processes
The same limitations are recommended as new source standards.
New source standards identical to BPCTCA limitations are recommended
for the following industry categories:
Iron-Ore Mines
Iron-Ore Mills - Physical Methods/Chemical Reagents
Copper-Ore Mines
Lead and Zinc Mines (No Solubilization Potential)
Gold Mines
Gold Mills (Mine/Mills) - Gravity Separation
Silver Mines
Silver Mills (Mine/Mills) - Gravity Separation
Mercury Mines
Antimony Mines
Titanium Mines (Lode Ore)
Titanium Mills - Physical Processes with Dredge Mining
New source standards identical to BATEA limitations are
recommended for:
Lead and Zinc Mines (Solubilization Potential)
Bauxite Mines
Ferroalloy-Ore Mills (less than 5,000 metric tons (5,512
short tons)) per year
Ferroalloy-Ore Mills (more than 5,000 metric tons (5,512
short tons) per year) - Physical Methods
Ferroalloy-Ore Mills - Leaching Processes
Platinum Mines/Mills
Separate new source standards are recommended for the following
categories or subcategories as discussed on the pages which follow:
Ferroalloy-Ore Mines
Ferroalloy-Ore Mills (more than 5,000 metric tons (5,512
short tons) per year) - Flotation Processes
Uranium Mines
NOTICE- THESE ARC TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASEDUPON^OMMENTTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-3
DRAFT
-------
DRAFT
Category; Ferroalloy Ores
Subcategory; Mines Extracting Ores of Ferroalloy Metals
Identification of NSPS. For new operations, based upon infor-
mation contained in Sections III - X, a determination has been
made that the technology applicable to new sources is identical
to BATEA with the exception of copreclpltation or ion exchange
for molybdenum or vanadium removal. Therefore, the technology
recommended for use is lime precipitation in conjunction with
a settling pond, flocculant addition, and secondary settling.
Rationale for Selection. The selection of the above technology
is made on the basis of the best available, demonstrated
technology. The use of coprecipitation or ion exchange is not
recommended for a new source performance standard because neither
of these technologies has as yet been demonstrated, and both will
require some development prior to application in this subcategory.
Level of_ Effluent Reduction Attainable. The parameters selected
for control and the levels of effluent reduction attainable by
Implementation of the above technology are presented in Table XI-1,
Subcategory; Mills Processing More Than 5,OOP Metric Tons
(5.500 Short Tons) o.f Ferroalloy Ores per Year by_ Flotation
Methods
Identification of NSPS. The information contained in Sections
III through X indicates that the best available, demonstrated
technology applicable to new sources in this subcategory is
settling, process-water recycle, and oxidation (aeration,
chlorination, or ozonatlon). This technology is Identical to
BATEA with the exception of ion exchange or coprecipitation.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANQE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
XI-4
DRAFT
-------
DRAFT
TABLE XI-1. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR NSPS-FERROALLOY-ORE MINES
PARAMETER
PH
TSS
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
V
Zn
Fluoride
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6* to 9"
20
3
0.4
0.5
0.03
0.7
t
0.10
t
0.1
10
24-hour maximum
6"to9*
30
5
0.8
0.1
0.06
1.0
t
0.15
t
0.2
20
*Value in pH units
*No limitation proposed for NSPS
NOTICE- THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-5
DRAFT
-------
DRAFT
Rationale for Selection. The reasons for selection are discussed
in detail in Section X. The use of ion exchange or coprecipitstion
for removal of molybdenum and vanadium is not specified for this
level because the technologies have not yet been demonstrated
and will require some development prior to application in this
subcategory.
Level of Effluent Reduction Attainable. The parameters selected
for control and the levels of effluent reduction attainable by
implementation of the above technology are presented in Table XI-2.
Category; Uranium Ores
Subcategory; Uranium Mines
Identification of NSPS. Based on information contained in Sections
III through X of this report, the best available, demonstrated
technology applicable to new sources in this subcategory is the use
of settling ponds, lime precipitation, sulfide precipitation,
ion exchange (for uranium removal), barium chloride coprecipi-
tation (for radium'removal), secondary settling, and aeration.
Rationale for Selection. All technology selected for use in this
subcategory to attain NSPS levels has been demonstrated, in the
ore mining and dressing industry or in the chlor-alkalai industry.
The requirement for ion-exchange treatment (for molybdenum and
vanadium removal) is not included at this level because this
technology has not yet been demonstrated and will require some
development prior to application in this subcategory.
Levels of Effluent Reduction Attainable. The parameters selected
and the levels of effluent reduction attainable by implementation
of the above technology are presented in Table XI-3.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPOHT AND AH.
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-6
DRAFT
-------
DRAFT
TABLE XI-2. PARAMETERS SELECTED AND EFFLUENT LIMITATIONS
RECOMMENDED FOR NSPS-FERROALLOY ORE MILLS
USING FLOTATION PROCESS
PARAMETER
pH
TSS
Oil and Grease
COD
Cyanide
As
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
CONCENTRATION (mg/fc)
IN EFFLUENT
30-day average
6" to 9'
20
3
25
0.02
0.4
O.OS
0.03
0.7
0.5
0.1
0.1
10
24-hour maximum
6* to 9*
30
5
50
0.04
0.8
0.1
0.06
1.0
1.0
0.2
0.2
20
WASTE LOAD per unit ore milled
30-day average
kg/1000
metric tons
_
10
1.5
13
0.01
0.2
0.03
0.02
0.4
t
0.05
0.05
5
lb/1000
short tons
_
20
3.0
25
0.02
0.4
0.05
0.03
0.7
t
0.1
0.1
10
24-hour maximum
kg/1000
metric tons
—
15
2.5
25
0.02
0.4
0.05
0.03
0.5
t
0.1
0.1
10
lb/1000
short tons
-
30
5.0
50
0.04
0.8
0.1
0.06
1.0
t
0.2
0.2
20
•Value in pH uniti
fNo limitation proposed for NSPS
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANOE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-7
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TABLE XI-3. PARAMETERS SELECTED AND EFFLUENT
LIMITATIONS RECOMMENDED FOR
NSPS-URANIUM MINES
PARAMETER
pH
TSS
COD
TOC
As
Cd
Cr
Cu
Mo
Ni
Pb
V
Zn
Ra226
Total U
CONCENTRATION (mg/£)
IN EFFLUENT
30-day average
6» to 9*
20
50
20
0.1
0.004
0.05
0.01
••
0.1
0.05
**
0.1
1*
1
24-hour maximum
6" to 9"
30
100
100
0.2
0.008
0.1
0.02
••
0.2
0.1
••
0.2
31
3
•Values in pH units
^Values in picocuries per liter
••No limitation proposed for NSPS
NOTICE. THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
XI-8
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PRETREATMENT STANDARDS
Recommended pretreatment guidelines for discharge of plant
wastewater into public treatment works conform in general
with EPA Pretreatment Standards for Municipal Sewer Works
as published in the July 19, 1973 Federal Register and "Title
40 - Protection of the Environment, Chapter 1 - Environmental
Protection Agency, Subchapter D - Water Programs - Part 128 -
Pretreatment Standards," a subsequent EPA publication. The
following definitions conform to these publications.
Compatible Pollutant
The term "compatible pollutant" means biochemical oxygen
demand, suspended solids, pH and fecal coliform bacteria,
plus additional pollutants identified in the NPDES permit,
if the publicly owned treatment works was designed to treat
such pollutants, and, in fact, does remove such pollutants
to a substantial degree. Examples of such additional
pollutants may include.
chemical oxygen demand
total organic carbon
phosphorus and phosphorus compounds
nitrogen and nitrogen compounds
fats, oils, and greases of animal or vegetable
origin except as defined below in Prohibited
Wastes.
Incompatible Pollutant
The term "incompatible pollutant" means any pollutant which
is not a compatible pollutant as defined above.
Joint Treatment Works
Publicly owned treatment works for both non-industrial and
industrial wastewater.
Major Contributing Industry
A major contributing industry is an industrial user of the
publicly owned treatment works that: has a flow of 189.2 cubic
meters (50,000 gallons) or more per average work day; has a
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-9
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flow greater than five percent of the flow carried by the
municipal system receiving the waste; has, in its waste, a
- toxic pollutant in toxic amounts as defined in standards
Issued under Section 307 (a) of the Act; or is found by the
permit Issuance authority, in connection with the issuance
of an NPDES permit to the publicly owned treatment works
receiving the waste, to have significant impact, either singly
or in combination with other contributing industries, on that
treatment works or upon the quality of effluent from that
treatment works.
Pretreatment
Treatment of wastewaters from sources before introduction into
the Joint treatment works.
Prohibited Wastes
No waste introduced into a publicly owned treatment works
shall interfere with the operation or performance of the
works. Specifically, the following wastes shall not be
introduced into the publicly owned treatment works:
a. Wastes which create a fire or explosion hazard
in the publicly owned treatment works;
b. Wastes which will cause corrosive structural
damage to treatment works, but in no case wastes
with a pH lower than 5.0, unless the works are
designed to accommodate such wastes;
c. Solid or viscous wastes in amounts which would
cause obstruction to the flow in sewers, or other
interference with the proper operation of the
publicly owned treatment works; and
d. Wastes at a flow rate and/or pollutant discharge
rate which is excessive over relatively short time
periods so that there is a treatment process upset
and subsequent loss of treatment efficiency.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-10
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Pretreatment for Incompatible Pollutants
In addition to the above, the pretreatment standard for
incompatible pollutants introduced into a publicly owned
treatment works by a major contributing industry shall be
best practicable control technology currently available;
provided that, if the publicly owned treatment works which
receives the pollutants is committed, in its NPDES permit,
to remove a specified percentage of any incompatible
pollutant, the pretreatment standard applicable to users of
such treatment works shall be correspondingly reduced for
that pollutant; and provided further that the definition of
best practicable control technology currently available for
Industry categories may be segmented for application to
pretreatment if the Administrator determines that the
definition for direct discharge to navigable waters is not
appropriate for Industrial users of joint treatment works.
Recommended Pretreatment Guidelines
In accordance with the preceding Pretreatment Standards for
Municipal Sewer Works, the following are recommended for
Pretreatment Guidelines for the wastewater effluents:
a. No pretreatment is required for removal of
compatible pollutants—biochemical oxygen demand,
total organic carbon, chemical oxygen demand, and
fecal coliform bacteria.
b. Suspended-solid concentrations should be 500 mg/1
or less. If discharge to municipal treatment
systems is possible, many of the wastewaters
encountered in this study would require settling
treatment or sedimentation prior to conveyance to
a treatment plant.
c. Pollutants such as chemical oxygen demand; total
organic carbon; phosphorus and phosphorus com-
pounds; nitrogen and nitrogen compounds; and fats,
oils, and greases need not be removed, provided
that the publicly owned treatment works were
designed to treat such pollutants and will accept
them. Otherwise, levels should be at or below
BPCTCA guideline recommendations.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
XI-H
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d. A pH range of 6 to 9 is desirable for wastewater
treatment by biological methods.
e. Hazardous pollutants such as cyanides, chromates,
heavy metals, and other substances which would
interfere with microorganisms responsible for
organic-substance degradation in a treatment
facility should be restricted to those quantities
recommended in Section IX Guidelines for Best
Practicable Control Technology Currently Available.
Most of the mining and milling operations are located in
isolated, rural regions and have no access to treatment
facilities.
In addition,, the hydraulic loading to the treatment systems
should be as uniform as possible to maximize treatment
efficiency; therefore, the large volumes and high seasonal
discharges encountered in the ore mining and dressing industry
may have adverse effects upon treatment efficiencies.
In the relatively few instances where municipal treatment
systems may be used because of proximity, it may be necessary
to use chemical treatment and settling, pH control, and flow
equalization or regulation.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
XI-12
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SECTION XII
ACKNOWLEDGMENTS
The preparation of this report was accomplished through the
efforts of the staff of Calspan Corporation, Buffalo, N.Y.,
under the overall direction of Mr. D. Barry Dahm and Mr. Roland^
J. Pilie, Head and Assistant Head (respectively) of the company's
Environmental Systems Department. Dr. P. Michael Terlecky, Jr.,
directed the technical program, supervised the preparation of
the final report, and edited the final report.
Mr. Donald Gipe, Project Officer, Effluent Guidelines Division,
provided technical direction during all phases of the program
and throughout preparation of the final report. Mr. Gipe pro-
vided timely and careful review of the draft report and suggested
organizational, as well as technical and editorial, changes.
Mr. Ronald Kirby and Mr. William Renfro, Effluent Guidelines
Division, were most helpful in providing historical data,
data searches, and other technical assistance during all
phases of this program.
Mr. Allen Cywin and Mr. Ernst P. Hall, Jr.,- Director and
Assistant Director (respectively) of the Effluent Guidelines
Division, provided overall guidance and leadership, as well
as helpful suggestions, during the course of the program.
Mr. Harold B. Couglin, Chief, Effluent Guidelines Implemen-
tation Branch, provided technical, as well as administrative,
guidance throughout the course of this program.
Acknowledgment and appreciation are also given to the secre-
tarial staffs of the Effluent Guidelines Division and the
Calspan Corporation Environmental Systems Department for
their efforts in typing of drafts, necessary revisions, and
final preparation of the effluent guidelines document. Credit
is also due the composition, illustration, photomechanics, and
duplicating personnel of the company's Technical Services
Department for preparation of all illustrations and tables
in the report.
Appreciation is extended to the following trade associations
and individual cooperations for assistance and cooperation
during the course of this program:
XII-1
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Aluminum Association
American Iron Ore Association
American Mining Congress
Lead and Zinc Institute
Aluminum Company of America
Amax Lead Company of Missouri
American Exploration and Mining Company
American Smelting and Refining Company
Anaconda Copper Company
Atlas Corporation
Bethlehem Mines Corporation
Brush Wellman Company
Bunker Hill Company
Carlin Gold Mining Company
Cities Service Company
Cleveland-Cliffs Iron Company
Climax Molybdenum Company
Cominco American, Inc.
Continental Materials Corporation
Copper Range Company
Curtis Nevada Mines, Inc.
Cyprus-Bagdad Copper Corporation
Eagle Pitcher Industries, Inc.
E.I. duPont de Nemours and Company, Inc.
Erie Mining Company
Goodnews Bay Mining Company
Hanna Mining Company
Homestake Mining Company
Idarado Mining Company
Inspiration Consolidated Copper Company
Jones and Laughlin Steel Corporation
Kennecot Copper Corporation
Kerramerican, Inc.
Kerr McGee Corporation
Knob Hill Mines, Inc.
Magma Copper Company
Marquette Iron Mining Company
Molybdenum Corporation of America
National Lead Industries, Inc.
New Jersey Zinc Company
Oat Hill Mining Company
Olgleby-Norton Company - Eveleth Taconite
Phelps Dodge Corporation
Pickands Mather and Company - Erie Mining Company
Ranchers Exploration and Development Corporation
XI1-2
DRAFT
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Rawhide Mining Company
Reynolds Mining Corporation
Standard Metals Corporation
St. Joe Minerals Company
Sunshine Mining Company
Titanium Enterprises, Inc.
Union Carbide Corporation
United Nuclear Corporation
U.S. Antimony Corporation
U.S. Steel Corporation
White Pine Copper Company
The assistance of Regional Offices of the USEPA is greatly
appreciated. Assistance from the U.S. Geological Survey
and the Bureau of Mines is also gratefully acknowledged.
Mr. James Scott, of Environment Canada, provided helpful
information on current practices within the Canadian Mining
and Milling Industry. His consultation during this program
was of great assistance.
Our appreciation is extended to the staff of Calspan Corpora-
tion for their assistance during this program. Specifically,
thanks are due:
Mr. John Michalovic
Mr. Ditmar Bock
Mr. Hans Relf
Mr. Leo Ehrenreich
Mr. Edward Hodgson
Mrs. Sharron Pek
Mr. Kent Bainbridge
Mr. Michael Wilkinson
Mr. George Niesyty, Jr.
Mr. Milton D. Smith
Mr. Richard G. Swan
Dr. Jerry Huang
Mr. Robert Lockemer
Dr. John Y. Yang
Ms. Ginger Leu
Ms. Anne Meyer
Mr. James Terry
Mr. Jack Fisher
Mr. James C. Kasprzyk
Mr. Raymond King
XII-3
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Mrs. Helen D. Fabrizzi
Mrs. Joyce L. Terrano
Mrs. Mary A. Polino
Mrs. Betty L. Smith
Mrs. Diane J. Lavelle
Mrs. Alice I. Michalski
Mrs. Victoria H. Wienke
Mrs. M. Evelyn Perlmutter
Mrs. Mildred C. O'Rourke
Miss Donna L. Zurbrlck
In addition, the following consultants assisted in the tech-
nical program:
Colorado School ojf Mines Research Institute
Mr. Franklin Davis
Dr. David Hyatt
Mr. Wayne Whitacre
Michigan Technological University - Institute of
Mineral Research
Professor Duane Thayer
Mr. Carlo Karkanian
University of_ Missouri at^ Rolla
Dr. Bobby Wixson
Dr. J. Charles Jennett
Dr. Michael Hardie
£ &. M Corporation
Mr. Andrew McCord
XII-4
DRAFT
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DRAFT
SECTION XIII
REFERENCES
1. "Iron Ore," American Iron Ore Association, Cleveland,
Ohio, 1973.
2. Minerals Yearbook. Bureau of Mines, U.S. Department of
the Interior, Washington, 1972.
3. "Metals Statistics," American Metal Market, Fairchild
Publications, Inc., New York, 1974.
4. "Mineral Supply Bulletin," Mineral Industrial Surveys,
Bureau of Mines, U.S. Department of the Interior, Issues
of August 2 and 9, 1974.
5. Columbium and Tantalum. F.T. Slsco and E. Epremian,
John Wiley & Sons, New York, 1963.
6. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 556, 1963.
7. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 650, 1970.
8. Metallurgy of the Rarer Metals - _3_, Manganese, A.H. Sully,
Butterworth's Scientific Publications, London, 1955.
9. "Domestic Manganese from Butte Helps in Emergency,"
J.B. Huttl, Engineering and Mining Journal, Vol. 143,
No. 1, January 1942.
10. "Upgrading Manganese Ore," S.J. McCarroll, Mining
Engineering. March 1954.
11. "Complex Three Kids Ore," J.B. Hattl, Engineering and
Mining Journal. Vol. 156, No. 11, November 1955.
12. "Agglomeration Flotation of Manganese Ore," E.H. Gates,
Mining Engineering. December 1957.
XIII-1
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DRAFT
13. "Manganese from Low Grade Ores by the Ammonium
Carbonate Process," J.Y. Welsh and D.W. Patterson,
Journal of_ Metals. June 1957.
14. The Mining of Nickel. J.R. Boldt and P. Quieneau,
D. Van Nostrand, Inc., New York, 1967.
15. Personal Communication from Colorado School of Mines
Research Institute to Calspan Corporation, November 1974.
16. Tungsten. K.C. Li and C.Y. Wang, Reinhold Publishing
Corporation, New York, 1947.
17. Extractive Metallurgy, J. Newton, Wiley Books,
New York, 1959.
18. Technical Information on the Climax Property, Climax
Molybdenum Company, Climax, Colorado, 1974.
19. "Gravity Plant Flowsheet for Nigerian Columbite," F.A.
Williams, Transactions £f_ the Institute of Mining
Metallurgy. Vol. 67, 1957.
20. Thorium Production Technology, F.L. Cuthbert, Addison-
Wesley Publications Co., Reading, Mass., 1958.
21. "The Extractive Metallurgy of Uranium," R.C. Merritt,
Colorado School of Mines Research Institute, Assigned to
the General Manager of the USAEC, Library of Congress,
No. 71-157076, 1971.
22. "The Outlook for Mercury in the United States," V.A.
Cammarota, Proceedings of_ the First International
Mercury Congress, Barcelona, Spain, Vol. 1, pp. 93-98,
May 6-10. 1974.
23. "Copper Leaching Practices in the Western United States,"
Bureau of Mines, U.S. Department of the Interior, Washington,
Information Circular No. 8341, 1968.
XIII-2
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24. "Copper in the United States - A Position Survey,"
J.V. Beall, Mining Engineering, pp. 35-47, April 1973.
25. Mineral Processing, E.J. Pryor, Elsevier Publishing
Co., Ltd., New York, 1965.
26. "Reagent Control in Flotation," C.H.G. Bushell and
N. Malnarich, Mining Engineering, 1956.
27. "Polarographlc Determination of Xanthates in Flotation
Liquors, "Shuig-Chuan Sun and R.P. Holtzman, Analytical
Chemistry. Vol. 29, No. 9, September 1957.
28. "State-of-the-Art of Uranium Mining, Milling, and
Refining Industry," D.A. Clark, U.S. Environmental
Protection Agency, Washington, EPA-660/2-74-038, June 1974.
29. "Base Metal Mine Waste Management in Northeastern New
Brunswick," Environmental Impact and Assessment Report
EPS 8-WP-73-1, Water Pollution Control Directorate,
Environment Canada, Ottawa (Ont.), Canada, June 1973.
30. "The Solubility of Heavy Metal Hydroxides in Water,
Sewage, and Sewage Sludge - I, the Solubility of Some
Metal Hydroxides," S.H. Jenkins, D.G. Keight, and R.E.
Humphrey, Journal on_ Air and Water Pollution, No. 8,
pp. 537-556, 1964.
31. "Processes, Procedures, and Methods to Control Pollution
from Mining Activities," U.S. Environmental Protection
Agency, Washington, EPA-430/9-73-011, 1973.
32. "Development Document for Proposed Effluent Limitation
Guidelines and New Source Performance Standards for the
Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Non
Ferrous Metals Manufacturing Point Source Category,"
U.S. Environmental Protection Agency, Washington,
EPA-440/l-74-032b, November 1974.
XIII-3
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33. "Case History, Non Ferrous Metal Mining - Smelting
Industry, Process Wastewater Treatment at Bunker Hill
Co.," G.M. Baker, A.H. Larsen, Paper presented at AMC
Mining Convention, September 9-12, 1973, Denver, Colorado.
34. "Mercury Recovery from Contaminated Waste Water and
Sludges," U.S. Environmental Protection Agency, Washington,
EPA-660/2-74-086, December 1974.
35. "An Interdisciplinary Study of Transport and Biological
Effects of Molybdenum in the Environment," Colorado
University, Boulder, Colorado, Molybdenum Project, NSF
RA/E-72-005, December 1971.
36. "Removal of Dissolved Molybdenum from Industrial Wastewaters
by Ferric Oxyhydroxide," George R. LeGendre and D.D. Runnels,
Dept. of Geological Sciences, University of Colorado,
Boulder, Colorado, September 1, 1974.
37. "Transport and Biological Effects of Molybdenum in the
Environment," Colorado State University, Fort Collins,
Colorado, NSF RA/E-73-002, January 1973.
38. "Neutradesulfating Treatment Process for Acid Mine
Drainage," Catalytic, Inc., EPA-14010-DYH-12/71, 1971.
39. Water and Wastewater Engineering, Volume 2, G.M. Fair,
J.C. Geyer, J.C. Prun, John Wiley & Sons, New York, 1968.
40. "Traces of Heavy Metals in Water Removal Processes and
Monitoring," U.S. Environmental Protection Agency,
Washington,1 EPA-902/9-74-001, 1974.
41. "Precipitation Scavenging of Inorganic Pollutants from
Metropolitan Sources," Battelle-Paclfic, NW Laboratories,
EPA-650/3-74-005, 1974.
42. "Development Document for Effluent Limitations Guidelines
and New Source Performance Standards for the Smelting and
Slag Processing Segments of the Ferroalloy Manufacturing
Point Source Category," U.S. Environmental Protection
Agency, Washington, EPA-440/l-74-008a, February 1974.
XIII-4
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43. "Methods to Treat, Control and Monitor Spilled Hazardous
Materials," Monthly Progress Report No. 6, EPA Contract
68-01-0110, Calspan Corporation, Buffalo, New York,
December 1971.
44. "An Investigation of Techniques for Removal of Chromium
from Electroplating Wastes," Battelle Memorial Institute,
EPA-12010-EIE-03/71, 1971.
45. "Adsorption/Filtration Plant Cuts Phenols From Effluent,"
J.B. Henshaw, Chemical Engineering. May 31, 1971.
46. "Treatment of Ferrous Acid Mine Drainage with Activated
Carbon," C.T. Ford and J.F. Boyer, Jr., U.S. Environmental
Protection Agency, Washington, EPA-R2-73-150, 1973.
47. "Removal of Dissolved Contaminants from Mine Drainage,"
J.P. Miller, U.S. Environmental Protection Agency,
Washington, EPA-R2-72-130, 1972.
48. "Waste Wool as a Scavenger for Mercury Pollution in
Waters," J.P. Tratnyek, U.S. Environmental Protection
Agency, Washington, EPA-16080-HUB-02/72, 1972.
49. Water Treatment. G.V. James, Technical Press, Ltd., 1965.
50. "Peat Moss Bed Filtration Removes Heavy Metals from
Wastewaters," Environmental Science and Technology.
Vol. 8, No. 598, 1974.
51. "Adsorption of Hg (II) by Hydrous Manganese Oxides,"
R.A. Lockwood and K.Y. Chen, Environmental Science
and Technology. Vol. 7, No. 1028, 1973.
52. "Study of Reutilization of Wastewater Recycled through
Groundwater," D.F. Boen, e£ a^L., U.S. Environmental
Protection Agency, Washington, EPA-16060-DDZ-07/71,
1971.
53. "Process Design Manual for Sludge Treatment and Disposal,"
U.S. Environmental Protection Agency, Technology Trans-
fer Office, Washington, EPA-625/1-74-006, 1974.
XIII-5
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DRAFT
54. Ion Exchange Technology. F.C. Nachod and J. Schubert,
Academic Press, Inc., New York, 1956.
55. "Application of Reverse Osmosis to Acid Mine Drainage
Treatment," R.C. Wilmoth, U.S. Environmental Protection
Agency, Washington, EPA-670/2-73-100, December 1973.
56. "Evaluation of Ion Exchange Processes for Treatment of
Mine Drainage Waters," Office of Saline Water, Envlrex,
Inc., Milwaukee, Wisconsin, PB-22773A, January 1974.
57. "Mine Drainage Pollution Control via Reverse Osmosis,"
R.C. Wilmoth and R.D. Hill, Mining Engineering, pp.
45-47, March 1973.
58. "Development of a High Product Water Recovery System
for the Treatment of Acid Mine Drainage by Reverse
Osmosis," M.K. Gupta, Office of Saline Water, Envirex,
Inc., Milwaukee, Wisconsin, 1974.
59. "The Use, Characteristics, and Toxicity of Mine-Mill
Reagents in the Province of Ontario," J.R. Hawley,
Ontario Ministry of the Environment, Toronto, Canada,
1972.
60. "Liquid-Liquid Solvent Extraction in the Mineral Indus-
tries," J.O. Golden, Mineral Industries Bulletin, March
1973.
61. "Metallurgical Application of Solvent Extraction,"
J.B. Rosenbaum, et. al.., Bureau of Mines, U.S. Depart-
ment of the Interior, Washington, Information Circular
8502, 1971.
62. "Solvent Extraction," C. Hanson, Chemical Engineering,
August 26, 1968.
XIII-6
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63. "Multi-Stage Flash Evaporation System for the Purifi-
cation of Acid Mine Drainage," D.R. Maneval and S.
Lenezis, Proceedings of Society of Mining Engineers.
AIME Fall Meeting, 1970.
64. "Base Metal Mine Waste Management in Northeastern New
Brunswick," Montreal Engineering Company, Ltd., Freder-
ickton, N.B., Canada, EPS 8-WP-73-1, June 1973.
65. "Report On Lead and Zinc Ores," to Calspan Corporation,
B.C. Wixson, C.J. Jennett, and M.G. Hardie, University
of Missouri at Rolla, Missouri, August 1974.
66. "The Problem of Acid Mine Drainage in the Province of
Ontario," J.R. Hawley, Ministry of the Environment,
Toronto, Ontario, Canada, 1972.
67. "Effects of Recycling Mill Water in the New Lead Belt
of Southeast Missouri," F.H. Sharp and K.L. Clifford,
Proceedings of AIME Meeting. Chicago, Illinois,
February 26, 1973.
68. Data from Colorado School of Mines Research Institute,
Project J31120, October 15, 1974.
69. "Operational Experience with a Base Metal Mine Drainage
Pilot Plant," Technology Development Report EPS 4-WP-
74-8, Environment Canada, Environmental Protection
Service, Water Pollution Control Directorate, Ottawa,
Ontario, Canada, September 1974.
70. "The Development of National Waste Water Regulations
and Guidelines for the Mining Industry," CIM Bulletin.
November 1974.
XIII-7
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SECTION XIV
GLOSSARY
absorption - The process by which a liquid is drawn into and
tends to fill permeable pores in a porous
solid body; also the increase in weight of a
porous solid body resulting from the penetra-
tion of liquid Into its permeable pores.
acid copper - Copper electrodeposited from an acid solution
of a copper salt, usually copper sulfate.
acid cure - In uranium extraction, sulfation of moist ore
before leach.
acid leach - (a) Metallurgical process for dissolution of
values by means of acid solution (used on
sandstone ores of low lime content); (b)
In the copper industry, a technology employed
to recover copper from low grade ores and
mine dump materials when oxide (or mixed
oxide-sulflde, or low grade sulfide) minerali-
zation is present, by dissolving the copper
minerals with either sulfurlc acid or sulfuric
acid containing ferric iron. Four methods of
leaching are employed: dump, heap, in-situ.
and vat (see appropriate definitions).
acid mine water - (a) Mine water which contains free sulfuric
acid, mainly due to the weathering of iron
pyrites; (b) Where sulfide minerals break down
under the chemical influence of oxygen and
water, the mine water becomes acidic and can
corrode ironwork.
activator, activating agent - A substance which when added to
a mineral pulp promotes flotation in the
presence of a collecting agent. It may be
used to increase the floatability of a mineral
in a froth, or to reflect a depressed (sunk
mineral).
XIV-1
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adit - (a) A horizontal or nearly horizontal passage driven
from the surface for the working or unwatering
of a mine; (b) A passage driven into a mine
from the side of a hill.
adsorption - The adherence of dissolved, colloidal, or finely
divided solids on the surface of solids with
which they are brought into contact.
aeroflocs - Synthetic water-soluble polymers used as floccu-
lating agents.
all sliming - (a) Crushing all the ore in a mill to so fine a
state that only a small percentage will fail
to pass through a 200-mesh screen; (b) Term
used for treatment of gold ore which is ground
to a size sufficiently fine for agitation as a
cyanide pulp, as opposed to division Into
coarse sands for static leaching and fine slimes
for agitation.
alluminothermic process - The reduction of oxides in an exo-
thermic reaction with finely divided aluminum.
alluvial deposit; placer deposit - Earth, sand, gravel or
other rock or mineral materials transported
by and laid down by flowing water. Alluvial
deposits generally take the form of (1) sur-
face deposits; (2) river deposits; (3) deep
leads; and (4) shore deposits.
alunite - A basic potassium aluminum sulfate, KAia.COH)^ 804)2..
Closely resembles kaolinlte and occurs in simi-
lar locations.
amalgamation - The process by which mercury is alloyed with
some other metal to produce amalgam. It
was used extensively at one time for the
extraction of gold and silver from pulverized
ores, now is largely superseded by the cyanide
process.
XIV-2
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AN-FO - Ammonium nitrate - fuel oil blasting agents.
asbestos minerals - Certain minerals which have a fibrous
structure, are heat resistant, chemically
inert and possessing high electrical Insulat-
ing qualities, and are of sufficient flexi-
bility' to be woven. The two main groups are
serpentine and amphiboles. Chrysotile (fibrous
serpentine, 3MgO . 2Si02_ . 2H20) is the prin-
cipal commercial variety. Other commercial
varieties are amosite, crocidolite, actinolite,
anthophyllite, and tremolite.
azurite - A blue carbonate of copper, Cu3.(C03)2_(OH)2_, crystal-
lizing in the monoclinic system. Found as an
alteration product of chalcopyrite and other
sulflde ores of copper in the upper oxidized
zones of mineral veins.
bastnasite; bastnaesite - A greasy, wax-yellow to reddish-brown
weakly radioactive mineral, (Ce.La) (C03_)F, most
commonly found in contact zones, less often in
pegmatites.
bauxite - (a) A rock composed of aluminum hydroxides, essen-
tially A12_CO . 2H20. The principal ore of
aluminum; also used collectively for lateritic
aluminous ores, (b) Composed of aluminum
hydroxides and impurities in the form of free
silica, clay, silt, and iron hydroxides. The
primary minerals found in such deposits are
boehmite, gibbsite, and diaspore.
Bayer Process - Process in which impure aluminum in bauxite
is dissolved in a hot, strong, alkalai solu-
tion (normally NaOH) to form sodium aluminate.
Upon dilution and cooling, the solution hydro-
lyzes and forms a precipitate of aluminum
hydroxide.
bed - The smallest division of a stratified series and marked
by a more or less well-defined divisional plane
from the materials above and below.
beneflciation - (a) The dressing or processing of ores for
the purpose of (1) regulating the size of a
desired product, (2) removing unwanted constitu-
ents, and (3) improving the quality, purity,
assay grade of a desired product; (b) Concentration
XIV-3
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or other preparation of ore for smelting by
drying, flotation, or magnetic separation.
Best Available Technology Economically Achievable - The
level of technology applicable to effluent
limitations to be achieved by July 1, 1983,
for industrial discharges to surface waters
as defined by Section 301(b)(1)(A) of the Act.
Best Practicable Control Technology Currently Available -
The level of technology applicable to effluent
limitations to be achieved by July 1, 1977,
for industrial discharges to surface waters
as defined by Section 301(b)(l)(A) of the Act.
byproduct - A secondary or additional product.
carbon absorption - A process utilizing the efficient absorp-
tion characteristics of activated carbon to
remove both dissolved and suspended substances.
carnotite - A bright yellow uranium mineral, K2/U02)2.(V04)£ . 3H20.
cationic collectors - In flotation, amines and related organic
compounds capable of producing positively
charged hydrocarbon-bearing ions for the pur-
pose of floating miscellaneous minerals,
especially silicates.
cationic reagents - In flotation, surface active substances
which have the active constituent in the posi-
tive ion. Used to flocculate and to collect
minerals that are not flocculated by the rea-
gents, such as oleic acid or soaps, in which
the surface-active ingredient is the negative
ion.
cement copper - Copper precipitated by iron from copper sulfate
solutions.
cerium metals - Any of a group of rare-earth metals separable
as a group from other metals occurring with
them and in addition to cerium includes lan-
thanum, praseodymium, neodymium, promethlum,
samarium and sometimes europium.
cerium minerals - Rare earths; the important one is monazite.
chalcoclte - Copper sulfide, Cu2S.
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chalcopyrite - A sulfide of copper and iron,
chert - Cryptocrystalline silica, distinguished from flint by
flat fracture, as opposed to conchoidal fracture.
chromlte - Chrome iron ore, FeCr204^
chrysocolla - Hydrated copper silicate, CuSiCG . 2H20.
chrysotile - A metamorphic mineral, an asbestos, the fibrous
variety of serpentine. A silicate of magnesium,
with silica tetrahedra arranged In sheets.
cinnabar - Mercury sulfide, HgS.
claim - The portion of mining ground held under the Federal
and local laws by one claimant or association,
by virtue of one location and record. A claim
is sometimes called a 'location*.
clarification - (a) The cleaning of dirty or turbid liquids by
the removal of suspended and colloidal matter;
(b) The concentration and removal of solids from
circulating water in order to reduce the sus-
pended solids to a minimum; (c) In the leaching
process, usually from pregnant solution, e.g.,
gold-rich cyanide prior to precipitation.
classifier - (a) A machine or device for separating the con-
stituents of a material according to relative
sizes and densities thus facilitating concen-
tration and treatment. Classifiers may be
hydraulic or surface-current box classifiers.
Classifiers are also used to separate sand from
slime, water from sand, and water from slime;
(b) The term classifier is used in particular
where an upward current of water is used to
remove fine particles from coarser material;
(c) In mineral dressing, the classifier Is a
device that takes the ball-mill discharge and
separates it into two portions—the finished
product which is ground as fine as desired,
and oversize material.
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coagulation - The binding of Individual particles to form floes
or agglomerates and thus increase their rate of
settlement in water or other liquid (see also
flocculate) .
coagulator - A soluble substance, such as lime, which when
added to a suspension of very fine solid parti-
cles in water causes these particles to adhere
in clusters which will settle easily. Used to
assist in reclaiming water used in flotation.
collector - A heteropolar compound containing a hydrogen-carbon
group and an ionizing group, chosen for the
ability to adsorb selectively in froth flotation
processes and render the adsorbing surface
relatively hydrophoblc. A promoter.
columbite; tantalite; nlobite - A natural oxide of niobium
(columbium) , tantalum, ferrous iron, and
manganese, found in granites and pegmatites,
(Fe,Mn)(Nb, 18)206^.
concentrate - (a) In mining, the product of concentration;
(b) To separate ore or metal from its contain-
ing rock or earth; (c) The enriched ore after
removal of waste in a beneficiation mill, the
clean product recovered in froth flotation.
concentration - Separation and accumulation of economic min-
erals from gangue.
concentrator - (a) A plant where ore is separated into values
(concentrates) and rejects (tails). An appli-
ance in such a plant, e.g., flotation cell,
Jig, electromagnet, shaking table. Also called
mill; (b) An apparatus in which, by the aid of
water or air and specific gravity, mechanical
concentration of ores is performed.
conditioners - Those substances added to the pulp to maintain
the proper pH to protect such salts as NaCN,
which would decompose in an acid circuit, etc.
and CaO are the most common conditioners
conditioning - Stage of froth-flotation process in which the
surfaces of the mineral species present in a
pulp are treated with appropriate chemicals to
influence their reaction when the pulp is
aerated.
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copper minerals - Those of the oxidized zone of copper deposits
(zone of oxidized enrichment) include azurite,
chrysocolla, copper metal, cuprite, and mala-
chite. Those of the underlying zone (that of
secondary sulfide enrichment) include bornite,
chalcocite, chalcopyrlte, covellite. The zone
of primary sulfides (relatively low in grade)
includes the unaltered minerals bornite and
chalcopyrite.
crusher - A machine for crushing rock or other materials. Among
the various types of crushers are the ball-mill,
gyratory crusher, Hadsel mill, hammer mill, jaw
crusher, rod mill, rolls, stamp mill, and tube
mill.
cuprite - A secondary copper mineral, Cu20.
cyanidation - A process of extracting gold and silver as
cyanide slimes from their ores by treatment
with dilute solutions of potassium cyanide
and sodium cyanide.
cyanidation vat - A large tank, with a filter bottom, in
which sands are treated with sodium cyanide
solution to dissolve out gold.
cyclone - (a) The conical-shaped apparatus used in dust
collecting operations and fine grinding appli-
cations; (b) A classifying (or concentrating)
separator into which pulp is fed, so as to
take a circular path. Coarser and heavier
fractions of solids report at the apex of long
cone while finer particles overflow from cen-
tral vortex.
daughter - Decay product formed when another element undergoes
radioactive disintegration.
decant structure - Apparatus for removing clarified water from
the surface layers of tailings or settling
ponds. Commonly used structure include decant
towers in which surface waters flow over a gate
(adjustable in height) and down the tower to a
conduit generally buried beneath the tailings,
decant weirs over which water flows to a channel
external to the tailings pond, and floating
decant barges which pump surface water out of
the pond.
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dense-media separation - (a) Heavy-media separation, or sink/
float. Separation of heavy sinking from light
floating mineral particles in a fluid of Inter-
mediate density; (b) Separation of relatively
light (floats) and heavy ore particles (sinks),
by immersion in a bath of intermediate density.
Denver cell - A flotation cell of the subaeration type, in
wide use. Design modifications include receded-
disk, conical-disk, and multibladed impellers,
low-pressure air attachments, and special froth
withdrawal arrangements.
Denver Jig - Pulsion-suction diaphragm Jig for fine material,
in which makeup (hydraulic) water is admitted
through a rotary valve adjustable as to portion
of Jigging cycle over which controlled addition
is made.
deposit - Mineral or ore deposit is used to designate a natural
occurrence of a useful mineral or an ore, in
sufficient extent and degree of concentration
to invite exploitation.
depressing agent; depressor - In the froth flotation process,
a substance which reacts with the particle
surface to render it less prone to stay in
the froth, thus causing it to wet down as a
tailing product (contrary to activator).
detergents, synthetic - Materials which have a cleansing action
like soap but are not derived directly from fats
and oils. Used in ore flotation.
development work - Work undertaken to open up ore bodies as
distinguished from the work of actual ore
extraction or exploratory work.
dewater - To remove water from a mine usually by pumping,
drainage or evaporation.
differential flotation - Separating a complex ore into two
or more valuable minerals and gangue by flo-
tation; also called selective flotation.
This type of flotation is made possible by
the use of suitable depressors and activators.
discharge - Outflow from a pump, drill hole, piping system,
channel, weir or other discernible, confined
or discrete conveyance (see also point source).
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dispersing agent - Reagent added to flotation circuits to
prevent flocculation, especially of objection-
able colloidal slimes. Sodium silicate is
frequently added for this purpose.
dredge; dredging - A large floating contrivance for underwater
excavation of materials using either a chain of
buckets, suction pumps, or other devices to ele-
vate and wash alluvial deposits and gravel for
gold, tin, platinum, heavy minerals, etc.
dressing - Originally referred to the picking, sorting, and
washing of ores preparatory to reduction. The
term now includes more elaborate processes of
milling and concentration of ores.
drift mining - A term applied to working alluvial deposits
by underground methods of mining. The paystreak
is reached through an adit or a shallow shaft.
Wheelbarrows or small cars may be used for
transporting the gravel to a sluice on the
surface.
dump leaching - Term applied to dissolving and recovering
minerals from subore-grade materials from a
mine dump. The dump is irrigated with water,
sometimes acidified, which percolates into
and through the dump, and runoff from the
bottom of the dump is collected, and a mineral
in solution is recovered by chemical reaction.
Often used to extract copper from low grade,
waste material of mixed oxide and sulfide
mineralization produced in open pit mining.
effluent - The wastewater discharged from a point source to
navigable waters.
electrowinnlng - Recovery of a metal from an ore by means of
electrochemical processes, i.e., deposition of
a metal on an electrode by passing electric
current through an electrolyte.
eluate - Solutions resulting from regeneration (elution) of ion-
exchange resins.
eluent - A solution used to extract collected Ions from an
ion-exchange resin or solvent and return the
resin to its active state.
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exploration - Location of the presence of economic deposits and
establishing their nature, shape, and grade;
the investigation may be divided into (1) pre-
liminary and (2) final.
extraction - (a) The process of mining and removal of ore from
a mine. (b) The separation of a metal or
valuable mineral from an ore or concentrate.
(c) Used in relation to all processes that are used
in obtaining metals from their ores. Broadly,
these processes involve the breaking down of the
ore both mechanically (crushing) and chemically
(decomposition), and the separation of the metal
from the associated gangue.
ferruginous - containing iron.
ferruginous chert - A sedimentary deposit consisting of
chalcedony or of fine-grained quartz and variable
amounts of hematite, magnetite, or limonite.
ferruginous deposit - A sedimentary rock containing enough
iron to justify exploitation as iron ore.
The iron is present, in different cases, in
silicate, carbonate, or oxide form, occurring
as the minerals chamosite, thuringite, siderlte,
hematite, limonite, etc.
flask - A unit of measurement for mercury; 76.5 pounds (34.73 kg).
flocculant - An agent that induces or promotes flocculation
or produces floccules or other aggregate
formation, especially in clays and soils.
flocculate - To cause to aggregate or to coalesce into
small lumps or loose clusters, e.g., the cal-
cium ion tends to flocculate clays.
flocculating agent; flocculant - A substance which produces
flocculation.
flotation - The method of mineral separation in which a froth
created in water by a variety of reagents
floats some finely crushed minerals, whereas
other minerals sink.
XIV-10
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flotation agent - A substance or chemical which alters the
surface tension of water or which makes it
froth easily. The reagents used in the flo-
tation process include pH regulators, slime
dispersants, resurfacing agents, wetting agents,
conditioning agents, collectors, and frothers.
friable - Easy to break, or crumbling naturally.
froth, foam - In the flotation process, a collection of bubbles
resulting from agitation, the bubbles being
the agency for raising (floating) the particles
of ore to the surface of the cell.
frother(s) - Substances used in flotation processes to make air
bubbles sufficiently permanent principally by
reducing surface tension. Common frothers are
pine oil, creyslic acid, and amyl alcohol.
gangue - Undesirable minerals associated with ore.
glory hole - A funnel-shaped excavation, the bottom of which is
connected to a raise driven from an under-
ground haulage level or is connected through
a horizontal tunnel (drift) by which ore may
also be conveyed.
gravity separation - Treatment of mineral particles which
exploits differences between their specific
gravities. Their sizes and shapes also play
a minor part in separation. Performed by
means of Jigs, classifiers, hydrocyclones,
dense media, shaking tables, Humphreys spirals,
sluices, vanners and briddles.
grinding - (a) Size reduction into relatively fine particles.
(b) Arbitrarily divided into dry grinding per-
formed on mineral containing only moisture as
mined, and wet grinding, usually done in rod,
ball or pebble mills with added water.
heap leaching - A process used in the recovery of copper
from weathered ore and material from mine
dumps. The liquor seeping through the beds
Is led to tanks, where it is treated with
scrap iron to precipitate the copper from
solution. This process can also be applied
to the sodium sulflde leaching of mercury
ores.
heavy-media separation - See dense-media separation.
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hematite - One of the most common ores of iron, Fe20_3, which
when pure contains about 70% metallTc~iron
and 30% oxygen. Most of the iron produced
in North America comes from the iron ranges
of the Lake Superior District, especially the
Mesabi Range, Minnesota. The hydrated variety
of this ore is called limonite.
Huntington-Heberlein Process - A sink/float process employing
a galena medium and utilizing froth flotation
as the means of medium recovery.
hydraulic mining - (a) Mining by washing sand and soil away
with water which leaves the desired mineral.
(b) The process by which a bank of gold-bear-
ing earth and rock is excavated by a Jet of
water, discharged through the converging nozzle
of a pipe under great pressure. The debris is
carried away with the same water and discharged
on lower levels into watercourses below.
hydrolysate; hydrolyzate - A sediment consisting partly of chemi-
cally undecomposed, finely ground rock powder
and partly of insoluble matter derived from
hydrolytic decomposition during weathering.
hydrometallurgy - The treatment of ores, concentrates, and
other metal-bearing materials by wet processes,
usually involving the solution of some component,
and its subsequent recovery from the solution.
ilmenite - An iron-black mineral, FeO . TiO£. Resembles
magnetite in appearance but is readily dis-
tinguished by feeble magnetic character.
in-situ leach - Leaching of broken ore in the subsurface as
it occurs, usually in abandoned underground
mines which previously employed block-caving
mining methods.
ion(ic) exchange - The replacement of ions on the surface, or
sometimes within the lattice, of materials
such as clay.
iron formation - Sedimentary, low grade, iron ore bodies
consisting mainly of chert and fine-grained
quartz and ferric oxide segregated in bands
or sheets irregularly mingled (see also
taconite).
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Jaw crusher - A primary crusher designed to reduce large
rocks or ores to sizes capable of being
handled by any of the secondary crushers.
Jig - A machine in which the feed is stratified in water
by means of a pulsating motion and from
which the stratified products are separately
removed, the pulsating motion being usually
obtained by alternate upward and downward
currents of the water.
Jigging - (a) The separation of the heavy fractions of an
ore from the light fractions by means of a
Jig. (b) Up and down motion of a mass of
particles in water by means of pulsion.
laterite - Red residual soil developed in humid, tropical,
and subtropical regions of good drainage.
It is leached of silica and contains con-
centrations particularly of iron oxides
and hydroxides and aluminum hydroxides.
It may be an ore of iron, aluminum, man-
ganese, or nickel.
launder - (a) A trough, channel, or gutter usually of wood,
by which water is conveyed; specifically in
mining, a chute or trough for conveying pow-
dered ore, or for carrying water to or from
the crushing apparatus, (b) A flume.
leaching - (a) The removal in solution of the more soluble
minerals by percolating waters. (b) Extract-
ing a soluble metallic compound from an ore
by selectively dissolving it in a suitable
solvent, such as water, sulfuric acid, hydro-
chloric acid, etc. The solvent is usually
recovered by precipitation of the metal or
by other methods.
leach ion-exchange flotation process - A mixed method of
extraction developed for treatment of copper
ores not amenable to direct flotation. The
metal is dissolved by leaching, for example,
with sulfuric acid, in the presence of an
ion exchange resin. The resin recaptures the
dissolved metal and is then recovered in a
mineralized froth by the flotation process.
XIV-13
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leach precipitation float - A mixed method of chemical reac-
tion plus flotation developed for such copper
ores as chryaocolla and the oxidized minerals.
The value Is dissolved by leaching with acid,
and the copper is reprecipitated on finely
divided particles of Iron, which are then
recovered by flotation, yielding an impure
concentrate in which metallic copper predomi-
nates.
lead minerals - The most important industrial one is galena
(PbS), which Is usually argentiferous. In
the upper parts of deposits the mineral may
be altered by oxidation to cerussite (PbC03)
or anglesite (PbS04_) . Usually galena occurs
in intimate association with sphalerite (ZnS).
leucoxene - A brown, green, or black variety of sphene or
titanite, CaTiSiO_5, occurring as monoclinic
crystals. An earthy alteration product con-
sisting in most instances of rutlle; used
in the production of titanium tetrachloride.
lime - Quicklime (calcium oxide) obtained by calcining lime-
stone or other forms of calcium carbonate.
Loosely used for hydrated lime (calcium hydrox-
ide) and incorrectly used for pulverized or
ground calcium carbonate in agricultural lime
and for calcium in such expressions as car-
bonate of lime, chloride of lime, and lime
feldspar.
lime slurry - A form of calcium hydroxide in aqueous suspen-
sion that contains considerable free water.
limonite - Hydrous ferric oxide, FeO(OH) . nH20. An important
ore of iron, occurring in stalactltic, mammillary,
or earthy forms of a dark brown color, and as
a yellowish-brown powder. The chief constituent
of bog iron ore.
liquid-liquid extraction, solvent extraction - A process
in which one or more components are removed
from a liquid measure by intimate contact
with a second liquid, which is itself nearly
insoluble In the first liquid and dissolves
Lhe Impurities and not the substance that is
to be purified.
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lode - A tabular deposit of valuable mineral between definite
boundaries. Lode, as used by miners, is nearly
synonymous with the term vein as employed by
geologists.
magnetic separation - The separation of magnetic materials
from nonmagnetic materials using a magnet.
An important process in the benef iciation of
iron ores in which the magnetic mineral is
separated from nonmagnetic material, e.g.,
magnetite from other minerals, roasted pyrite
from sphalerite.
magnetic separator - A device used to separate magnetic from
less magnetic or nonmagnetic materials. The
crushed material is conveyed on a belt past
a magnet.
magnetite, magnetic iron ore - Natural black oxide of iron,
FeJJO^. As black sand, magnetite occurs in
placer deposits, and also as lenticular bands.
Magnetite is used widely as a suspension solid
in dense-medium washing of coal and ores.
malachite - A green, basic cupric carbonate, Cu2^(OH)2.C03^,
crystallizing in the monoclinic system. It
is a common ore of copper and occurs typically
in the oxidation zone of copper deposits.
manganese minerals - Those in principal production are pyro-
lusite, some psilomelane, and wad (impure
mixture of manganese and other oxides) .
manganese nodules - The concretions, primarily of manganese
salts, covering extensive areas of the ocean
floor. They have a layer configuration and
may prove to be an important source of man-
ganese.
manganese ore - A term used by the Bureau of Mines for ore
containing 35 percent or more manganese and
may include concentrate, nodules, or synthetic
ore.
manganiferous iron ore - A term used by the Bureau of Mines
for ores containing 5 to 10 percent manganese.
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manganiferous ore - A term used by the Bureau of Mines for
any ore of importance for its manganese con-
tent containing less than 35 percent manganese
but not less than 5 percent manganese.
mercury minerals - The main source is cinnabar, HgS.
mill - (a) Reducing plant where ore is concentrated and/or
metals recovered. (b) Today the term has
been broadened to cover the whole mineral
treatment plant in which crushing, wet grind-
Ing, and further treatment of the ore is con-
ducted, (c) In mineral processing, one machine,
or a group, used In comminution.
minable - (a) Capable of being mined, (b) Material that can
be mined under present day mining technology
and economics.
mine - (a) An opening or excavation in the earth for the pur-
pose of excavating minerals, metal ores or
other substances by digging, (b) A word for
the excavation of minerals by means of pits,
shafts, levels, tunnels, etc., as opposed to
a quarry, where the whole excavation is open.
In general the existence of a mine is deter-
mined by the mode in which the mineral is
obtained, and not by its chemical or geologic
character. (c) An excavation beneath the
surface of the ground from which mineral matter
of value is extracted. Excavations for the
extraction of ore or other economic minerals
not requiring work beneath the surface are
designated by a modifying word or phrase as:
(1) opencut mine - an excavation for removing
minerals which is open to the weather; (2)
steam shovel mine - an opencut mine in which
steam shovels or other power shovels are used
for loading cars; (3) strip mine - a stripping,
an openpit mine In which the overburden is
removed from the exploited material before the
material Is taken out; (4) placer mine - a
deposit of sand, gravel or talus from which
some valuable mineral Is extracted; and (5)
hydraulic mine - a placer mine worked by means
of a stream of water directed against a bank
of sand, gravel, or talus. Mines are commonly
known by the mineral or metal extracted, e.g.,
bauxite mines, copper mines, silver mines,
etc. (d) Loosely, the word mine is used to
XIV-16
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mean any place from which minerals are extracted,
or ground which it is hoped may be mineral
bearing, (e) The Federal and State courts
have held that the word mine, in statutes
reserving mineral lands, included only those
containing valuable mineral deposits. Dis-
covery of a mine: In statutes relating to
mines the word discovery is used: (1) In the
sense of uncovering or disclosing to view ore
or mineral; (2) of finding out or bringing to
the knowledge the existence of ore, or mineral,
or other useful products which were unknown;
and (3) of exploration, that is, the more
exact blocking out or ascertainment of a deposit
that has already been discovered. In this sense
it is practically synonymous with development,
and has been so used in the U.S. Revenue Act
of February 19, 1919 (Sec. 214, subdiv. A10,
and Sec. 234, subdiv. A9) in allowing depletion
to mines, oil and gas wells. Article 219 of
Income and War Excess Profits Tax Regulations
No. 45, construes discovery of a mine as:
(1) The bona fide discovery of a commercially
valuable deposit of ore or mineral, of a value
materially in excess of the cost of discovery
in natural exposure or by drilling or other
exploration conducted above or below the ground;
and (2) the development and proving of a mineral
or ore deposit which has been apparently worked
out to be a mineable deposit or ore, or mineral
having a value in excess of the cost of improv-
ing or development.
mine drainage - (a) Mine drainage usually implies gravity flow
of water to a point remote from mining operation.
(b) The process of removing surplus ground or
surface water by artificial means.
mineral - An inorganic substance occurring in nature, though
not necessarily of inorganic origin, which
has (1) a definite chemical composition, or
more commonly, a characteristic range of chemi-
cal composition, and (2) distinctive physical
properties, or molecular structure. With few
exceptions, such as opal (amorphous) and mer-
cury (liquid), minerals are crystalline solids.
XIV-17
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mineral processing; ore dressing; mineral dressing - The dry
and- wet crushing and grinding of ore or other
mineral -bear ing products for the purpose of
raising concentrate grade; removal of waste
and unwanted or deleterious substances from
an otherwise useful product; separation into
distinct species of mixed minerals; chemical
attack and dissolution of selected values.
modifier(s) - (a) In froth flotation, reagents used to control
alkalinity and to eliminate harmful effects of
colloidal material and soluble salts. (b)
Chemicals which Increase the specific attrac-
tion between collector agents and particle
surfaces, or conversely which increase the
wettabillty of those surfaces.
molybdenite - The most common ore of molybdenum, MoS:^.
molybdenite concentrate - Commercial molybdenite ore after
the first processing operations. Contains
about 90% Mo S2_ along with quartz, feldspar,
water, and processing oil.
monazite - A phosphate of the cerium metals and the principal
ore of the rare earths and thorium. Monoclinic.
One of the chief sources of thorium used in the
manufacture of gas mantles. It is a moderately
to strongly radioactive mineral, (Ce,La,Y,Th)P04^.
It occurs widely disseminated as an accessory
mineral in granitic igneous rocks and gneissic
metamorphic rocks. Detrital sands in regions
of such rocks may contain commercial quantities
of monazite. Thorium-free monazite Is rare.
New Source Performance Standard - Performance standards for the
Industry and applicable new sources as defined
by Section 306 of the Act.
nlccollte - A copper-red arsenide of nickel which usually
contains a little iron, cobalt, and sulfur.
It is one of the chief ores of metallic nickel.
nickel minerals - The nickel-iron sulfide, pentlandlte ((Ye,
Ni)_9S8) is the principal present economic
source of nickel, and garnierite (nickel-
magnesium hydrosilicate) is next in economic
Importance.
XIV-18
DRAFT
-------
DRAFT
.eic acid - A mono-saturated fatty acid, CH3_(CH2)7CH:CH(CH2.)7.
COOH. A common component of almost all naturally
occurring fats as well as tall oil. Most com-
mercial oleic acid is derived from animal tallow
or natural vegetable oils.
pen-pit mining, open cut mining - A form of operation designed
to extract minerals that lie near the surface.
Waste, or overburden, is first removed, and
the mineral is broken and loaded. Important
chiefly in the mining of ores of iron and
copper. '
ore - (a) A natural mineral compound of the elements of which
one at least is a metal. Applied more loosely
to all metalliferous rock, though it contains
the metal in a free state, and occasionally
to the compounds of nonmetallic substances,
such as sulfur, (b) A mineral of sufficient
value as to quality and quantity which may be
mined with profit.
ore dressing - The cleaning of ore by the removal of certain
valueless portion as by Jigging, cobbing,
vanning and the like. Synonym for concentra-
tion. The same as mineral dressing.
ore reserve - The term is usually restricted to ore of which
the grade and tonnage have been established
with reasonable assurance by drilling and other
means.
oxidized ores - The alteration of metalliferous minerals by
weathering and the action of surface waters,
and the conversion of the minerals into oxides,
carbonates, or sulfates.
oxidized zone - That portion of an ore body near the surface,
which has been leached by percolating water
carrying oxygen, carbon dioxide or other gases.
pegmatite - An igneous rock of coarse grain size usually
found as a crosscutting structure in a larger
igneous mass of finer grain size.
pelletizing - A method in which finely divided material is
rolled in a drum or on an inclined disk, so
that the particles cling together and roll up
into small, spherical pellets.
XIY-19
DRAFT
-------
DRAFT
pH modifiers - Proper functioning of a catlonic or anlonic
flotation reagent is dependent on the close
control of pH. Modifying agents used are
soda ash, sodium hydroxide, sodium silicate,
sodium phosphates, lime, sulfurlc acid, and
hydrofluoric acid.
placer mine - (a) A deposit of sand, gravel, or talus from
which some valuable mineral is extracted.
(b) To mine gold, platinum, tin or other
valuable minerals by washing the sand, gravel,
etc.
placer mining - The extraction of heavy mineral from a placer
deposit by concentration in running water.
It includes ground sluicing, panning, shovel-
Ing gravel into a sluice, scraping by power
scraper, excavation by dragline or extraction
by means of various types of dredging activities.
platinum minerals - Platinum, ruthenium, rhodium, palladium,
osmium, and iridium are members of a group
characterized by high specific gravity, unusual
resistance to oxidizing and acidic attack, and
high melting point.
point source - Any discernible, confined and discrete convey-
ance, including but not limited to any pipe,
ditch, channel, tunnel, conduit, well, discrete
fissure, container, rolling stock, concentrated
animal feeding operation, or vessel or other
floating craft, from which pollutants are or
may be discharged.
pregnant solution - A value bearing solution in a hydrometal-
lurgical operation.
pregnant solvent - In solvent extraction, the value-bearing
solvent produced in the solvent extraction
circuit.
prompter - A reagent used in froth-flotation process, usually
called the collector.
rare-earth deposits - Sources of cerium, terbium, yttrium,
and related elements of the rare-earth's group,
as well as thorium.
raw mine drainage - Untreated or unprocessed water drained,
pumped or siphoned from a mine.
XIV-20
DRAFT
-------
DRAFT
reagent - A chemical or solution used to produce a desired
chemical reaction; a substance used in assay-
ing or in flotation.
reclamation - The procedures by which a disturbed area can be
reworked to make it productive, useful, or
aethetically pleasing.
recovery - A general term to designate the valuable constitu-
ents of an ore which are obtained by metal-
lurgical treatment.
x
reduction plant - A mill or a treatment place for the extrac-
tion of values from ore.
roast - To heat to a point somewhat short of fuzing in order to
expel volatile matter or effect oxidation.
rougher cell - Flotation cells in which the bulk of the gangue
is removed from the ore.
roughing - Upgrading of run-of-mill feed either to produce a
low grade preliminary concentrate or to reject
valueless tailings at an early stage. Performed
by gravity on roughing tables, or in flotation
in a rougher circuit.
rutile - Titanium dioxide, T102^
scintillation counter - An Instrument used for the location of
radioactive ore such as uranium. It uses a trans-
parent crystal which gives off a flash of light
when struck by a gamma ray, and a photomultiplier
tube which produces an electrical impulse when the
light from the crystal strikes it.
*?
•elective flotation - See differential flotation.
lettling pond - A pond, natural or artificial, for recovering
solids from an effluent.
iderlte - An iron carbonate, FeCO_3-
*
Lime, slimes - A material of extremely fine particle size
encountered in ore treatment.
.udge - The precipitant or settled material from a wastewater.
urry - (a) Any finely divided solid which has settled out as
from thickeners. (b) A thin watery suspension.
XIV-21
DRAFT
-------
DRAFT
solvent extraction - See liquid-liquid extraction.
sphalerite - Zinc sulfide, ZnS.
stlbnite - An antimony sulfide, Sb2S3^. The most important
ore of antimony.
suction dredge - (a) Essentially a centrifugal pump mounted on a
barge. (b) A dredge in which the material is
lifted by pumping through a suction pipe.
sulfide zone - That part o'f a lode or vein not yet oxidized by
the air or surface water and containing sulfide
minerals.
surface active agent - One which modifies physical, electrical,
or chemical characteristics of the surface of
solids and also surface tensions of solids
or liquid. Used in froth flotation (see also
depressing agent, flotation agent).
tabling - Separation of two materials of different densities
by passing a dilute suspension over a slightly
inclined table having a reciprocal horizontal
motion or shake with a slow forward motion and
a fast return.
taconite - (a) The cherty or jaspery rock that encloses the
Mesabi iron ores in Minnesota. In a somewhat
more general sense, it designates any bedded
ferruginous chert of the Lake Superior District.
(b) In Minnesota practice, is any grade of
extremely hard, lean iron ore that has its
iron either in banded or well-disseminated
form and which may be hematite or magnetite,
or a combination of the two within the same
ore body (Bureau of Mines).
taconite ore - A type of highly abrasive iron ore now exten-
sively mined in the United States.
tailing pond - Area closed at lower end by constraining wall
or dam to which mill effluents are run.
XIV-22
DRAFT
-------
DRAFT
tailings - (a) The parts, or a part, of any incoherent or
fluid material separated as refuse, or
separately treated as inferior in quality or
value; leavings; remainders; dregs. (b) The
gangue and other refuse material resulting
from the washing, concentration, or treatment
of ground ore. (c) Those portions of washed
ore that are regarded as too poor to be
treated further; used especially of the debris
from stamp mills or other ore dressing machin-
ery, as distinguished from concentrates.
/
tall oil - The oily mixture of rosin acids, and other materials
obtained by acid treatment of the alkaline
liquors from the digesting (pulping) of pine
wood. Used in drying oils, in cutting oils,
emulsifiers, and in flotation agents.
tantalite - A tantalate of iron and manganese (Fe,Mn)Ta206^
crystallizing in the orthorhombic system.
tetrahedrlte - A mineral, the part with Sb greater than As
of the tetrahedrite-tenantite series,
Cu3_(Sb,As)S,3_. Silver, zinc, iron and mer-
cury may replace part of the copper. An
important ore of copper and silver.
thickener - A vessel or apparatus for reducing the amount
of water in a pulp.
thickening - (a) The process of concentrating a relatively
dilute slime pulp into a thick pulp, that is,
one containing a smaller percentage of mois-
ture, by rejecting liquid that is essentially
solid free. (b) The concentration of the
solids in a suspension with a view to recover-
ing one fraction with a higher concentration
of solids than in the original suspension.
tin minerals - Virtually all the industrial supply comes from
casslterite(Sn02J, though some has been
obtained from the sulfide minerals stannlte,
cylindrite, and frankeite. The bulk of cas-
siterite comes from alluvial workings.
titanium minerals - The main commercial minerals are rutile
(T102J and ilmenite (FeTiOS).
tyuyamunite - A yellow uranium mineral, Ca(UC£)2/V04_)2^ . 3H20.
It is the calcium analogue of carnotite.
XIV-23
DRAFT
-------
DRAFT
uraninite - Essentially U02_. It is a complex uranium mineral
containing also rare earths, radium, lead,
helium, nitrogen and other elements.
uranium minerals - More than 150 uranium bearing minerals
are known to exist, but only a few are common.
The five primary uranium-ore minerals are
pitchblende, uraninite, davidlte, coffinite,
and brannerite. These were formed by deep-
seated hot solutions and are most commonly
found in veins or pegmatites. The secondary
uranium-ore minerals, altered from the primary
minerals by weathering or other natural pro-
cesses, are carnotite, tyuyamunite and meta-
tyuyamunite (both very similar to carnotite),
torbernite and metatorbernite, autunlte and
meta-autunite, and uranophane.
vanadium minerals - Those most exploited for industrial use
are patronite (VS^), roscoelite (vanadium
mica), vanadinite (Pb5Cl(V04)3>), carnotite
and chlorovanadinlte.
vat leach - Employs the dissolution of copper oxide minerals
by sulfuric acid from crushed, non-porous ore
material placed in confined tanks. The leach
cycle is rapid and measured in days.
weir - An obstruction placed across a stream for the purpose
of diverting the water so as to make it flow
through a desired channel, which may be an
opening or notch in the weir itself.
wetting agent - A substance that lowers the surface tension
of water and thus enables it to mix more
readily. Also called surface active agent.
Wilfley table - A widely used form of shaking table. A plane
rectangle is mounted horizontally and can be
sloped about its long axis. Tt is covered
with linoleum (occasionally rubber) and has
longitudinal riffles dying at the discharge
end to a smooth cleaning area, triangular in
the upper corner. Gentle and rapid throwing
motion is used on the table longitudinally.
Sands, usually classified for size range are
fed continuously and worked along the table
with the aid of feedwater, and across riffles
downslope by gravity tilt adjustment, and
XIV-24
DRAFT
-------
DRAFT
added washwater. At the discharge end, the
sands have separated into bands, the heaviest
and smallest uppermost, the lightest and larg-
est lowest.
xanthate - Common specific promoter used in flotation of sul-
fide ores. A salt or ester of xanthlc acid
which is made of an alcohol, carbon disulfide
and an alkalai.
xenotime - A yttrium phosphate, YPOA^ often containing small
quantities of cerium, terbium, and thorium,
closely resembling zircon in crystal form
and general appearance.
yellow cake - (a) A term applied to certain uranium concen-
trates produced by mills. It is the final
precipitate formed in the milling process.
It is usually considered to be ammonium
diuranate, (NH4)2U2_07_, or sodium diuranate,
Na2U2p2.> but the composition is variable and
depends upon the precipitating conditions.
(b) A common form of triuranium octoxide,
U2I08., is yellow cake, which is the powder
obtained by evaporating an ammonia solution
of the oxide.
zinc minerals - The main source of zinc is sphalerite (ZnS),
but some smithsonite, hemimorphite, zinclte,
willemlte, and franklinite are mined.
zircon - A mineral, ZrSiO^. The chief ore of zirconium.
zircon, rutile, ilmenite, monazite - A group of heavy min-
erals which are usually considered together
because of their occurrence as black sand in
natural beach and dune concentration.
XIV-25
DRAFT
-------
DRAFT
CHEMICAL ELEMENTS
Values in parentheses represent the most stable known isotopes.
ATOMIC ATOMIC
SYMBOL NUMBER UEKJHT
Actinium
Aluminum
Americiiim
Antimony
Argon
Arsenic
Astatine
Dariuin
Derkelium
Beryllium
Bismuth
Boron
Bromine
Cadmium
Calcium
Californium
Carbon
Cerium
Cesium
Chlorine
Chromium
Cobalt
Columbium (Niobium)
Copper
Curium
Dysprosium
Erbium
Europium
Fluorine
Francium
Gadolinium
Gallium
Germanium
Gold
Hafnium
Helium
Hnlmimn
Hydrogen
Illinium (Promethium)
Indium
lodmo
Indium
Iron
Krypton
Lunllmnum
Ixiiid
Lithium
Lutecium
Magnesium
Manganese
Ac
Al
Am
Sb
A
As
At
Ba
Bk
Be
Bi
B
Br
Cd
Ca
Cf
C
Ce
Cs
Cl
Or
Co
Cb
Cu
Cm
Dy
Er
Eu
F
Fr
Gd
Ga
Ge
Au
II f
He
Ho
II
II
In
I
Ir
1'c
Kr
La
Pl>
Li
Lu
Mg
Mn
89
13
95
51
18
33
85
56
97
4
83
5
35
48
20
98
6
58
55
17
24
27
41
29
96
66
68
63
9
87
64
31
32
79
72
2
67
1
61
40
53
77
25
30
57
82
3
71
12
25
227
20 97
(241)
121.76
30944
7491
(211)
137.36
243(')
9.013
20900
1082
7991R
11241
4008
244O
12010
140.13
13291
35 457
5201
5894
9291
63 54
(242)
16246
167.2
1520
1900
(223)
15A9
69 72
7260
197.2
1786
4003
164 94
10080
(147)
11471!
126 '.HZ
193 1
55 sr»
s:i7
138 9'2
207 -Jl
(i 0-10
174 99
24.32
54.93
SYMBOL
Mercury
Molybdenum
Neodymiura
Neptunium
Neon
Nickel
Niobium
NitroKcn
Osmium
Oxygen
Palladium
Phosphorus
Plntinimi
Plutonium
Polonium
Potassium
Praseodymium
Prorr.cthium (Illinium)
Protactinium
Radium
Radon
Rhenium
Rhodium
Rubidium
Ruthenium
Samarium
Scandium
Selenium
Silicon
Silver
Sodium
Strontium
Sulfur
Tantalum
Tpchnctium
Tellurium
Trrhium
Thallium
Thorium
Thulium
Tin
Titanium
TuiiKston (Wolfram)
UiMiiium
V:iiiadium
Wolfiam (Tungsten)
Xonon
Yttcibium
Yttrium
Zinc
Zirconium
UK
Mo
Nd
Np
Nc
Ni
Nb
N
Os
O
Pd
P
Pt
Pu
Po
K
Pr
Pm
Pa
Ra
Rn
Re
Rh
Rb
Ru
Sm
Sc
So
Si
Ag
Na
Sr
R
Ta
Tc
Tc
Tb
Tl
Th
Tm
SH
Ti
W
U
V
W
Xo
Yb
Y
Zn
Zr
ATOMIC ATOMIC
NUMBER WEIUHT
80
42
60
93
10
28
41
7
76
8
46
15
78
94
84
19
59
61
91
88
80
75
45
37
44
62
21
34
14
47
11
38
16
73
43
52
65
81
90
69
50
22
74
92
23
74
54
70
39
30
40
200 61
!).-> 95
144 27
(237)
20 183
58 (i'J
02m
14 008
190 2
160000
10(i 7
30 98
195 23
(230)
210
39096
14092
(147)
231
226.05
222
18631
102 91
8548
101 7
150.43
4510
7896
2806
107880
22.997
8763
32 066
18088
(99)
127 61
1592
20-1 39
232 12
Ili04
US 70
4790
IM'.I'J
2.W 07
50 95
1S.19J
l.tl It
17.UI1
SS '12
<>r, :
-------
DRAFT
CONVERSION TABLE
MULTIPLY (ENGLISH UNITS) by TO OBTAIN (METRIC UNITS)
ENGLISH UNIT ABBREVIATION CONVERSION ABBREVIATION METRIC UNIT
acres
acre • feet
British Thermal
Units
British Thermal
Units/pound
cubic feet
cubic feet
cubic feet/minute
cubic feet/second
cubic inches
cubic yards
degrees Fahrenheit
feet
flask of mercury
gallons
gallons
gallons/day
gallons/minute
horsepower
inches
inches of mercury
miles (statute)
million gallons/day
ounces (troy)
pounds
pounds/square
inch (gauge)
pounds/square
inch (gauge)
square feet
square inches
tons (short)
tons (long)
yards
ac
acft
BTU
BTU/lb
cu ft
cu ft
cfm
cfs
cu in.
cu y
oF
ft
(76.5 Ib)
gal
gal
gpd
gpm
hp
in.
in. Hg
mi
mgd
troy oz
Ib
psig
psig
sqft
sq in.
t
longt
y
0405
1,233.5
0.252
0.555
0.028
28.32
0.028
1.7
16.39
0.76456
0.555 (OF-32)1
0.3048
34.73 '
0.003785
3.785
0.003785
0.0631
0.7457
2.54
0.03342
1.609
3,785 '
31.10348
0.454
(0.06805 psig + 1)1
5.1715
0.0929
6.452
0907
1.016
0.9144
ha
cu m
kgcal
kg cal/kg
cu m
1
cu m/min
cu m/mm
cu cm (or cc)
cu m
°C
m
kgHg
cu m
1
cu m/day
I/sec
kW
cm
aim
km
cu m/day
g
kg
atm
cm Hg
sq m
sqcm
kkg
kkg
m
hectares
cubic meters
kilogram - calorics
kilogram calorics/kilogram
cubic meters
liters
cubic meters/minute
cubic meters/minute
cubic centimeters
cubic meters
degrees Celsius
meters
kilograms of mercury
cubic meters
liters
cubic meters/day
liters/second
kilowatts
centimeters
atmospheres
kilometers
cubic meters/day
grams
kilograms
atmospheres (absolute)
centimeters of mercury
square meters
square centimeters
metric tons (1000 kilograms)
metric tons (1000 kilograms)
meters
Actual conversion, not a multiplier
XIV-27
DRAFT
------- |