EPA 440/1-75/061
GROUP II
Development Document for
Interim Final and Proposed Effluent
Limitations Guidelines and New Source
Performance Standards for the
Ore Mining and Dressing Industry
Point Source Category
Vol. I
vF
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DEVELOPMENT DOCUMENT
for
INTERIM FINAL AND PROPOSED
EFFLUENT LIMITATIONS GUIDELINES
and
NEW SOURCE PERFORMANCE STANDARDS
for the
ORE MINING AND DRESSING
POINT SOURCE CATEGORY
VOLUME I - SECTIONS I - VI
Russell E. Train
Administrator
Andrew W. Breidenbach, Ph.D
Acting Assistant Administrator for
Water and Hazardous Materials
\
^
Allen Cywin
Director, Effluent Guidelines Division
Donald C. Gipe
Project Officer
Ronald G. Kirby
Assistant Project Officer
Effluent Guidelines Division
Office of Water and Hazardous Materials
U.S. Environmental Protection Agency
Washington, D.C. 20460
October 1975
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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
application of the best practicable control technology
currently available (BPCTCA) and the degree of effluent
reduction attainable 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 technology, processes,
operating methods, or other alternatives.
Based upon the application of the best practicable control
technology currently available, 14 of the 41 subcategories
for which separate limitations are suggested can be operated
with no discharge of process waste water. With the best
available technology economically achievable, 21 of the 41
subcategories for which separate limitations are proposed
can be operated with no discharge of process waste water to
navigable waters. No discharge of process waste water
pollutants is also achievable as a new source performance
standard for 21 of the 41 subcategories.
Supporting data and rationale for development of the
proposed effluent limitation guidelines and standards of
performance are contained in this report (Volumes I and II) .
111
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CONTENTS (VOLUME I)
Section Pa§e
I CONCLUSIONS !
II RECOMMENDATIONS 3
III INTRODUCTION H
PURPOSE AND AUTHORITY H
SUMMARY OF METHODS USED FOR DEVELOPMENT 13
OF EFFLUENT LIMITATION GUIDELINES AND
STANDARDS OF TECHNOLOGY
SUMMARY OF ORE-BENEFICIATION PROCESSES 17
GENERAL DESCRIPTION OF INDUSTRY BY ORE 29
CATEGORY
IV INDUSTRY CATEGORIZATION 141
INTRODUCTION 141
FACTORS INFLUENCING SELECTION OF 143
SUBCATEGORIES IN ALL METAL ORE CATEGORIES
DISCUSSION OF PRIMARY FACTORS INFLUENCING 148
SUBCATEGORIZATION BY ORE CATEGORY
SUMMARY OF RECOMMENDED SUBCATEGORIZATION 169
FINAL SUBCATEGORIZATION 169
V WASTE CHARACTERIZATION 173
INTRODUCTION
SPECIFIC WATER USES IN ALL CATEGORIES 175
PROCESS WASTE CHARACTERISTICS BY ORE 176
CATEGORY
VI SELECTION OF POLLUTANT PARAMETERS 373
INTRODUCTION 373
GUIDELINE PARAMETER-SELECTION CRITERIA 373
v
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SIGNIFICANCE AND RATIONALE FOR SELECTION 374
OF POLLUTION PARAMETERS
SIGNIFICANCE AND RATIONALE FOR REJECTION 398
OF POLLUTION PARAMETERS
SUMMARY OF POLLUTION PARAMETERS SELECTED 400
BY CATEGORY
VI
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CONTENTS (VOLUME II)
Section Page
VII CONTROL AND TREATMENT TECHNOLOGY 403
INTRODUCTION 403
CONTROL PRACTICES AND TECHNOLOGY 404
TREATMENT TECHNOLOGY 419
EXEMPLARY TREATMENT OPERATIONS BY ORE 460
CATEGORY
VIII COST, ENERGY, AND NONWATER-QUALITY ASPECTS 567
INTRODUCTION 567
SUMMARY OF METHODS USED 567
WASTEWATER-TREATMENT COSTS FOR IRON-ORE 573
CATEGORY
WASTEWATER TREATMENT COSTS FOR COPPER-ORE 581
CATEGORY
WASTEWATER-TREATMENT COSTS FOR LEAD- AND 588
ZINC-ORE CATEGORY
WASTEWATER-TREATMENT COSTS FOR GOLD-ORE 600
CATEGORY
WASTEWATER-TREATMENT COSTS FOR SILVER-ORE 621
CATEGORY
WASTEWATER-TREATMENT COSTS FOR BAUXITE 631
CATEGORY
WASTEWATER-TREATMENT COSTS FOR FERROALLOY- 634
ORE CATEGORY
WASTEWATER TREATMENT COSTS FOR MERCURY- 658
ORE CATEGORY
WASTEWATER TREATMENT COSTS FOR URANIUM- 670
ORE CATEGORY
WASTEWATER TREATMENT COSTS FOR METAL 685
ORES, NOT ELSEHWERE CLASSIFIED
NON-WATER QUALITY ASPECTS 6"
vii
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IX BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY 703
AVAILABLE, GUIDELINES AND LIMITATIONS
INTRODUCTION 703
GENERAL WATER GUIDELINES 705
BEST PRACTICABLE CONTROL TECHNOLOGY 707
CURRENTLY AVAILABLE BY ORE CATEGORY
AND SUBCATEGORY
X BEST AVAILABLE TECHNOLOGY ECONOMICALLY 763
ACHIEVABLE, GUIDELINES AND LIMITATIONS
INTRODUCTION 753
GENERAL WATER GUIDELINES 764
BEST AVAILABLE TECHNOLOGY ECONOMICALLY 766
ACHIEVABLE BY ORE CATEGORY AND SUBCATEGORY
XI NEW SOURCE PERFORMANCE STANDARDS AND 795
PRETREATMENT STANDARDS
INTRODUCTION 795
GENERAL WATER GUIDELINES 796
NEW SOURCE STANDARDS BY ORE CATEGORY 796
PRETREATMENT STANDARDS 801
XII ACKNOWLEDGMENTS 809
XIII REFERENCES 813
XIV GLOSSARY 821
LIST OF CHEMICAL SYMBOLS 846
CONVERSION TABLE 847
Vlll
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TABLES (VOLUME I)
No., Title Pa2<
II-l Summary of Recommended BPCTCA Effluent Limitations 4
By Category and Subcategory—Ores for Which
Separate Limitations Are Proposed
II-2 Summary of Recommended BATEA Effluent Limitations 6
By Category and Subcategory—Ores for Which
Separate Limitations Are Proposed
II-3 Summary of Recommended NSPS Effluent Limitations 8
By Category and Subcategory—Ores for Which
Separate Limitations Are Proposed
HI-1 Iron-Ore Shipments for United States 30
III-2 Crude Iron-Ore Production for U.S. 31
III-3 Reagents Used for Flotation of Iron Ores 36
III-4 Various Flotation Methods Available for Pro- 37
duction of High-Grade Iron-Ore Concentrates
III-5 Total Copper-Mine Production of Ore by Year 42
III-6 copper-Ore Production from Mines by State (1972) 42
III-7 Average Copper Content of Domestic Ore 44
III-8 Average Concentration of Copper in Domestic ores 44
by Process (1972)
III-9 Copper Ore Concentrated in the United States 45
by Froth Flotation, Including LPF Process
(1972)
111-10 Heap or Vat Ore Leached in the United States (1972) 48
III-11 Average Price Received from Copper in the 50
United States
III-12 production of Copper from Domestic Ore by 51
Smelters
111-13 Mine Production of Recoverable Lead in the 53
United States
111-14 Mine Production of Recoverable Zinc in the 54
United States (Preliminary)
III-15 Domestic Silver Production from Different 66
Types of Ores
111-16 Silver Produced at Amalgamation and Cyanidation 67
Mills in the U.S. and Percentage of Silver
Recoverable from All Sources
III-17 Production of Bauxite in the United states 71
111-18 Production of Ferroalloys by U. S. Mining and 73
Milling Industry
111-19 Observed Usage of Some Flotation Reagents 95
III-20 Probable Reagents Used in Flotation of Nickel gg
and Cobalt ores
IX
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TABLES (cont.)
— Title
111-21 Domestic Mercury Production Statistics
III-22 Isotopic Abundance of Uranium
III-23 Uranium Milling Activity by State, 1972
111-24 Uranium Concentration in IX/SX Eluates
III-25 Decay Series of Thorium and Uranium
111-26 Uranium Milling Processes
III-27 Uranium Production
III-28 Vanadium Production
111-29 Vanadium Use 125
III-30 Production of Antimony from Domestic Sources
111-31 Domestic Platinum-Group Mine Production and Value
111-32 Production and Mine Shipments of Titanium 136
Concentrates from Domestic Ores in the US
I v-1 summary of Industry Subcategorization Recomended
IV-2 Final Subcategorization
V-l Historical Constituents of Iron-Mine Discharges
V-2 Historical Constituents of Waste water from Iron-
Ore Processing
V-3 Chemical Compositions of Sampled Mine Waters
V-4 Chemical Compositions of Sampled Mill Waters
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-6 Chemical Characteristics of Discharge Water
from Mine 1108
V-7 Characteristics of Mill 1108 Discharge Water
V-8 Principal Copper Minerals Used in the United States
V-9 Mine-Water Production from Selected Major Copper-
Producing Mines and Fate(s) of Effluent
V-10 Summary of Solid Wastes Produced by Plants
Surveyed
V-ll Raw Waste Load in Water Pumped from Selected
Copper Mines
V-12 1973 Water Usage in Dump, Heap, and In-Situ
Leaching Operations
V-l3 Chemical Characteristics of Barren Heap, Dump, or
In-Situ Acid Leach Solutions (Recycled: No
Waste Load)
V-14 Water Usage in Vat Leaching Process as a Function 71Q
of Amount of Product (Precipitate or Cathode
Copper) Produced
V-15 Chemical Characteristics of Vat-Leach Barren
Acid Solution (Recycled: No Waste Load)
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TABLES (cont.)
No. Title
V-16 Miscellaneous Wastes from Special Handling of 221
Ore Wash Slimes in Mine 2124 (No Effluent)
V-17 Examples of Chemical Agents Which May be Employed 225
In copper Flotation
V-18 Water Usage in Froth Flotation of Copper 227
V-19 Raw Mill Waste Loads Prior to Settling in Tailing 228
Ponds
V-20 Waste water Constituents and Waste Loads Resulting 233
from Discharge of Mill Process Waters
V-21 Range of Chemical Characteristics of Sampled Raw 241
Mine Water from Lead/Zinc Mines 3102, 3103,
and 3104
V-22 Range of Chemical Characteristics of Raw Mine 242
Waters from Four Operations in Solubiliza-
tion-Potential Subcategory
V-23 Ranges of constituents of Waste waters and Raw Waste248
Loads for Mills 3102, 3103, 3104, 3105, and
3106
V-24 Chemical Composition of Raw Mine Water from Mines 253
4105 and 4102
V-25 Process Reagent Use at Various Mills Beneficiating256
Gold Ore
V-26 Minerals Commonly Associated with Gold Ore 256
V-27 Waste characteristics and Raw Waste Loads at Four 258
Gold Milling Operations
V-28 Raw Waste Characteristics of Silver Mining 263
Operations
V-29 Major Minerals Found Associated with Silver Ores 266
V-30 Flotation Reagents Used by Three Mills to Bene- 267
ficiate Silver-Containing Mineral Tetrahedrite
(Mills 4401 and 4403) and Native Silver and
Argentite (Mill 4402)
V-31 Waste characteristics and Raw Waste Loads at Mills 268
4401, 4402, 4403, and 4105
V-32 Concentrations of selected Constituents in Acid 275
Raw Mine Drainage from Open-Pit Mine 5101
V-33 Concentrations of Selected constituents in Acid 275
Raw Mine Drainage from Open-Pit Mine 5102
V-34 concentrations of Selected constituents in Alkaline276
Raw Mine Drainage from Underground Mine 5101
V-35 Waste water and Raw waste Load for Open-Pit Mine 5101 278
V-36 Waste water and Raw Waste Load for Underground 278
Mine 5101
V-37 Types of Operations Visited and Anticipated— 279
Ferroalloy-Ore Mining and Dressing Industry
XI
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TABLES (cont.)
No.
Page
V~38 Chemical Characteristics of Raw Mine Water in
Ferroalloy Industry
V-39 Reagent Use in Molybdenum Mill 6101 289
V"40 *aw Waste Characterization and Raw Waste Load 289
V~41 Reagent Use for Rougher and Scavenger Flotation 292
at Mill 6102
Use for Cleaner Flotation at Mill 6102
^ Reagent Use at Byproduct Plant of Mill 6102 (Based 293
on Total Byproduct Plant Feed)
V~44 Mflj- 6102 Effluent Chemical Characteristics (com- 293
bined-Tailings Sample)
Y~f5 Chemical Characteristics of Acid-Flotation Step 295
V~46 Composite Waste Characteristics for Beneficiatior 299
at Mill 6104 (Samples 6, 8, 9, and 11)
V~47 Waste Characteristics from Copper-Thickener over- 299
flow for Mill 6104 (Sample 5)
V-48 Scheelite-Flotation Tailing Waste Characteristics 300
and Loading for Mill 6104 (Sample 7)
JJ"?« 50-Foot-Thickener Overflow for Mill 6104 (Sample 10)300
V 50 waste Characteristics of Combined-Tailing Discharge 301
for Mill 6104 (Samples 15, 16, and 17)
V~5i Waste Characteristics and Raw Waste Load at Mill
6105 (Sample 19)
v~52 Chemical Composition of Waste water, Total Waste, 302
and Raw Waste Loading from Milling and Smelter
Effluent for Mill 6106
V~53 Waste Characterization and Raw Waste Load for 306
Mill 6107 Leach and Solvent-Extraction Effluent
(Sample 80)
v~54 Waste Characteristics and Waste Load for Dryer 307
Scrubber Bleed at Mill 6107 (Sample 81)
v~55 Waste Characteristics and Loading for Salt-Roast 3ns
Scrubber Bleed at Mill 6107 (Sample 77)
v~56 Expected Reagent Use at Mercury-Ore Flotation
Mill 9202
v~57 Waste Characteristics and Raw Waste Loadings at
Mills 9201 and 9202
V-58 Waste Constituents Expected
v~59 Chemical and Physical Waste Constituents observed
in Representative Operations
Xll
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TABLES (cont. )
No.
V-60 Water Use and Flows at Mine/Mills 9401, 9402, 9403, 326
and 9404 .
V-61 Water Treatment Involved in U/Ra/V Operations J^b
V-62 Radionuclides in Raw Waste waters from Uranium/ 334
Radium/Vanadium Mines and Mills
V-63 Organic Constituents in U/Ra/V Raw Waste water 334
V-64 Inorganic Anions in U/Ra/V Raw Waste water 336
V-65 Light-Metal Concentrations Observed in U/Ra/V 336
Raw Waste water
V-66 Concentrations of Heavy Metals Forming Anionic 336
Species in U/Ra/V Raw Waste water
V-67 concentrations of Heavy Metals Forming Cationic 337
Species in U/Ra/V Raw Waste water
V-68 Other constituents Present in Raw Waste water in 337
U/Ra/V Mines and Mills
V-69 Chemical Composition of Waste water and Raw Waste 339
Load for Uranium Mines 9401 and 9402
V-70 Chemical Composition of Raw Waste water and Raw 339
Waste Load for Mill 9401 (Alkaline-Mill
Subcategory)
V-71 Chemical Composition of Waste water and Raw Waste 340
Load for Mill 9402 (Acid- or Combined Acid/
Alkaline-Mill Subcategory)
V-72 Chemical Composition of Waste water and Raw Waste 341
Load for Mine 9403 (Alkaline-Mill Subcategory)
V-73 Chemical Composition of Waste water and Raw Waste 342
Load for Mill 9404 (Acid- or Combined Acid/
Alkaline-Mill Subcategory)
V-74 Reagent Use at Antimony-Ore Flotation Mill 9901 349
V-75 Chemical Composition of Raw Waste water Discharged 350
From Antimony Flotation Mill 9901
V-76 Major Waste Constituents and Raw Waste Load at 351
Antimony Mill 9901
V-77 Chemical Composition of Raw Waste water from 353
Beryllium Mill 9902 (No Discharge from
Treatment)
V-78 Chemical Composition of Raw Waste water from 353
Rare-Earth Mill 9903
V-79 Results of Chemical Analysis for Rare-Earth 359
Metals (Mill 9903 — No Discharge)
xin
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TABLES (cont.)
— ^^ Page
V~80 Chemical Composition and Raw Waste Load from 361
Rare-Earth Mill 9903
V~81 Chemical Composition and Loading for Principal 363
Waste Constituents Resulting from Platinum
Mine/Mill 9904 (Industry Data)
V~82 Chemical Composition of Raw Waste water from 364
Titanium Mine 9905
v~83 Chemical Composition of Raw Waste water from 366
Titanium Mill 9905
V~8£* Reagent Use in Flotation Circuit of Mill 9905 ^*
V~°l Principal Minerals Associated with Ore of Mine 9905 367
v~bb Major Waste Constituents and Raw Waste Load at 367
Mill 9905
v~87 Chemical Composition of Raw Waste water at Mills 371
9906 and 9907
V~88 Raw Waste Loads for Principal Waste water Consti- 372
tuents from Sand Placer Mills 9906 and 9907
VI~1 Known Toxicity of Some Common Flotation Reagents 396
Used in Ore Mining and Milling Industry
VI~2 Summary of Parameters Selected for Effluent Limi- 401
tation by Metal Category
xiv
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FIGURES (VOLUME I)
Mo. Title
III-l Beneficiation of Iron Ores ^
III-2 Iron-Ore Flotation-Circuit Flowsheet *>
III-3 Magnetic Taconite Beneficiation Flowsheet J°
III-U Agglomeration Flowsheet ^9
III-5 Major copper Mining and Milling Zones of the U.S. 4J-
III-6 General Outline of Methods for Typical Recovery 46
of Copper from Ore
III-7 Major Copper Areas Employing Acid Leaching in
Heaps, in Dumps, or In Situ
III-8 Lead/Zinc-Ore Mining and Processing Operations 56
III-9 cyanidation of Gold Ore: Vat Leaching of Sands 61
and 'Carbon-in-Pulp1 Processing of Slimes
111-10 Cyanidation of Gold Ore; Agitation/Leach 63
Process
III-11 Flotation of Gold-Containing Minerals with 64
Recovery of Residual Gold Values by
Cyanidation
111-12 Recovery of Silver .Sulfide Ore by Froth 68
Flotation
III-13 Gravity-Plant Flowsheet for Nigerian Columbite 81
111-14 Euxenite/Columbite Beneficiation-Plant Flowsheet 82
III-15 Representative Flow Sheet for Simple Gravity 83
Mill
111-16 Simplified Molybdenum Mill Flowsheet 86
111-17 Simplified Molybdenum Mill Flow Diagram 88
111-18 Simplified Flow Diagram for Small Tungsten 91
Concentrator
III-19 Mill Flowsheet for a Canadian columbium 92
Operation
III-20 Flowsheet of Tristage crystallization Process 95
for Recovery of Vanadium, Phosphorus, and
Chromium from Western Ferrophosphorus
III-21 Arkansas Vanadium Process Flowsheet 96
III-22 Flowsheet of Dean-Leute Ammonium Carbamate 97
Process
111-23 Pachuca Tank for Alkaline Leaching 108
111-24 Concentration Processes and Terminology 112
III-25 Simplified Schematic Diagram of Sulfuric Acid 118
Digestion of Monazite Sand for Recovery
of Thorium, Uranium, and Rare Earths
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FIGURES (cont.)
— Title
111-26 Simplified Schematic Diagram of Caustic Soda 119
Digestion of Monazite Sand for Recovery
TTT __ _. of Thorium, Uranium, and Rare Earths
III-27 Effect of Acidity on Precipitation of Thorium, 120
Rare Earths and Uranium from a Monazite/
Sulfuric Acid Solution of Idaho and
Indian Monazite Sands
111-28 Generalized Flow Diagram for Production of 124
Uranium, Vanadium, and Radium
111-29 Beneficiation of Antimony Sulfide Ore by TOO
Flotation
TTT"^? Gravity Concentration of Platinum-Group Metals 133
TTT oi Beneficiation of Heavy-Mineral Beach Sands 130
111-32 Beneficiation of Ilmenite Mined from a Rock Deposit 140
v-i Flow Scheme for Treatment of Mine Water 183
aca 183
v-j Water Balance for Mine/Mill 1105 (September 1974)
v-4 Concentrator Flowsheet for Mill 1105
V-5 Flowsheet for Mill 1104 (Heavy-Media Plant) 19Q
V-6 Simplified Concentration Flowsheet for Mine/Mill 1108 194
V-7 Waste water Flowsheet for Plant 2120-B Pit TQQ
V-8 Flowsheet of Hydrometallurgical Process Used in 207
Acid Leaching at Mine 2122
V-9 Reactions by Which Copper Minerals Are Dissolved in 209
Dump, Heap, or In-Situ Leaching
V~10 Typical Design of Gravity Launder/Precipitation ?in
Plant
v~11 Cutaway Diagram of Cone Precipitator 211
V~12 Diagram of Solvent Extraction Process for Recovery 213
„ __ of Copper by Leaching of Ore and Waste
v~13 Vat Leach Flow Diagram (Mill 2124) 217
V~1{| Flow Diagram for Flotation of Copper (Mill 2120) 222
v 15 Addition of Flotation Agents to Modify Mineral 994
Surface
V~16 Flowsheet for Miscellaneous Handling of Flotation
Tails (Mill 2124)
v~17 Dual Processing of Ore (Mill 2124) 23fi
v~18 Leach/Precipitation/Flotation Process 007
v~19 Water Flow Diagram for Mine 3105
. .
v~20 Water Flow Diagram for Mine 3104
v~21 Flow Diagram for Mill 3103 247
xvi
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FIGURES (cont. )
No.
V-22 Water Flow in Four selected Gold Mining and 251
Milling Operations
V-23 Water Flow in Silver Mines and Mills
V-24 Process and Waste water Flow Diagram for Open-Pit
Bauxite Mine 5101
V-25 Mill 6601 Flowsheet
V-26 Simplified Mill Flow Diagram for Mill 6102
V-27 Internal Water Flow For Mill 6101 Through
Molybdenum Separation
V-28 Internal Water Flow for Mill 6104 Following 297
Molybdenum Separation
V-29 Water Use and Waste Sources for Vanadium Mill 6107 304
V-30 Water Flow in Mercury Mills 9101 and 9102 310
V-31 Typical Water-Use Patterns 316
V-32 Alkaline-Leach Water Flow 322
V-33 Ammonium Carbonate Leaching Process 324
V-34 Water Flow in Mills 9401, 9402, 9403, and 9404 327
V-35 Flowchart of Mill 9401 328
V-36 Flow Chart for Mill 9402 329
V-37 Flow Chart of Mill 9403 330
V-38 Flow Chart of Mill 9404 331
V-39 Water Flows and Usage for Mine/Mills 9901 (Antimonyp44
and 9902 (Beryllium)
V-40 Water Flows and Usage for Mine/Mills 9903 345
(Rare Earths) and 9904 (Platinum)
V-41 Water Flows and Usage for Titanium Mine/Mills 346
9905 and 9906
V-42 Beneficiation of Bertrandite, Mined from a Lode 355
Deposit by Flotation (Mill 9903)
V-43 Beneficiation of Rare-Earth Flotation Concentrate 356
by Solvent Extraction (Mill 9903)
V-44 Beneficiation and Waste Water Flow of Ilmenite 365
Mine/Mill 9905 (Rock Deposit)
V-45 Beneficiation of Heavy-Mineral Beach Sands (Rutile,370
Ilmenite, Zircon, and Monazite) at Mill 9906
xvn
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SECTION I
CONCLUSIONS
To establish effluent limitation guidelines and standards of
performance, the ore mining and dressing industry was
divided into 41 separate categories and subcategories for
which separate limitations were recommended. This report
deals with the entire metal-ore mining and dressing industry
and examines the industry by ten major categories: iron
ore; copper ore; lead and zinc ores; gold ore; silver ore;
bauxite ore; ferroalloy-metal ores; mercury ores; uranium,
radium and vanadium ores; and metal ores, not elsewhere
classified (ores of antimony, beryllium, platinum, rare
earths, tin, titanium, and zirconium). The
subcategorization of the ore categories is based primarily
upon ore mineralogy and processing or extraction 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 14 of 41 subcategories for which separate limitations
are proposed can be operated with no discharge of process
waste water. With the best available technology
economically achievable, facilities in 21 of the 41
subcategories can be operated with no discharge of process
waste water to navigable waters. No discharge of process
waste water is also achievable as a new source performance
standard for facilities in 21 of the 41 subcategories.
Examination of the waste water treatment methods employed in
the ore mining and dressing industry indicates that tailing
ponds or other types of sedimentation impoundments are the
most commonly used methods of suspended-solid removal, and
that these impoundments provide the additional benefit of
reduction of dissolved parameters as well. Tailing impound-
ments also serve to equalize flow rates and concentrations
of waste water parameters.
It is concluded that, for areas of excess water balance, the
practices of runoff diversion, segregation of waste streams,
and reduction in the use of process water will assist in the
attainment of no discharge for the specified subcategories.
Effective chemical-treatment methods which will result in
significant improvement in discharge-water quality and
pollutant waste loads beyond those attained by the
application of impoundment and settling are identified in
this report.
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SECTION II
RECOMMENDATIONS
The recommended effluent limitation guidelines based on the
best 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 14 of the 41 subcategories achieve no
discharge of process waste water.
The recommended effluent limitation guidelines based upon
the best available technology economically achievable
(BATEA) are summarized in Table II-2. Of the 41
subcategories listed for which separate limitations are
recommended, it is recommended that facilities in 21
subcategories achieve no discharge of process waste water by
1983.
The new source performance standards (NSPS) recommended for
operations begun after the proposal of recommended guide-
lines for the ore mining and dressing industry are
summarized in Table II-3. With the exception of four
subcategories, new source performance standards are
identical to BPCTCA and BATEA recommended effluent
limitations.
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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
X
IX-1
IX-2
COPPER ORES
Mines
Mills
( Open-Pit, Underground, Stripping
| Hydrometallurgical (Leaching)
( Vat Leaching
' Flotation
X
X
IX-3
IX-4
LEAD AND ZINC ORES
Mines
Mills
IX-5
1X6
GOLD ORES
Mines
Mills
{CyankJation Process
Amalgamation Process
Flotation Process
• Gravity Separation
X
IX-7
IX-8
1X9
IX 10
SILVER ORES
Mine*
Mills
{Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
X
IX 11
IX 12
IX 13
IX-14
BAUXITE ORE
I
IX-15
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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 DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mines/Mills
Mills (
> 6,000 metric tontt/yeer
< 5,000 metric tonst/year
> 5,000 metric tonsVyear 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 Leaching
Alkaline Leaching
X
X
IX-22
ANTIMONY ORES
Mines
Mills -
Flotation Process
X
IX-23
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
Mines
Mills
-------
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 < Pnvsical/Cnemical Separation
| Magnetic and Physical Separation
X
X-1
X-2
COPPER ORES
Mines / Open-Pit, Underground, Stripping
} Hydrometallurgical (Leaching)
{Vat Leaching
Flotation
X
X
X
X-3
LEAD AND ZINC ORES
Mines
Mills
X
X-4
GOLD ORES
Mines
{Cyarudation Process
Amalgamation Process
Flotation Process
Gravity Separation
X
X
X
X-5
(Same as BPCTCA)
SILVER ORES
Mines
! Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
X
X
X
X-6
(Same as BPCTCA)
BAUXITE ORE
Mines I!
X-7
-------
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
EFFLUENT
ZERO LIMITATIONS
DISCHARGE RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mine/Mills
Mills /
> 5,000 metric tonst/year
< 5,000 metric tons^/year
> 5,000 metric tons* /year by Physical Processes
> 5,000 metric tonsVyear by Flotation
Leaching
X-8
(SameasBPCTCA)
X-9
X-10
X 11
MERCURY ORES
Mines
Mills <
II X'12
Gravity Separation II X
Flotation Process II X
URANIUM, RADIUM, VANADIUM ORES
Mines
Mills \
Acid or Acid/Alkaline Leaching
Alkaline Leaching
X-13
X
X
ANTIMONY ORES
Mines
Mills
Flotation Process
(SameasBPCTCA)
X
BERYLLIUM ORES
Mines
Mills
X
X
PLATINUM ORES
Mines or Mine/Mills
(SameasBPCTCA)
RARE EARTH ORES
Mines
Mills -
Flotation or Leaching
X
X
TITANIUM ORES
Mines
Mills 1
Electrostatic/Magnetic and Gravity/Flotation Processes
Physical Processes with Dredge Mining
(SameasBPCTCA)
X
(SameasBPCTCA)
-------
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
j Physical/Chemical Separation
| Magnetic and Physical Separation
X
(SameasBATEA)
(Same as BATEA)
COPPER ORES
Mines
Mills
{Open-Pit, Underground .Stripping
Hydrometallurgical (Leaching)
f Vat Leaching
< Flotation
X
X
X
(SameasBATEA)
LEAD AND ZINC ORES
Mines
Mills
X
(Same as BATEA)
GOLD ORES
Mines
Mills
{Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation
X
X
X
(Same as BATEA)
(Same as BPCTCA)
SILVER ORES
Mines
Mills
(Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
X
X
X
(SameasBATEA)
(Same as BPCTCA)
BAUXITE ORE
(Same as BPCTCA)
-------
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
Mino/MilK
Mills |
> 5,000 metric tons'/veer
< 5,000 metric tonst/year
> 5,000 metric tons^/year by
> 5,000 metric tonsVyear by
Leaching
Physical Processes
Flotation
XI-1
XI-2
XI-3
(Same as BATEA)
MERCURY ORES
Mines
Mills <
Gravity Separation
Flotation Process
X
X
(Same as BPCTCAl
URANIUM, RADIUM, VANADIUM ORES
Mines
Mills j
Acid or Acid/Alkaline Leaching
Alkaline Leaching
X
X
XI-4
ANTIMONY ORES
Mines
Mills
Flotation Process
X
(Same as BPCTCA)
BERYLLIUM ORES
Mines
Mills
X
X
PLATINUM ORES
Mines or Mine/Mills
(Same as BPCTCA)
RARE-EARTH ORES
Mines
Mills -
Flotation or Leaching
X
X
TITANIUM ORES
Mines
Mills J
Electrostatic/Magnetic and Gravity/Flotation Processes
Physical Processes with Dredge Mining
X
(Same as BPCTCA)
(Same as BPCTCA)
-------
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
waters 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
lr 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 determined 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 standard
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
11
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for the Ore Mining and Dressing Industry point source
category.
Section 306 of the Act requires the Administrator, within
one year after a category of sources is included in a list
published pursuant 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 Administrator to propose
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 27 source
categories. Publication of an amended list will constitute
announcement 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 Management 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
12
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pollutants, the level of pollutant reduction attainable
through application of the best practicable control
technology currently available, and best available
technology economically achievable. Standards of
performance 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).
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 categories.
Development of reasonable industry categories and
subcategories and establishment of effluent guidelines and
treatment standards require a sound understanding and know-
ledge of the Ore Mining and Dressing Industry, the mining
techniques and milling processes involved, the mineralogy of
the ore deposits, water use, waste water 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 and beneficiating of
metals and ore minerals for the ore mining and dressing
industry. The survey and sampling and analysis covered a
wide range of processes, products, and types of wastes. In
each SIC category, slightly different evaluation criteria
were applied initially, depending upon the nature of the
extraction processes 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.
13
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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
waste water control, summaries of production of the
particular metals of interest, U.S. Bureau of Mines annual
summaries, 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 technology. Various mining
company personnel, independent researchers, and state and
federal environmental officials also supplied requested
information. In addition, Environment Canada provided
information on current practices within the Canadian Mining
and Dressing Industry.
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 waste water 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 waste waters 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 the industry and present in measurable
14
-------
quantities were selected as pollutants subject to effluent
limitation guidelines and standards.
Site Visits and Sampling Program
Based upon information gathered as part of the assembly of a
data base, examination of NPDES permits and permit appli-
cations, surveys by trade associations, and examination of
texts, journals, and the literature available on treatment
practices in the industry, selection of mining and milling
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, waste water control, and water
treatment practices was obtained. As a result of the
visits, many subcategories which had been tentatively
determined were found to be unnecessary. Flow diagrams were
obtained indicating the course of waste water 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 intermediate process or treatment
steps were performed as part of the site visits. In-situ
analyses for selected parameters such as temperature, pH,
dissolved oxygen, and specific conductance were performed
whenever possible. Historical data for the same waste
streams was obtained when available.
Raw waste characteristics were then identified for each sub-
category. This included an analysis of all constituents of
waste waters which might be expected in effluents from
mining and milling operations. In addition to examination
of candidate 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
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 Base
Cost information contained in this report was obtained
directly from industry during plant visits, from engineering
firms, equipment suppliers, and from the literature. The
15
-------
information obtained from these sources has been used to
develop general capital, operating and overall costs for
each treatment and control method. Where data was lacking,
costs were developed parametrically from knowledge of
equipment required, processes employed, construction, and
maintenance requirements. This generalized cost data plus
the specific information obtained from plant visits was then
used for cost effectiveness estimates in Section VIII and
wherever else costs are mentioned in this report.
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 environmental 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 treatment 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
control 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
16
-------
presently employed in the industry, but included also those
processes in pilotplant or laboratory-research stages.
SUMMARY OF ORE-BENEFICIATION 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 either 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 concentrate produced.
Frequently, depending on end use, a particular minimum grade
of concentrate is required, and only limited amounts of
specific gangue components are acceptable without penalty.
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-
17
-------
inclusive; rather, its purpose is to discuss the 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."
Gravity-Concentration Processes
General. Gravity-concentration processes exploit
differences 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 particle in a gravity separator. Smaller
particles of ore mineral may be 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. For
sufficiently small particles, even the smallest turbulence
suspends the particle for long periods of time, regardless
of density. Such slimes, once formed, cannot be recovered
by gravity techniques and may cause very low recoveries in
gravity processing of highly friable ores, such as scheelite
(calcium tungstate, CaWCW).
18
-------
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 lighter 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 concentrate 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), while 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. Reprocessing 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 provide an efficient
means of gravity separation for large volumes of material
between 0.1 mm and 2 mm (up to approximately 0.01 in.) in
diameter and have been widely applied—particularly, in the
processing of heavy sands for ilmenite (FeTiO_3) and monazite
(a rare-earth phosphate). They consist of a helical conduit
(usually, of five turns) about a vertical axis. A slurry of
ore is fed to the conduit 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
19
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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 (heavy media separation)
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 ore
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
"HuntingtonHeberlein" 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 ore and for
the separation of different valuable minerals recovered from
complex ores. Extensive use of magnetic separation is made
in the processing of ores of iron, columbium and tantalum,
and tungsten, to name a few. The separation is 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
20
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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 is added which prevents the bubbles from bursting when
they reach the surface, a layer of mineral-laden foam is
built up at the surface of the flotation cell which may be
removed to recover the mineral. Requirements for the
success of the operation are that particle si2e be small,
that reagents compatible with the mineral to be recovered be
used, and that water conditions in the cell not interfere
with attachment of reagents to mineral or to air bubbles.
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
21
-------
phenomenon, it can often be made highly specific, allowing
production of high-grade concentrates from very-low-grade
ore (e.g., over 95-percent MoS_2 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 five
basic types of compounds: pH conditioners (regulators,
modifiers) , 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 the collectors to specific kinds
of particles 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 (NaS^cOC2H5). Frothers are generally organics with
a soluble hydroxyl group and a "non-wettable" hydrocarbon.
Sodium cyanide is widely used as a pyrite depressant.
Activators useful in sulfide-ore flotation may include
cuprous sulfide and sodium sulfide. Other pyrite
depressants which are less damaging to the environment may
be used to replace the sodium cyanide. 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
22
-------
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, oxidized ores 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 leached 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 anionic 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 leaching in water.
(3) Uranium ores are only mildly soluble in water, but
they dissolve quickly in acid or alkaline
solutions.
(4) Native gold which is found in a finely divided
state is soluble in mercury and can be extracted by
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.
23
-------
Leaching solutions can be categorized as strong, general
solvents (e.g., acids) and weaker, specific solvents (e.g.,
cyanide). The acids dissolve certain 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 leach in a variety of ways. In vat
leaching, the process is carried out in a container (vat),
often equipped with facilities for agitation, heating,
aeration, and pressurization (e.g., Pachuca tanks). In-situ
leaching takes place in the ore body, with the leaching
solution applied either by plumbing or by percolation
through overburden. The pregnant leach solution is pumped
to the recovery facility and can often be recycled. In-situ
leaching is most economical when the ore body is surrounded
by impervious strata. When water suffices 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),
uranium using (sulfuric acid leach), and copper (using
sulfuric acid or acid ferric sulfate leach), are recovered
in this fashion.
24
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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 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 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. is 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 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
by the solution according to the reaction:
25
-------
+ 8NaCN + 2H20 + 02 — > 4NaAu (CN) 2 +
and subsequently sorbed onto activiated carbon ("Carbon-in-
J*ulp» process) or precipitated with metallic zinc according
to the reaction:
2NaAu(CN)2 + Zn --- > Na2Zn(CN) + 2Au
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
solution 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
SK>lubilized 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
Jmall 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 with 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 are deposited onto
stainless steel wool cathodes. The cathodes are then sent
to the refinery for processing.
j?retreatment of ores containing only finely divided gold and
Silver usually includes multistage crushing, fine grinding,
and classification of the ore pulp into sand and slime
fractions. The sand fraction then is leached in vats with
dilute, well aerated cyanide solution. The slime fraction,
after thickening, 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 countercurrent decantation
processing. Gold or silver is then recovered from the
^regnant leach solutions by the methods discussed above.
Different types of gold/silver ore reguire 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 cyanidation, the refractory graphitic
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
26
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"kerosene or fuel oil to inhibit adsorption of gold from
•solution.
Other refractory ores are those which contain sulfides.
Roasting to liberate the sulf ide-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 only 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 chemical oxidation.
(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 at a level that will minimize
consumption of cyanide by the dissolution of other
metal- bearing minerals.
(4) 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 the same
principle: Polar organic molecules tend to exchange a
mobile ion in their structure — typically, Cl- , NO3-, HSOU-,
or CO3_ — (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 + (U02 (804)3) ---- > R^UO^ (S(Ht) 3 + 4X-
27
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This reaction proceeds from left to right in the
process. Typical resins adsorb about ten percent e
mass in uranium and increase by about ?en percent in
density. m a concentrated solution of the mobile ion (for
rSveTsId - N-^drochl0-- -eld). the reaction 'can (* be
reversed, and the uranxum values are eluted (in this
acid) ' m general. the
for a metaiiic cation
Cr++ + > Mg-n- > Na+
andr because of decreasing ionic radius, with atomic number:
92U > 42Mo > 23V
and the separation of hexavalent 92U cations by ion 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) almost always appear in aqueous solutions
as oxidized ions (uranyl, vanadyl, or molybdate radicals),
with uranium and vanadium additionally complexed with
anionic 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 (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 pH 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
solventextraction concentrators that are useful even when
each stage effects only partial separation of a value from
an inter ferent. Unfortunately, the solvents are easily
polluted by slimes, and complete liquid/solid separation is
necessary. lonexchange and solvent-extraction circuits can
28
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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).
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 and
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 processing 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%
(Reference 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 breakdown of crude
iron-ore production in the U.S. is shown in Table III-2. _A
breakdown of U.S. iron-ore shipments by producing company is
given in Supplement B to this document. Except for a very
small tonnage, iron ores are beneficiated before shipping.
Beneficiation of iron ore includes such operations as crush-
ing, screening, blending, grinding, concentrating, classify-
ing, briquetting, sintering and agglomerating and is 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. As with many other natural
resources, future availability will largely be a matter of
cost rather than of absolute depletion as these lower-grade
ores are utilized. Benefication methods have been developed
to upgrade 20-30% iron 'taconite1 ores into high-grade
materials.
29
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TABLE II1-1. IRON-ORE SHIPMENTS FOR UNITED STATES
METRIC TONS LONG TONS
METRIC TONS LONG TONS
REGION
Great Lakes
Northeastern
Southern
Western
TOTAL U.S.
REGION
a. QUANTITIES SHIPPED BY REGION
AMOUNT SHIPPED
1968
65,093.239
3,602,706
3.474,203
10.566,860
82,736,905
64,065,185
3,545,805
3,419,333
10,399,972
1969
72,534,630
3,453,486
4,733,087
10,454,364
71,389,050
3,398,943
4,658,335
10,289,252
70,180,666
3,043,857
5,022,369
10,544,782
69,072,263
2,995,784
4,943,048
10,378,242
AMOUNT SHIPPED
1971
62,766,873
2,859,973
4,240,720
8,253,243
61,775,561
2,814,804
4,173,744
8.122,895
1972
METRIC TONS LONG TONS
METRIC TONS LONG TONS
METRIC TONS LONG TONS
65,759,357
2,362,067
4,032,651
7,397,815
64,720,783
2,324,762
3,968,961
7,266,471
77,504,865
2,405,456
3,923,518
8,462,579
76,280,787
2,367,465
3,861,552
8,328,925
b. SHIPMENTS FROM GREAT LAKES REGION AS PERCENTAGES OF TOTAL U S SHIPMENTS
YEAR
••
1968
1969
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
73.5
GREAT LAKES AGGLOMERATES
AS PERCENTAGE OF TOTAL
U.S. SHIPMENTS
487
506
52 3
56 3
61 8
61.7
CATEGORY
=======
Direct Shipping
Coarse Ores
Fine Ores
Screened Ores
Concentrates
Agglomerates
c. PERCENTAGES OF TOTAL U.S. SHIPMENTS
1968
8.2
3.2
28.3
60.3
100.0
1969
='" ' — -
7.0
3.1
27.5
62.4
100.0
YEAR
1970
5.0
2.7
28.2
64.1
100.0
1971
4.3
3.1
23.7
68.9
100.0
1972
;
2 0
12.8
11.9
73.3
100.0
1973
•
y A.
12.9
12.9
71.8
100.0
SOURCE: Reference 1
30
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TABLE II1-2. CRUDE IRON-ORE PRODUCTION FOR U.S.
QUANTITIES PRODUCED
YEAR
=====
1968
1969
1970
1971
1972
1973
PRODUCTION BY REGION
GREAT LAKES
METRIC TONS
159,349,027
169,328,525
172.799,898
161,947,509
158,183.907
186,627.840
LONG TONS
156,832,339
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
6,915,338
LONG TONS
10,075,038
9,575.011
9,028.913
7,651,428
6,615,513
6,806,120
SOUTHERN
METRIC TONS
^^
7,743,542
9,135,951
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
207,463,916
212.498,366
197,558,596
187,586.069
220,253.451
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
5.0
9.4
100.0
1971
82.0
3.9
4.8
9.3
100.0
1972
84.3
3.6
5.0
7.1
100.0
1973
84.7
3.2
3.9
8.2
100.0
SOURCE: Reference 1
31
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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
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 oresr 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 mav
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
blasting are usually 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 conveyors, skip hoists, or a
combination of these. It is then transferred to a crushing
plant for size reduction, to a screening plant for sizing,
or to a concentrating plant for treatment by washing (wet
size classification and tailings rejection) or by gravity
separation. ^ y
Special problems are associated with the mining of taconite.
The extreme hardness of the ore necessitates additional
drilling/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
techniques 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 workings, and elaborate pumping facilities
32
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usually requires expensive capital investments. Production
in terms of iron ore/day is much lower than in the case of
open pit production, necessitating the presence of very high
grade ores for economic recovery. 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.
Crude ore not requiring further processing may be crushed
and screened 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 product 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
attached to ores conditioned with flotation reagents
separate out iron ore during the flotation process, while
magnetic separation techniques are used where ores
containing magnetite are encountered.
At the present time, there are only three 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
III-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. Higher grade ores
containing hematite cannot be upgraded much above 55% iron.
Agglomerating processes follow concentration operations and
increase the particle size of iron ore and reudces "fines"
which normally would be lost in the flue gases. Sintering,
pelletizing, briquetting, and nodulizing 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. Pelletizing involves the
33
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Figure 111-1. BENEFICIATIOIM OF IRON ORES
t t
WASHING JIGG
SINTERING
1
ORE
CRUSHING AND
SCREENING
t
BLENDING
V
CONCENTRATING PROCESSES:
PHYSICAL
1 t
|NG MAGNETIC IfAXr
SEPARAT.ON SEp^^ON
+
AGGLOMERATION
PROCESSES
PELLETIZING NODULIZING
CHEMICAL
|
FLOTATION
1
BRIQUETTING
f
TO STOCK PILE AND/OR SHIPPING
34
-------
Figure 111-2. IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
DENSIFYING THICKENER
UNDERFLOW
CONDITIONERS
1
ROUGHER FLOTATION
ROUGHER
CONCENTRATE
(10 CELLS)
^.,1
I
1
FROTH OF
FIRST 2 CELLS
I
CLEANER FLOTATION
CLEANER
TAIL CLE/
1 CONCEIT
ROUGHER
TAIL
1
FROTH OF
XNER FIRST 2 CELLS
JTRATE | _
(8 CELLS)
RECLEANER
1
RECLEANER
TAIL
FLOTATION
1
RECLEANER
CONCENTRATE
(7 CELLS)
I
to»
TO
TAILING
BASIN
TOTAL
FLOTATION
CONCENTRATE
J
TO AGGLOMERATION
(FIGURE 111-4)
35
-------
TABLE 111-3. REAGENTS USED FOR FLOTATION OF IRON ORES
represent approximate maximum usages. Exact chemical composition of reagent
'• Anionic 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 lb/,hort 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)
*• Anionic Flotation of Iron Oxides (from crude ore)
Low-rosin tall oil fatty acid: 0.5 kg/metric ton (1 Ib/short ton)
3- Cationic 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)
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
36
-------
TABLE 111-4. VARIOUS FLOTATION METHODS AVAILABLE FOR PRODUCTION
OF HIGH-GRADE IRON-ORE CONCENTRATE
1. Anionic flotation of specular hematite
2. Upgrading of natural magnetite concentrate by cationic flotation
3. Upgrading of artificial magnetite concentrate by cationic flotation
4. Cationic flotation of crude magnetite
5. Anionic flotation of silica from natural hematite
6. Cationic flotation of silica from non-magnetic iron formation
37
-------
Figure 111-3. MAGNETIC TACONITE BENEFICIATION FLOWSHEET
CRUSHED CRUDE ORE
i
>M
I
COBBER MAGNETIC SEPARATION
CONCENTRATE
BALL MILL
±
CLEANER MAGNETIC SEPARATION
CONCENTRATE
I
HYDROCYCLONE
OVERSIZE UNDERSIZE
HYDROSEPARATOR
CONCENTRATE
i
FINISHER MAGNETIC SEPARATION
CONCENTRATE
i
KEI
I
FILTER
T
TO PELLETIZING
(FIGURE 111-1}
TAILING
TO TAILING BASIN
38
-------
Figure 111-4. AGGLOMERATION FLOWSHEET
CONCENTRATE FILTER CAKE
BALLING DRUM
SCREEN
UNDERSIZE OVERSIZE
FUEL
I
AGGLOMERATION FURNACE
PELLETS EXHAUST GASES
t I
TO STOCK PILE TO ATMOSPHERE
AND/OR SHIPPING
39
-------
formation of pellets or balls of iron ore fines, followed by
heating. (Figure III-4 illustrates a typical pelletizing
operation.) Nodules or lumps are 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 grinding 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 processing the ores and wastes
by hydrometallurgical and/or physical-chemical processes.
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
Montana
Nevada
Michigan
Tennessee
40
-------
Figure 111-5. MAJOR COPPER MINING AND MILLING ZONES OF THE U.S.
MINING AND MILLING COPPER AS A PRIMARY METAL
I:*:*:] MINING AND MILLING COPPER AS A COPRODUCT
-------
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 III-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
-------
A series of tables follow which give statistics for the U.S.
copper industry. Table III-5 lists total copper mine
production of ore by year, and Table III-6 gives copper ore
production by state for 1972. The average copper content of
domestic ores is given by Table III-7. Th- average
concentration of copper recovered from domestic ores,
classified by extraction process, is listed in Table III-8.
Copper concentrate production by froth flotation is given in
Table III-9, while production 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 remainder 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; recovery of copper by
electromining amounted to 12.5%.
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 processing. 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
molybdenum byproduct ore, 18,824 metric tons (20,750 short
tons) of byproduct molybdenum were produced.
Processes Employed to Extract Copper from Ore. The mining
methods 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.
43
-------
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 III-8. AVERAGE CONCENTRATION OF COPPER IN DOMESTIC ORES
BY PROCESS (1972)
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
—
0.07*
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
44
-------
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
45
-------
Figure 111-6. GENERAL OUTLINE OF METHODS FOR TYPICAL RECOVERY OF
COPPER FROM ORE
ORE K01%Cu)
»- OVERBURDEN AND
WASTE DUMPS
TO MARKET
(OR REFINERY)
46
-------
Processing 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. Hydrometallurgical pro-
cesses currently employ sulfuric acid (5-10%) or iron
sulfate to dissolve copper from the oxide or mixed oxide-
sulfide ores in dumps, heaps, vats or in-situ (Table III-
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 extract!on-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, tellurium, and iron. These are
removed in the refinery.
One 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. Also, 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 electrowinning 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-11). 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 six
47
-------
TABLE 111-10. COPPER ORE HEAP OR VAT 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
48
-------
Figure 111-7. MAJOR COPPER AREAS EMPLOYING ACID LEACHING IN HEAPS,
IN DUMPS, OR IN SITU
LEACHING ZONES
-------
TABLE 111-11. AVERAGE PRICE RECEIVED FROM COPPER
IN THE UNITED STATES
YEAR
1865 - 1874
1907
1910
1915
1917
1920
1925
1930
1932
1935
1940
1945
1950
1955
1960
1965
1970
1972
1973
PRICE IN CENTS PER KILOGRAM (CENTS PER POUND)
LAKE COPPER*
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 (14.44)
29.48 (13.40)
13.00 ( 5.91)
19.62 ( 8.92)
25.65(11.66)
26.40 (12.00)
40.96- 54.16(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(53.00-60.00)
109.7 -114.7(49.88-52.13)
110.3 -159.2(50.13-72.38)
ELECTROLYTIC COPPERt
.
-
-
.
.
.
-
-
.
-
-
.
42.90- 53.90(19.50-24.50)
69.30 - 94.60 (31.50 - 43.00)
66.00 - (30.00)
77.00- 81.40(35.00-37.00)
116.9 -132.3(53.12-60.12)
111.4 -115.8(50.63-52.63)
116.9 - 151.1 (53.13 - 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
50
-------
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
51
-------
thousand metric tons of ore per day. Lead and zinc ores are
produced almost exclusively from underground mines. There
are some deposits which are amenable to open pit operations;
a number of mines during their early opening stages of
operation are started as open-pit mines and then developed
into underground mines. At present, only one small open-pit
mine is in operation, and its useful life is estimated in
months. Therefore, for all practical purposes, all mining
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
conversion 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 sulfide). 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 zincite (zinc oxide),
willemite (zinc silicate), and franklinite (an iron, zinc,
manganese oxide complex). Sphalerite is often found in
association with sulfides of iron and lead. Other elements
often found in association 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
52
-------
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.-DEC.
METRIC TONS
5
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
251,571
51
1,078
1,331
5,674
1,359
443
596
486
300,163
1,658
SHORT TONS
...
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.
53
-------
TABLE 111-14. MINE PRODUCTION OF RECOVERABLE ZINC
IN THE UNITED STATES (PRELIMINARY)
STATE
Arizona
California
Colorado
Idaho
Illinois
Kentucky
Maine
Missouri
Montana
New Jersey
New Mexico
New York
Pennsylvania
Tennessee
Utah
Virginia
Washington
Wisconsin
1973
RANK
4
5
7
1
6
2
3
9
8
%
11.94
9.55
4.13
17.27
6.94
17.4
13.32
3.48
3.51
Total
Daily average*
1973
JAN.-DEC. TOTALS
METRIC TONS
7,638
16
51^33
41,216
4,823
245
17343
74,576
379
29355
11,147
73361
17,104
57,474
15,023
15,131
5,768
7365
431,599
1,183
SHORT TONS
8,421
18
56317
45,442
5,318
270
19,672
82,223
418
33,027
12,290
81,435
18358
63,367
16364
16,682
6,359
8,672
475,853
1,304
1
1974
JAN. TOTALS
METRIC TONS
600
_.
3,961
3,279
224
_.
1,238
6^89
82
2,361
863
6,961
1,575
7,239
1,130
1,281
528
733
38,644
1,246
SHORT TONS
662
4,367
3,615
247
1,365
7,265
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 adjustment for Sundays or holidays.
54
-------
preliminary 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 area 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, 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
mineralcontaining 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 skips.
Primary or rough crushing equipment is often operated
underground. The drilling and transportation equipment is,
of course, highly mechanized and employs the diesel power.
At some locations, the equipment is maintained in
underground shops, constructed in mined-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 locations.
55
-------
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
GRINDING AND
CLASSIFICATION
LEAD ROUGHER
LEAD FLOTATION
ZINC ROUGHER
FINAL LEAD
CONCENTRATE
ZINC ROUGHER
CONCENTRATE
THICKENING
AND FILTRATION
USUALLY
RECYCLED
TO PROCESS
WATER SYSTEM
AND FILTRATION
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS
EXCESS REAGENTS
TO SUBSURFACE
DRAINAGE
CONCENTRATE
TO ZINC
SMELTER
USUALLY RECYCLED
TO PROCESS
WATER SYSTEM
56
-------
water enters with sand or tailings used in hydraulic
backfill operations.
The water is pumped from the mine at a rate necessary to
maintain operations in the mine. The amount of water pumped
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 million liters (30 to 40
million gallons) per day. In many cases, there is a sub-
stantial seasonal variation in the amount of water which
which must be pumped.
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 preconcentrated using
mechanical devices based on specific gravity principles.
The ore is initially crushed to a size suitable for
introduction into fine grinding equipment, such as rod mills
and ball mills. These mills run wet and are usually run in
circuit with rake or cyclone classifers 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 (CaOH2) are common. Other
reagents commonly used and their purposes are:
57
-------
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 Xanthate Collector
Dixanthogen Collector
Isopropyl Ethyl Thionocarbonate Collectors
Sodium Diethyl-dithiophosphate Collectors
Zinc Sulfate Zinc Depressant
Sodium Cyanide Zinc Depressant
Copper Sulfate Zinc Activant
Sodium Dichromate 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 hydrophobic (non water accepting) 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.
58
-------
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 tailings from a lead/zinc flotation mill contains the
residual solids from the original ore which have been finely
ground to allow mineral recovery. The tailings also
contains dissolved solids and excess mill reagents. 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
industry is concentrated in eight states: Alaska, Montana,
New Mexico, Arizona, Utah, Colorado, Nevada, and South
Dakota. 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 either
underground or open-pit methods, the particular method
chosen depending on such factors as size and shape of the
59
-------
deposit, ore grade, physical and mineralogical character of
the ore and surrounding rock, and depth of the deposit.
Milling Practices. Milling practices for the processing and
recovery of gold and gold-containing ores are cyanidation
amalgamation, flotation, and gravity concentration All
these processes have been employed in the beneficiation of
ore mined from lode deposits. Placer operations, however,
employ only gravity methods, sometimes in connunction with
amalgamation.
Prior to 1970, amalgamation was the process used to recover
nearly 1/4 of the gold produced domestically. Since that
txme, 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 0.395 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 domestically was recovered by cyanidation, and this
increased to 54.695 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 g9ld/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
60
-------
Figure 111-9. CYANIDATION OF GOLD ORE: VAT LEACHING OF SANDS
AND 'CARBON-IN-PULP' PROCESSING OF SLIMES
ORE
TO REFINERY
TO SMELTER
61
-------
agitation 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
SiSkeners? 10n ±S C°llected *» Banting frU, the
Cyanidation of slimes generated in the course of wet
±SK cufrently bei"9 done by a recently developed
carbon-in-pulp (FigUre II:[-9). The slimes are
a^CyanXdS solution in large tanks, and the
gold cyanide is collected by adsorption onto
activated charcoal. Gold is stripped7 from the charcSa?
using a small volume of hot caustic; an electrowinning
process is used for final recovery of the gold in the SI?
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 go Id- containing minerals (Figure III-
11). The single operation currently using this method
further processes the tailings from the flotation circuit bv
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 trov
ounces), over 38% of this production came from Idaho, and
most of this, from the rich Coeur d'Alene district in the
Idaho panhandle. 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
62
-------
Figure 111-10. CYANIDATION OF GOLD ORE: AGITATION/LEACH PROCESS
ORE
J_
CRUSHING
GRINDING
CONDITIONING
COUNTERCURRENT
LEACHING IN
THICKENERS
PRECIPITATION
OF GOLD FROM
LEACHATE BY
ADDITION OF
ZINC DUST
I
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
1
PRECIPITATE
FILTERED AND
THICKENED
TO SMELTER
REAGENTS (CN)
BARREN
PULP
TAILING
POND
TAILING-POND
DECANT
RECYCLED
BARREN SOLUTION
RECYCLED
J
63
-------
Figure 111-11. FLOTATION OF GOLD-CONTAINING MINERALS WITH RECOVERY OF
RESIDUAL GOLD VALUES BY CYAIMIDATION
ORE
CRUSHING
GRINDING
CONDITIONING
i
SELECTIVE
FROTH
FLOTATION
CONCENTRATE
FILTERED
AND THICKENED
TO SMELTER
FLOTATION CIRCUIT
TAILINGS
-fr
REAGENTS (CN)
LEACHING IN
THICKNERS
.BARREN
PULP
TO TAILING
POND
PRECIPITATION OF GOLD
FROM LEACHATE BY
ADDITION OF ZINC DUST
I
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
PRECIPITATE FILTERED
AND THICKENED
T
TO SMELTER
64
-------
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 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 silver is produced are listed in Table III-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 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
liberated—gold/silver ores, processed by amalgamation and
cyanidation, now contribute only 1 percent of the domestic
silver produced. Primary sulfide ores, processed by
flotation and smelting, account for 99 percent (Table III-
16).
Selective froth flotation processing can effectively and
efficiently beneficiate almost any type and grade of sulfide
ore. This process employs various well-developed reagent
combinations 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, promoters,
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)12Sb5S13. A tetrahedrite concentrate contains
approximately 25 to 32% copper in addition to the 25.72 to
65
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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
66
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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
SILVER BULLION AND PRECIPITATES RECOVERABLE BY
AMALGAMATION
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
YEAR
1968
1969
1970
1971
1972
SILVER RECOVERABLE FROM ALL SOURCES (%)
AMALGAMATION
0.28
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
*Less than 1/2 unit
SOURCE: REFERENCE 2
67
-------
Figure 111-12. RECOVERY OF SILVER SULFIDE ORE BY FROTH FLOTATION
ORE
NO. 1
FLOTATION CIRCUIT
NO. 2
FLOTATION CIRCUIT
RETREATMENT
CIRCUIT
PYRITE
CONCENTRATE
NO. 3
FLOTATION CIRCUIT
I
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
FINAL PYRITE
CONCENTRATET
CONTAINS 3.43 KILOGRAMS PErt
METRIC TON (100 TROY OUNCES
PER SHORT TON)
68
-------
44.58 kilograms per metric ton (750 to 1300 troy ounce per
ton) of silver. A low-grade (3.43 kg per metric ton; 100
troy oz per ton) silver/pyrite concentrate is produced at
one mill. Antimony may comprise up to 18% of the
tetrahedrite concentrate and may or may not be extracted
prior to shipment to a smelter.
Various other silver-containing minerals are recovered as
byproducts of primary copper, lead, and/or zinc operations.
Where this occurs, the usual practice is to ultimately
recover the silver from the base-metal flotation
concentrates at the smelter or refinery.
Less than 1 percent of the current domestic production of
silver is recovered by amalgamation or cyanidation
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
(A12O_3) 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 chemical 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 mined 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 unconsolidated sands and clays, but some
slumping does occur.
69
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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 crusSng? ore
blending, and grinding in preparation for refining. in
1972, less than 10 percent of the bauxite used for primarv
aluminum production was of domestic origin with the
increasing 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
rS?i"rC:Lent P° meet ?reS6nt needS °f the nine domestic
™i? rieS'.^ Re°ent 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(OH)12(SO«)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. Additionally, the processing of alunite
creates no significant "red mud" (leach residue). Currently
alunite mining is in the exploratory stages, with a
commercial scale refinery 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
nJS"^ 5r°m % PSak 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
70
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TABLE 111-17. PRODUCTION OF BAUXITE IN THE UNITED STATES
YEAR
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
71
-------
these operations is based on the use of a portion of their
end product in the production of ferroalloys (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 the
past and are included in present reserves. Beneficiation
techniques include numerous gravity processes, jigging,
tabling, sink-float, Humphreys spirals; flotation, both
basic-sulfide and fatty-acid; and a variety of ore leaching
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. Production figures for the
ferroalloy mining and milling industry since 1945 are
summarized in Table III-18.
72
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TABLE 111-18. PRODUCTION OF FERROALLOYS BY
U.S. MINING AND MILLING INDUSTRY
COMMODITY
Chromium
Columbium 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
(15,150)
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
tt
Reference 5
73
-------
A~ ^ "T.I-18 shows, molybdenum mining and milling
constxtute the largest and most stable segment of the
ferroalloy ore mining and milling industry in the United
States. The U.S. produces over 85% of the world's
:uc 'bde-ium supply, with two mines doir.nating the industry.
, L c ,vio mines are among the 25 largest mining operations
in the U.S. Production is expected to increase in the near
future with expanded output from existing facilities, and at
least one major new operation in Colorado is expected to be
in operation soon.
The only commercially important ore of molybdenum is
molybdenite, MoS2. It is mined by both open-pit and under-
ground methods and is universally concentrated by flotation.
Commercially exploited ore currently ranges from 0.1 to 0.3
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 concentrate); with the sale of some
stockpiled material, only about 50 mines operated with a
price of $43 per unit (Reference 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
(CaWOU_) and the wolframite series, wolframite ((Fe, Mn)W04J,
ferberite (FeWOjf) , and huebnerite (MnWO^. Underground
mining predominates, 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
74
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about 0.6 percent tungsten, and concentrates containing
about 70 percent W0_3 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.
Manganese 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.
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 deoxidizer, 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
75
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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 carbamate leaching
(Reference 13). It is most likely that heavy-media
separation will be practiced in the immediate future.
Nickel ores are not currently being exploited in the U. S.
One nickel lateritic lateritic deposit is currently being
mined. Some sulfide nickel ore deposits with commercial
possibilities have been found in Alaska (Reference 2). If
they are developed, 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).
There are no mines or mills currently active in the U.S.
producing ores or concentrates of chromium, cobalt, colum-
bium, 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, as 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.
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
76
-------
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
influence on the cobalt market (Reference 5). Known
domestic ore from which cobalt might be recovered is a
complex copper 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. Concentration techniques varying widely from
fairly simple gravity and hand picking techniques through
magnetic and electrostatic 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 particular, hydrometallurgical processes
like those used in vanadium 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 metals may be beneficiated in
77
-------
the same type of process used for vanadium. There is also
some similarity to tungsten, molybdenum, and chromium.
Ferroalloy Ore Beneficiation Processes
Ore processing in the ferroalloys industry varies widely,
and even ores bearing the same ore mineral may be
concentrated by widely differing techniques. There is thus
no scheelite recovery process or pyrolusite concentration
technique per se. 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 functions (such as crushing
and grinding ore) that are common to nearly all ore
concentration procedures. Fundamental ore beneficiation
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
beneficiation 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, Humphreys spirals, and
tabling. Jigging is applicable to fairly coarse ore,
ranging in size from 1 mm to 13 mm (approximately O.OU 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.
78
-------
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 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
application 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 perfect, 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 O.OOU 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 cassiterite,
columbite, and monazite (Reference 5 ) or the separation of
cassiterite and wolframite (Reference 18). Although they
79
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are both most frequently implemented as dry processes, wet-
belt magnetic separators are used. Since ore particles are
charged to 20,000 to 40,000 volts for electrostatic
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 grinding 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 MoS2^
concentrate from 0.3 percent) (Reference 18 ). its
specificity also allows separation of different ore minerals
(e.g., CuS and MoS_2) where desired, and operation with
minimum reagent consumption since reagent interaction is
typically only with the particular materials to be floated
or depressed.
80
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Figure 111-13. GRAVITY-PLANT FLOWSHEET FOR NIGERIAN COLUMBITE
-OVERFLOW-
-OVERFLOW-
f T°
WASTE
— »J
15-cm (6-in.)
GRAVEL
PUMP
CYCLONES
{OPTIONALI
*
4
SHAKING TABLES
- OVERFLOW -
•MIDDLINGS
-MIDDUNGS-
I
-HEADS-
- OVERSIZE-
(BLED PERIODICALLY)-
SOURCE: REFERENCE 19
81
-------
Figure 111-14. EUXENITE/COLUMBITE BENEFICIATION-PLANT FLOWSHEET
J DREDGE
1
HEAVY MINERAL
CONCENTRATE
t
[_ STORAGE
1
o
L o^cc™ j -q nv,^™,i.L| -T """""=""= [""^STORAGE
1 | * ^ WASTE
TO WASTE-* — SLIME
TO -*- QUARTZ «
STORAGE ^^
STORAGE"^" GARNET «
WASTE -*-TAILINGS-*
-4 CLASSIFIER L>JATTRITIC
t
INDUCED-ROLL
MAGNETIC SEPARATOR
t_
)NER ^ SEP A^ATO^I -MCLASSIFIER ~»- DRYER
*
MAGNETIC SEPARATOR I
(LOW-INTENSITY)
1
| SCREEN • - . n
INDUCED-ROLL
MAGNETIC SEPARATOR
t | 1
r NONCONDUCT
ELECTROSTATIC
SEPARATOR
1
NONCONDUCTORS
. INDUCED-ROLL
MAGNETIC SEPARATOR
T
. CROSSBELT MAGNETIC
SEPARATOR
1 1 MID
NONMAGNETICS
SCREEN J
[ FURNACE |
ELECTROSTATIC
SEPARATOR
ORS 1 1 — CONDUCTORS 1
[ SCREEN |
< 28 MESH '
CROSSBELT MAGNETIC
SEPARATOR
DLINGS *-
r
CROSSBELT MAGNETIC ' I 1
SEPARATOR ' ILMENITE f—
> 35 MESH < 35 MESH 1 NONMAGNETICS
4 X NONMAGNETICS I
T T ALTERNATE | _ f
FEE
| WET TABLE j DRIED CONC
r
AIR TABLE
*
CONCENTRATE
.ENTRATE M
CROSSBELT MAGNETIC 1 NONMAGNETICS
SEPARATOR J AND MIDDLINGS *"^
I" — ' L- 1
fc TO
STORAGE
^_ TO
STORAGE
| MONAZITE | | EUXENITE 1 COLUMBITE
TO STORAGE
82
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Figure 111-15. REPRESENTATIVE FLOW SHEET FOR SIMPLE GRAVITY MILL
MINING
ORE
TO
WASTE
TAILS -
GRINDING AND
CRUSHING
I
SCREENING
FINE
MIDDLINGS
COARSE
- HEADS
•TAILS-
MIDDLINGS
SHAKING
TABLES
I
• CONCENTRATE
83
-------
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, activators, 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 (C2H50 . NaCS2).
Frothers are generally organics with a soluble hydroxyl
group and a "non-wettable" hydrocarbon (Reference 17 ).
Pine oil (C6H12OH), 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
111-16. Water use in this operation currently amounts to
approximately 1.8 tons of water per ton of ore processed,
essentially all of which is process water. Reclaimed water
from thickeners at the mill site (shown on the flowsheet)
amounts to only 10 percent of total use.
Where byproducts are recovered with molybdenite, a somewhat
more complex mill flowsheet results, although the
molybdenite recovery circuits themselves remain quite
similar. A very simplified flow diagram for such an opera-
84
-------
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
Nokes 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 amines
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 50*
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
85
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Figure 111-16. SIMPLIFIED MOLYBDENUM MILL FLOWSHEET
MINING
ORE
I
CRUSHINGS,
WEIGHING, AND
SCREENING
BALL
MILLS
CYCLONES
OVERFLOW
ROUGHER FLOAT
MIDDLINGS
MIDDLINGS —
UNDERFLOW
-CONCENTRATE-
SCAVENGER
FLOAT
(4 STAGES WITH
REGRIND AND
INTERNAL RECYCLEI
OVERFLOW
—]
-UNDERFLOW-
OVERFLOW
-RECLAIM —
WATER
TO TAILINGS
POND
— CONCENTRATE
MIDDLINGS -
CLEANER
FLOAT
(6 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
— TAILS-
TAILS
_*_
CONCENTRATE
CYCLONES
[ -
UNDERFLOW -
MOLYBDENUM
PRODUCT
86
-------
tion is shown in Figure III-17. Pyrite flotation and
monazite flotation are accomplished at acid pH (U.5 and 1.5,
respectively), somewhat increasing the likelihood of solu-
bilizing 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 flotattion from complex ores in processes
involving multiple flotation steps. Some of the most likely
reagents to be used in these operations are presented in
Table III-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 processes
to be, in general, somewhat more sensitive to feedwater
conditions than sulfide floats; consequently, they are less
frequently run with recycled water.
In current U.S. operations, scheelite is recovered by
flotation using fatty acids as collectors. A typical 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/short ton), and sodium oleate (0.1 to 0.2 kg/metric ton,
or 0.2 to O.U Ib/short ton). In addition, materials such as
copper sulfate or acid dichromate may be used in small to
moderate amounts as conditioners and gangue depressants.
Scheelite flotation circuits may run alkaline or acid,
depending primarily on the accessory mineralization in the
87
-------
Figure 111-17. SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM
' CONCENTRATE
LIGHT TO TAILS
MONAZITE
CONCENTRATE
TO TAILS
CRUSHING
(3 STAGES)
28% + 3 MESH
1
GRINDING
BALL MILLS
36% + 100 MESH
^l
+
FLOTATION
t
FLOTATION
I
96% OF MILL FEED
*
GRAVITY
HUMPHREY'S SPIRALS
*
PYRITE
FLOTATION
1
TAILS
*
TABLES
*
MONAZITE
FLOTATION
1
MAGNETIC
SEPARATION
1
NONMAGNETIC
TIN CONCENTRATE i r
1
CONCENTRATE
1
CLEANER
FLOTATION
(4 STAGES)
1
DRYING
*
MOLYBDENUM
CONCENTRATE
(93% + MoS2)
-TAILINGS-
MAGNETIC TUNGSTEN
CONCENTRATE
88
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TABLE 111-20. PROBABLE REAGENTS USED IN FLOTATION OF
NICKEL AND COBALT ORES
Lime
Amyl Xanthate
Isopropyl Xanthate
Pine Oil
Methyl Isobutyl Carbinol
Triethoxybutane
Dextrin
Sodium Cyanide
Copper Sulfate
Sodium Silicate
89
-------
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 IH-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. Approximately four tons of water are
used per ton of ore processed in this operation.
The basic flotation operations for manganese ores and colum-
bium 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 lb/short ton)
Soap skimmings no kg/metric ton
(80 lb/short ton)
Oronite S (wetting agent) 5 kg/metric ton
(10 Ib/short ton)
sol 5 kg/me trie 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 cleanina
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
separate the desired concentrates. This is illustrated in
the flowsheet for a Canadian pyrochlore (NaCaCb206F) mill in
Figure III-19.
90
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Figure 111-18. SIMPLIFIED FLOW DIAGRAM FOR SMALL TUNGSTEN CONCENTRATOR
ORE
i
SULFIDE
FLOTATION
CYCLONE
75% SANDS
i
GRAVITY
TABLES
TAILINGS
25%
SLIMES
OVERFLOW
THICKENER
SCHEELITE
FLOTATION
HCI LEACH
(15 TO 20% OF
FRACTION)
TUNGSTEN
CONCENTRATE
91
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Figure 111-19. MILL FLOWSHEET FOR A CANADIAN COLUMBIUM OPERATION
CONCENTRATE
~L
MAGNETIC
SEPARATION
DESLIMING
UNDERFLOW
I
BULK FLOTATION
CONCENTRATE
DESLIMING
UNDERFLOW
MAGNETIC
SEPARATION
PRIMARY CLEANING
(FIRST STAGE)
CONCENTRATE
PRIMARY CLEANING
(SECOND STAGE)
CONCENTRATE
SECONDARY CLEANING
(THREE STAGES)
CONCENTRATE
I
~l
- TAILS -
REVERSE
FLOTATION
— CONCENTRATE -
MAGNETITE
TO
STOCKPILE
—. OVERFLOW
I (CALCITE)
SPIRALS
TAILS
TABLING
CONCENTRATE
CONCENTRATE
SOURCE:REFERENCE 5
92
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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 legs easily beneficiated. A number of leaching
processes have been developed for manganese ores in the
search for methods of exploiting plentiful, low-grade,
difficult-toconcentrate 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
chromite 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 welldocumented 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% V205, and
some Cr, Ni, and Mo — yields vanadium as a byproduct of
phosphate fertilizer production. Ferrophosphate is first
prepared by smelting a charge 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 (Na^CO_3) 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 metavanadate, and sodium
93
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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 V205 (redcake) by acidification, and chromium
is precipitated as lead chromate. By this process, 85
percent of vanadium, 65 percent of chromium, and 91 percent
phosphorus can be extracted.
Another, basically non-radioactive, vanadium ore, with a
grade of 1 percent V205, is found in a vanidiferous, mixed-
layer montmorillonite/illite and goethite/montroseite
matrix. This ore is opened up by salt roasting, following
extrusion of pellets, to yield sodium metavanadate, which is
concentrated 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 carbamate 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 III-21).
During this time, the potential environmental problem and
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
Insectide, Fungicide, and Rodenticide Act. in addition,
registration has been suspended for mercury alkyl compounds
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
94
-------
Figure 111-20. FLOWSHEET OF TRISTAGE CRYSTALLIZATION PROCESS FOR
RECOVERY OF VANADIUM, PHOSPHORUS, AND CHROMIUM
FROM WESTERN FERROPHOSPHORUS
FERROPHOSPMORUSj
SETTLING AND OECANTATION
SOLIDS
WATER WASH
PRIMARY CRYSTALLIZATION
PRIMARY CENTRIFUGING
WASH
LIQUOR
PREGNANT
SOLUTION
CRYSTALS
-DISSOLUTION-
J_
| SECONDARY CRYSTALLIZATION [
RED CAKE MOTHER LIQUOR
[ FUSION j |
| FILTRATION |
VANADIUM PRODUCT
Ci«
T
WA
TO
• ^STOCKPlLC
3HI2 SOLUTION mnnf
j i
[ FILTRATION J
1 \ ^_ SOLUTION
I TO WASTF
0 CHROMIUM ^ TO
>TE PRODUCT ' STOCKPILE
-------
Figure 111-21. ARKANSAS VANADIUM PROCESS FLOWSHEET
1.5 -2.0% V2O5
t
6-10%
NaCL
TERTIARY
AMINES
GRINDING
PELLETIZING
ROASTING
850°C (1562°F)
NaVO,
£
H2S04
•
LEACHING AND
ACIDIFICATION
pH 2.5 - 3.5
(Na6V10028
SODIUM DECAVANADATE)
I
SOLVENT EXTRACTION
NH4OH
PRECIPITATION
I
NH4V03
PRECIPITATE
I
CALCINING
V2O5 PRODUCT
96
-------
Figure 111-22. FLOWSHEET OF DEAN-LEUTE AMMONIUM CARBAMATE PROCESS
RAW SIZED ORE -- 1.9 cm (0.75 in.)
REDUCING FURNACE
-CO,
i
GRINDING (30 MESH)
WEAK Mn
SOLUTION
i
LEACHING
(TWO 11.356- i (3000-GAL))
REACTION TANKS
IN SERIES
I
9.14-m (30-ft) CLARIFYING
THICKENER
7.6-m (25-ft) COUNTERCURRENT
WASHING THICKENERS
LIVE
STEAM
SLURRY
STILL
NH4NH2CO2
TAILINGS
NEW LEACH LIQUOR
LEACH LIQUOR
REGENERATION
TW011.356-X
(3000-GAL)
PRECIPITATION
TANKS (IN SERIES)
-NH,
T r
MnCO,
CLARIFYING
THICKENER
MOTHER
" LIQUOR "
70% SOLIDS
ROTARY DRYER
NH4NH2CO2
PRODUCT
97
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TABLE 111-21. DOMESTIC MERCURY PRODUCTION STATISTICS
CATEGORY
Production in metric tons
(flasks)
Dollar value (thousands)
YEAR
1969
1.029
(29,640)
$14,969
1970
948
(27,296)
$11,130
1971
56
621
(17,883)
$ 5,229
21
253
(7,286)
$1,590
6
SOURCE: REFERENCE 2
98
-------
growth in the consumption of mercury is anticipated for
electrical apparatus, instruments, and dental supplies.
From consideration 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 production 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 operation scheduled to begin in Nevada later in
1975 will concentrate mercury ore by flotation. This
concentrate will be furnaced, and annual production of
mercury from the operation is expected to reach 20,000
flasks (695 metric tons, or 765 short tons).
Uranium, Radium, and Vanadium Ores
The mining and milling of uranium, vanadium, and radium
constitute one industry segment, because uranium and
vanadium are sometimes found in the same ore and because
radium, resulting from the radioactive decay of uranium, has
99
-------
always been obtained from uranium ores. In the oast 20
years, the demand for radium has diminished as radioactive
isotopes (e.g.. Co 60, Pu 239) with tailored characteristics
as sources of radiation have become available. Radium is
orTmarltv for a P°llutant in the »*stes. Uranium is mLel
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 Is a
ferroalloying metal and, in the form of its oxide, as a
S?SiS A • V^adium 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) oxidation
state is easily oxidized and the resulting uranium (VI), o-
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
asnh^J10no With-.Karb°naCeOUS fossils- i.e., lignite and
asphalts, ores with a grade in excess of a fraction of a
oS^be uranium are rare (80% of the industry operates with
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
tnat yield the concentrate, and both processes normally
share a waste water 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 auite
profitable. ^
The uranium oxide concentrate, whose grade is usually quoted
in percent of U308 (although that oxide figures in the
assay, rather than in the product), is generated by one of
several hydrometallurgical processes. For purposes of waste
water 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
100
-------
bicarbonate, depending on the content of acid-
wasting limestone in the gangue.
(2) Values in the leachate are usually concentrated by
ion exchange (IX) or by solvent extraction (SX) .
They are then precipitated as the concentrate,
yellowcake.
Some vanadium finds are not associated with significant
uranium concentrations. 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 III-22.
Primary deposits of uranium ore contain uraninite, the U(IV)
compound UOJ2, 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 UO^ and UO^. The composition U.30§ — i.e., U02.2U03
is particularly stable. The process occasionally stops
at gummite (U0_2. H20) , an orange or red, waxy mineral, but
usually involves further oxidation and reactions with
alkaline and alkaline-earth oxides, silicates, and
phosphates. The transport leads to the surface uranium ores
of arid lands, including carnotite (K2(U02)2(VOU)2.3H20),
uranophane (CaU2Si2011.7H£0) , and autunite
(Ca(U02J2.(PO^)_2.10-12H20) 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.
-------
TABLE 111-22. ISOTOPIC ABUNDANCE OF URANIUM
ISOTOPE
U238
U 235
U234
U236
HALF-LIFE (YEARS)
451 x 109
7.13 x 108
2.48 x 105
2.39 x 107
ABUNDANCE
99.27%
0.72%
0.0057%
Traces Identified
(Moon-1972; Earth-1974)
102
-------
A reducing environment is often provided by decaying
biological 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
encountered 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 (200-ft) grid,
and development drilling, on a 15-m (50-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 price and the possibility of increased uranium
demand due to the current energy situation have resulted in
the mining, for storage, of ore below this threshold 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 (JAN. 1, 1975)
U3O8 No. of Known % of total
(Short Torts)
New Mexico 137,108 66 69
Wyoming 28,300 14 14
Utah and
Colorado 11,400 99 5
Texas 14,400 45 7
Others 8,800 60 5
The number of separate known deposits in the western United
States is 284, 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
103
-------
Texas gulf coast, dominate the industry. in 1974
Mexico provided 43 percent and Wyoming 32 percent of
production. In 1974, the U. S. production was 7.1
tons of ore with a U3O8 equivalent of 12,600 tons.
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
r££f£ X t0° XKW f°5 economic recovery of uranium, which is
recovered as a byproduct only in Florida. Vanadium, in ores
r2aJ ^ co?tain Cranium 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
?«r^?S Sf * ?anuary 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
shortlived, 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 bv
leaching and ion exchange.
Because of the small size of pockets of high-grade ore,
openpit 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 stockpiles. on January 1974, 33 open "pit
mines were being worked, and 20 other (e.g., heap-leachinq)
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.
Milling. Mills range in ore processing capacity from 450
"^cn J°nS (495 Short t0ns> 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 111-23 and
are tabulated by company in Supplement B.
104
-------
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
Kin nc MII i c
3
7
3
2
3
1
1
20
105
-------
CJu_§hirK[, ^nd Roasting. Ore from the mine tends
c& be quite viriaJle 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 "alu j the proper source
?nd to rchieve uniform ~r-\er and .*• blending to a:iire
aniform 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 to convert insoluble heavy-metal
vanadates (vanadium complex) and carnotite to more soluble
sodium vanadate, which is then extracted with water. 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 .07 mm (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 (CaCO_3) 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 U(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 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:
2U02: + 02 > 2UOJ3
2W)3 + 2H2SOU. + 5H_2O > 2 (UO2SOJ4) . 7H2O
106
-------
Uranyl sulfate (UO2SO4) forms a complex, hydrouranyl tri-
sulfuric acid (H^UO2(SOJ4) .3) r 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 raising the pH
with the addition of sodium hydroxide. The supernatant 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 central 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:
2UO2 + 02 > 2U03
(oxidation)
3Na2(CO.3) + UO3 + H2O > 2NaOH + Na4_ (U02J (C03)3
(leaching)
2NaOH + CO2 > Na2C0.3 + H20
(recarbonization)
2Na4JU Na.2U2O7 + 6Na2_CO3_ + 3H20
(precipitation)
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 waste water but are often ignored.
Purification to remove sodium ion is necessary both to meet
the specifications of American uranium processors and for
107
-------
Figure 111-23. PACHUCA TANK FOR ALKALINE LEACHING
LEACH
AIRLIFT
=////=:////=
108
-------
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., Na2lJ2p2 = Na_2O +
2UO3I.
The process of generating UO2: 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 UO3_
component to UO_2 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 UOJ and Na_2O 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 U Na2SO^ + 3H20 + 2 (UQ2) 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.
UO2SCW + H.2O + 2NH3_ > (NHU) .23(34 + UOJ3
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 an
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 ;o fer-
tilize alkaline southwestern soils. Ammonium sulfate could
be generated by neutralizing the wastes of the 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
109
-------
ammonium ion, present a present health hazard more severe
than that from sulfate ion.
Vanadium Recovery. Vanadium, found in carnotite CK2(\3O2)2
(VO4J 2 . 3H2!O) as well as in heavy metal vanadates—e.g.,
vanadinite (9PbO . 3V2O_5 . PbCl_)—is converted to sodium
orthovanadate (Na^SVO^) , which is water-soluble by roasting
with sodium chloride or soda ash (Na2CO3). After water
leaching, ammonium chloride is added, and poorly soluble
ammonium vanadates are precipitated:
NaSVOU + 3NH4C1 + H2O > 3NaOH +NIMVO3 + 2NH4OH
(ammonium metavanadate)
Na3VO4 * 3NH4C1 > 3NaCl + (NHU)3VOi»
(ammonium orthovanadate)
The ammonium vanadates are thermally decomposed to yield
vanadium pentoxide:
2(NH!)!VQ1 > 6NH3 + 3H2O + V2O5
A significant fraction (86 to 87%) of V2O5 is used in the
ferroalloys industry. There, ferrovanadium has been
prepared in electric furnaces by the reaction:
V2O5 + Fe_2O_3 + 8C > SCO + 2FeV
or by aluminothermic reduction (See Glossary) in the
presence of scrap iron.
Air pollution problems associated with the salt roasting
process have led many operators to a hydrometallurgical
process of vanadium recovery that is quite similar to
uranium recovery by acid leaching and solvent exchange. The
remainder of V2O5 production is used in the inorganic
chemical industry, and its processing is not within the
scope of these guidelines. Since the mining and
beneficiation of vanadium ores not containing 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.
110
-------
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 precipitated 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
concentrations in the eluate of some of these processes are
shown in Table III-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-
24a) . 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 III-2Ub,
c). Sands are often repulped, by the addition of some waste
water stream or another, to facilitate flow to the tailing
pond as much as a few kilometers away. Consequently, there
is some latitude for the selection of the waste water 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-, N0_3-r HSO_4-, C0_3— (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
111
-------
Figure 111-24. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 1 of 2)
FROM •
LEACH
PREGNANT
LEACH LIQUOR
SLIMES
CLEAR LEACH LIQUOR
TO COLUMN IX OR SX
a) LIQUID/SOLID SEPARATION
SLIMY PULP TO
RESIN-IN-PULP IX
SAND
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
112
-------
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
PREGNANT
LEACH
LIQUOR
1
°0
Oo
LOADING
SOLVENT
LOADED
ORGANIC
CD-
BARREN STRIPPED
ELUANT SOLVENT
SOLVENT
BARREN
LIQUOR
»
PHASE
SEPARATION STRIPPING
e) SOLVENT EXTRACTION
Vy ^PREGNANT
X-X ELUATE
If
PHASE
SEPARATION
PRECIPITATION
f) ELUEX PROCESS
RECYCLE
BARREN
ELUANT
IX
ELUANT
~\ f ^
1 1
/ \
IX
PREGNANT
ELUATE
PARTIALLY STORAGE)
STRIPPED \ / LOADED
RESIN ^ -~ RESIN
g) SPLIT ELUTION
113
-------
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
1.2 to 1.6
IX/SX combination
Eluex process
3.0 to 7.5
114
-------
brittle, radioactive, and magnetic, permitting concentration
by magnetic means. There are some deposits of consolidated
monazite sands in Wyoming.
Hydrometallurgical 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 III-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 lb) of monazite sulfate and is not very
economical. When used in conjunction 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
environmental 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
uranium diuranate in the 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. 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 other subcategories related to thorium ore may
generally apply.
115
-------
advantage of both the slime resistance of resin-in-pulp ion
exchange and the separatory efficiency of solvent exchange
(Eluex process). The uranium values are precipitated with a
base or a combination of base and 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 actinide (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 contain from 1 to 10 percent thoria (Tho;2) . American
ores from the North and South Carolinas, Florida, and Idaho
contain 1.2 to 7 percent ThO^, 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
concentration 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, cassiterite, or other
materials that concentrate in a similar way. Monazite is
116
-------
X be the mobile ion. Then, the exchange reaction for the
uranyltrisulfate complex is
URX + (U02 (S0.4) 3_) ---> R4U02(SOJi)l + 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+++ Mg+ + 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 anionic 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, 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, Figure
III-24d, arrives at the end of the column) are used to make
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
countercurrent 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
117
-------
Figure 111-25. SIMPLIFIED SCHEMATIC DIAGRAM OF SULFURIC ACID DIGESTION
OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
AND RARE EARTHS
MAIN STREAM
TO
STOCKPILE
RESIDUE
(UNDIGESTED MONAZITE SAND,
SIUICA, ZIRCON, AND RUTILE)
FILTRATE
(R.E., U, AND P205)
SELECTIVE
PRECIPITATION
AT pH 2.3
1
FILTRATE
*
PRECIPITATE OF
U AND P2OS
(BYPRODUCT)
TO WASTE
TO SHIPPING
MONAZITE
SAND
GRINDING
OPERATION
DIGESTION
DISSOLUTION
Th, R.E., U, AND P2O5
SELECTIVE
PRECIPITATION
AT pH 1.05
PRECIPITATE
(Th, R.E.. AND P2O5)
PURIFICATION BY SOLVENT EXTRACTION,
SELECTIVE PRECIPITATION, OR FRAC-
TIONAL CRYSTALLIZATION
CONCENTRATES
TO SHIPPING
SOURCE: REFERENCE 20
118
-------
Figure 111-26. SIMPLIFIED SCHEMATIC DIAGRAM OF CAUSTIC SODA DIGESTION
OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
AND RARE EARTHS
MAIN STREAM
MONAZITE
SAND
NaOH
H2O
I
FILTRATE
(NaOH AND Na3P04)
±
CRYSTALLIZATION
1
i
RESIDUE
(Na3P04)
(BY-PRODUCT)
1
FILTRATE
(RARE EARTHS)
SELECTIVE
PRECIPITATION
FILTRATE
i
1
GRINDING
OPERATION
DIGESTION
138°C
(280°F)
HYDROUS METAL-OXIDE CAKE
(Th, U, AND R.E.)
PRECIPITATE OF
RARE EARTHS
(BY-PRODUCT)
TO WASTE
SOURCE: REFERENCE 20
t
TO STOCKPILE
SELECTIVE
PRECIPITATION
PRECIPITATE
(Th AND U)
PURIFICATION BY
SOLVENT EXTRACTION
T
"— 1 TO
CONCENTRATES |-»" STOCKPILE
119
-------
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
Q
ui
<
oc
a.
UJ
D
O
IDAHO MONAZITE SAND
D INDIAN MONAZITE SAND
A
20 -
ACIDITY (pH)
AGITATION TIME: 5 MINUTES
DILUTION RATIO: H2O: SAND = 45:1 TO 50:1
DIGESTION RATIO: 93% H2SO4: SAND = 1.77
NEUTRALIZING AGENT: 3.1% NH4OH
SOURCE: REFERENCE 20
120
-------
Radiation parameters of thorium and uranium daughters are
somewhat different. The two decay series are compared in
Table III-25. The uranium series is dominated by radium,
which—with a halflife of 1620 years and chemical
characteristics that are distinctly different from those of
the actinides and lanthanides—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 being
Ra228 at 6.7 years.
Industry Flow Charts. Of the sixteen mills operating in
1967 (Table III-26), no two used identical leaching concen-
tration, and precipitation steps. The same was probably
true of the 15 mills operating in 1974 (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 production data (U.S.
Atomic Energy Commission, 1974) show that uranium production
has been relatively stable (between 12,600-14,000 ton U_3O8
per year) since 1968.
Table 111-27 shows uranium production for the period 1968
through 1972, expressed in terms of both ore movement and
UJOJ3 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-
tonium) will become available if and when environmental
problems of fuel recycling are resolved—particuarly, 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
differently, 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 illustrated in Table 111-29.
121
-------
TABLE 111-25. DECAY SERIES OF THORIUM AND URANIUM
ELEMENT OR
NAME
SYMBOL(S)
HALF-LIFE
ENERGY OF RADIATION
(MeV)
(X.
fl I r
Thorium Series
Thorium
Mesothorium 1
Mesothorium 2
Radiothorium
Thorium X
Thoron
Thorium A
Thorium B
Thorium C
Thorium C'
Thorium C"
Thorium D
90Th
88Ra228(MsTh1)
89Ac228 (MsTh2)
^Th228 (RdTh)
ggRa224 (ThX)
ggRn220 (Tn)
^Po216 (ThA)
82Pb212 (ThB)
83Bi212(ThC)
^Po^lThC')
81TI208 (ThC")
82Pb2°8(ThD)
1.34x 1010 years
6.7 years
6.13 hours
1.90 years
3.64 days
54.5 seconds
0.1 58 seconds
10.6 hours
60.5 min
3 x 10 second
3.1 minutes
Stable
4.20
-
4.5
5.42
5.68
6.28
6.77
-
6.05
8.77
-
-
-
0.053
1.55
-
-
-
ft
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
92u238 (uu
9QTh234(UX1)
91Pa234(UX2)
7*14.
92u234 (uio
90Th230(.o)
88Ra226
86Rn222
^Po218 (RaA)
82Pb214 (RaB)
O1A
83Bi2U (RaC)
MPo214 (RaC')
81TI210(RaC")
82Pb210 (RaD)
83Bi210(RaE)
84Po210 (RaF)
82Pb206 (RaC)
4.55 x 109 years
24.1 days
1.14 minutes
C
2.69 x 10° years
8.22 x 104 years
1600 years
3.825 days
3.05 minutes
26.8 minutes
19.7 minutes
1.5 xlO"4 second
1.32 minutes
22.2 years
4.97 days
139 days
Stable
4.21
-
-
4.75
4.66
4.79
5.49
5.99
-
5.50
7.68
-
-
-
5.30
-
-
0.13
2.32
—
-
-
-
ft
0.65
3.15
-
1.80
0.025
1.17
-
-
-
0.09
0.80
—
r
0.19
-
-
r
1.8
-
-
0.047
-
r
-
122
-------
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
Kerr-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 & 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 (H2O2)
VANADIUM RECOVERY
7
3
4
1
3
3
8
1
5
SOURCE: REFERENCE 21
123
-------
Figure 111-28. GENERALIZED FLOW DIAGRAM FOR PRODUCTION OF URANIUM
VANADIUM, AND RADIUM unMmum,
MINING
1
ORE TREATMENT
LEACHING
I
LIQUID/SOLID
SEPARATION
I"
I
ION EXCHANGE
SOLVENT EXTRACTION
PATH I
I
PATH IH
PATHn
PRECIPITATION
TO
STOCKPILE
1
•-H
URANIUM
CONCENTRATE
VANADIUM
BYPRODUCT
RECOVERY
TO
STOCKPILE
124
-------
TABLE 111-27. URANIUM PRODUCTION
YEAR
=
1968
1969
1970
1971
1972
1973
ORE MOVEMENT
1000
METRIC TONS
=====
5.861
5,367
5,749
5,708
5,834
6,152
1000
SHORT TONS
,
6,461
5,916
6,337
6,292
6,431
6,781
U308 PRODUCTION
1000
METRIC TONS
•
11.244
10.554
11.732
11.157
11.727
12.032
1000
SHORT TONS
•
12.394
11.634
12.932
12.298
12.927
13.263
U3O8 RESERVES*
1000
METRIC TONS
146
185
224
248
248
251
SHORT TONS
161
204
247
273
273
277
•At $18.000 per metric ton ($16,340 per short ton).
TABLE III-28. VANADIUM PRODUCTION
YEAR
1968
1969
1970
1971
1972
u.s. V2o5
PRODUCTION
1000
METRIC
TONS
5,590
5,369
5,085
4,812
4,771
1000
SHORT
TONS
6,192
5,918
5,605
5,304
5.259
%OF
WORLD
46
31
27
28
26
WORLD V2O5
PRODUCTION
1000
METRIC
TONS
12,119
16,892
18,337
16,883
18,135
1000
SHORT
TONS
13,359
18,620
20,213
18,610
19,990
V205 VALUE
PER
METRIC
TON
$3,910
$5,190
$7,216
$7,887
$6,941
PER
SHORT
TON
$3,547
$4.708
$6,546
$7,155
$6,297
TABLE III-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
METRIC
TONS
•'.•
4,084
172
43
453
1972
SHORT
TONS
4,502
190
47
499
%
86
4
1
9
125
-------
Radium is traded from foreign sources, but not mined, in
quantities of about 40 grams (or curies) (1.4 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 (reference 28).
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). The 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).
M§£ai 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
126
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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
5,263
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 LEADt
(ANTIMONY CONTENTI
METRIC TONS
1,179
1,065
542
751
468
SHORT TONS
1,300
1,174
598
828
516
includes product-on from antimony ores and concentrates and byproduct recovery from silver concentrates.
tByproduct produced at lead refineries in the United States.
127
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antimony content. Nearly all of this production can be
attributed to a single operation which is using a froth
flotation process to concentrate stibnite (Sb2S3) (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) .
Antimony is also contained in lead concentrates and is ulti-
mately recovered as a byproduct at lead smelters usually as
antimonial lead. 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
(Be3_Al_2(Si6O.18)) was produced in Colorado and South Dakota.
The largest domestic source of beryllium ore is a
bertrandite (Be4Si2O7 (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.
128
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Figure 111-29. BENEFICIATION OF ANTIMONY SULFIDE ORE BY FLOTATION
MINING
ORE
i
CRUSHING
GRINDING
1
CLASSIFICATION
ROUGHER
FLOTATION
FROTH
I
•TAILS
•TAILS-
SCAVENGER
FLOTATION
CLEANER
FLOTATION
FROTH
I
I
FROTH
TO
WASTE
FILTER
FILTRATE
1
THICKENER
WASTE
I
FINAL
CONCENTRATE
TO SHIPPING
129
-------
A sulfuric acid leach process is employed to recover
beryllium from the Spor Mountain bertrandrite. This is a
P^^1?^ary Process' however, and further details are
wxthheld. 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,
iridium, 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
platinum-group metals is principally as a byproduct of
SKSt*,??6^1??' 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 glaciofluvial outwash The
other domestic source of platinum is as a byproduct of
refining copper from porphyry and other copper deposits and
trom 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 (Reference 2) .
130
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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
538.6
560.8
532.2
TROY OUNCES
14,793
21.586
17,316
18,029
17,112
VALUE
$1,500,603
$2,094,607
$1,429,521
$1,359,675
$1,267,298
SOURCE: REFERENCE 2
131
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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 III-30.
In the Republic of South Africa, milling and beneficiation
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.
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 (CeFCO_3) and
monazite (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 rare-earth
132
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Figure 111-30. GRAVITY CONCENTRATION OF PLATINUM-GROUP METALS
DREDGE
(SCREENING,
JIGGING, AND
TABLING)
TABLING
MAGNETIC
SEPARATION
CMflOMITE/
MASMITITE
DRYING
SCREENING
SIZING
11
BLOWER
90% CONCENTRATE
(PLATINUM GROUP AND GOLD)
1
TO SHIPPING
TO
WASTE
TO
' SHIPPING
TO
WASTE
133
-------
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 (approximately 228,488
short tons) of bastnaesite ore mined and milled (Reference 2
). Monazite 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 118 metric tons (130 short tons) of
byproduct monazite 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 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 PEO.
Monzazite is recovered from heavy-mineral sands mined
primarily for their titanium content. Beneficiation of
monazite is by the wet-gravity, electrostatic, and magnetic
techniques discussed 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 presently no known exploitable 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. A small placer operation
began production in New Meixco in June 1973. 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-
134
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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 cassiterite 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.
Cassiterite concentrate recovery takes place after flotation
of molybdenum ore by 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 (FeTiO^) and rutile (Ti02).
The United States is a major source of ilmenite but not of
rutile. Since 1972, however, a new operation in Florida has
been producing (5,964 metric tons, or 6,575 short tons, in
1974) 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 operation 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^; however, some sand deposits yield altered
ilmenite (leucoxene) containing 60 percent or more TiO.2, as
well as rutile containing 90 percent or more Ti02-
135
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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
887,506
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
136
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The method of mining and beneficiating 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 Ti02 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.
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, therefore, 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.
137
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Figure 111-31. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS
WET MILL
DRY MILL
r
MAGNETICS
NONMAGNETICS
MONAZITE
TO
SHIPPING
FLOWS (ROUGHERS AND CLEANERS)
NONCONDUCTORS
138
-------
Ilmenite is also currently mined from a rock deposit in New
York by conventional open-pit methods. This ilmenite/
magnetite ore, averaging 18 percent T±02, 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. Beneficiation 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
mining and/or milling of zirconium or for its primary or
byproduct/coproduct recovery.
The principal mineral source of zirconium is zircon
(ZrSiO_4) , 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 monazite) which are
separated from one another by a combination of electrostatic
and magnetic 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 estimated 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) .
139
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Figure 111-32. BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
MINING
ORE
JL
CRUSHING
GRINDING
i
CLASSIFICATION
I
MAGNETIC
SEPARATION
I
MAGNETICS-
i
NONMAGNETICS
MAGNETITE
i
ILMENITE
AND GANGUE
DEWATERER
i
FLOTATION
CIRCUIT
THICKENER
i
TO
WASTE
FILTER
i
DRIER
CONCENTRATE
*
TO SHIPPING
140
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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
subcategorization 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.
141
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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 limitations
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, drainage
tunnels, 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, and loading facilities."
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 represented. 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
142
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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
subcategories which have been made as part of these
recommendations:
(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.
(U) 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 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.
Designation as a Mine or Mill
It is often desirable to consider mine water and mill
process water separately. 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
many instances, it is advantageous to separate mine water
from mill process waste water because of differing water
quality, flow rate or treatability. Levels of pollutants in
mine waters are generally lower or less complex than those
in mill process waste waters. Mine water contact with
finely divided ores, (especially oxidized ores) is minimal
143
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and mine water is not exposed to the suite of process water
reagents often added in milling. Waste water volume
reduction from a mine is seldom a viable option whereas the
technology is available to eliminate all discharge from many
milling operations.
While it is generally more efficient to treat mine waste
water and mill waste water separately, there are some
situations in which combining the mine waste water and mill
process waste water cause a co-precipitation of pollutants
with their resultant discharge being of higher quality than
either of the individual treated discharges. in some
instances, use of the mine waste water as mill process water
will also result in an improved quality of discharge because
of the interactions of the chemicals added to the process
water with the pollutants in the mine water.
Type of 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. 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 treatment
technologies for mine waters from either open pit or
underground mines, designation of the type of mine was not
selected as a suitable basis for general subcategorization
in the industry.
Type of Processing (Beneficiation. 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
pollution 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 generated 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 important parameters being
nearly identical for many subcategories.
144
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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 subcategories. Although usage is usually low,
its presence in effluent streams has potentially harmful
effects. The added reagents may have secondary effects on
the waste water 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.
Ore leaching operations differ substantially from physical
processing and flotation plants in waste water character and
treatment requirements. The use of large quantities (in
relation to ore handled) of reagents, and the deliberate
solubilization of ore components characterizes these opera-
tions, wide diversity of leaching and chemical extraction
processes, therefore, affects the character and quantities
of water quality parameters, 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 subcategorizing each major ore
category.
Mineralogy of the Ore
The mineralogy and host rock present greatly determine the
beneficiation 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 beneficiation 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
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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.
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 unlike many other industry
segments. Therefore, climate and rainfall variations must
be 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. Treatment of large volumes of water by
evaporation in many areas of the United States cannot be
utilized where topographic 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 difficult than in a semi-arid area.
Although climate, rainfall, and location were not used as
primary subcategorization factors, they were given
consideration when determining treatment technology and
effluent limitations (i.e., copper ore industries).
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
guality 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. Mines and mills processing less than 5,000 metric
tons (5,512 short tons) of ore per year in the ferroalloys
industry (most notably, tungsten) are typically intermittent
in operation, have little or no discharge, and are
economically marginal. Pollution potential for such
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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.
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 waste water. 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
subcategorization 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. This characteristic was not found to be a
basis for general subcategorization, however, it was
considered in all subcategories.
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
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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.
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 on 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. While not used for subcategorization,
topography has been considered in the determination of
effluent limits for each subcategory.
Facility Age
Many mines and mills are currently operating which have
operated 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 concentrating 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.
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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
characteristics 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
information 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 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.
Ila. Mine producing effluent - processing plant
producing an effluent - physical processing.
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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 subcategorization 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
Open-pit
Iron formation
Physical processing
Operation B
Medium tonnage
Open-pit
Iron formation
Physical processing
Operation C
Medium tonnage
Open-pit
Natural ore
Physical processing
Operation D
Low tonnage
Open-pit
Natural ore
Older plant (1957)
Mine produces effluent
Processing plant has closed water
system
Medium age plant (1965)
Mine produces effluent
Processing plant has closed water
system
Older plant
No effluent
(1948)
Older plant (1953)
Mine produces effluent
Processing plant produces effluent
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Physical processing
Operation E
High tonnage
Open-pit
Iron formation
Physical processing
Medium age plant (1967)
Mine produces effluent
Processing plant has closed
water system
Medium age plant (1967)
No effluent
Operation F
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
The mines visited and sampled had a 1973 production of
approximately 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
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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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 monitoring 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 presently used concentrating technology.
The milling categorized into three distinct classes based on
the type of ore and the type of processing.
Category Ila. 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 for the
benefication of iron formations.
Final Iron-Ore Subcategorization. Based on the types of
discharges found from all mills, the first two subcategories
can be grouped into a single segment. Mills employing
magnetic separation (No chemical separation) have
demonstrated that a distinct subcategory can be made because
of the type of ore, and the mode of beneficiation.
I. Mines Open-pit or underground, removing natural ores or
iron formations.
II. Iron ore mills employing physical and chemical
separation and iron ore mills employing only physical
separation (not magnetic)
III. Iron ore mills employing magnetic and physical
separation
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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 subcategories based upon
process resulted in grouping operations with similar water
quality characteristics. 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-Ore 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.
II. copper mines employing hydrometallurgical processes
III. Copper mills employing the vat-leaching process
IV. Copper mills employing froth flotation
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. The froth
flotation 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 a few mines send the drainage to the mill tailings
lagoon or use the water in the leach circuits. A decrease
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in excess water problems can be realized in many cases if
mine water is treated separately from mill process water.
Some industry personnel have indicated concern that
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; however,
data has not been provided to support this concern.
molybdenum mills in Canada indicate that the mill tailings
include a built-in blowdown in the form of water trapped in
the interstitial voids of the tailings and the product.
This blowdown removes part of the dissolved salts from a
recyle operation with the result that the circuits can
operate on a zero discharge. 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, land availability, and topography, a preliminary
scheme for subcategorization 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 flotation 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 of 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.
The preliminary subcategorization scheme utilized was
selected to provide subcategorization 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
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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
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
considerably. No basic differences in 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 dewatering 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 sulfide
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.
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The treatability of mine waste water is significantly
affected by the occurrence of local geological conditions
which cause solubilization of undesirable metals or salts.
A common, and well-understood, example is acid mine drainage
caused by the oxidation of pyrite (FeS^) to ferrous sulfate
and sulfuric acid. This oxidation requires both moisture
and air (oxygen source) to occur. The acid generated then
leaches hea\*y metals from the exposed rock on 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 by later contact with limestone or dolomitic
limestone, but the pH level attained may not be high enough
to cause precipitation 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 solubilization conditions do not occur. The treated
effluents from mines in this situation often exhibit higher
levels of heavy metals in solution than untreated mine
waters from mines where solubilization conditions do not
occur.
It has been determined that subcategorization on the basis
of solubilization potential is not justified, however, the
effluent limits recommended have taken into consideration
this factor.
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
subcategorization on this basis.
The final recommended subcategorization for the lead/zinc
mining and milling industry is, therefore, condensed to:
I. Lead and/or zinc mines
II. Lead and/or zinc mills
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
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processing, wastes and waste treatability, water balance,
and final product. Upon intensive background data
compilation (as discussed in Section III) , mill inspections,
and communications with the industry, most of the factors
were found to reduce to mineralogy of the ore (and, thus,
product) and milling 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 characteristic
difference between mine effluents and mill/mine-
mill effluents is their quantitative and
qualitative pollutant loadings. This difference
between mines and mills makes necessary the
application of differing waste-treatment tech-
nologies and/or the segregation of sources for
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 technology.
(2) The specific beneficiation process adapted is based
on the geology and mineralogy of the ore. The
waste characteristics and treatability of the mill
effluent are a function of the particular
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
treatability; 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
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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 Gold-Ore Subcateqorization
On the basis of the rationale developed above and previously
discussed in the introductory portion of this section, six
subcategories were identified for the gold mining and
milling industry:
I. Mine(s) alone.
II. Mill(s) or mine/mill complex(es) 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).
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 SiIver-Ore Subcateqorization
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.
Bauxite Ores
In the bauxite mining industry, most criteria for subcate-
gorization 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 mining
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 emitting no waste water.
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 minedrainage characteristics. In addition, an
evaluation of 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
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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
II. Mines and 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.
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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
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
reasonable. 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 components. Reagents added to effect the flotation may
include major pollutants. Cyanide, for example, is commonly
used and, though usage is low, may necessitate treatment.
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.
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Ore leaching operations differ substantially from physical
processors and flotation plants in effluent character and
treatment requirements. The use of larqe quantities (in
relation to ore handled) of reagents, and the deliberate
solubilization of ore components, characterizes these opera-
tions. The solubilization 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+
SOU., 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 leaching, as in the uranium industry, for example)
but the limited current activity and data available at this
time do not support such a division.
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
discharge 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 the
tails and product. It is anticipated that, under the
impetus of effluent limitations established under PL 92-500
and the resultant pollution control costs, many mills in the
defined subcategories will choose the often less expensive
option of discharge elimination.
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 beneficiating 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
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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 projection).
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 exceptions. 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.
Uranium, Radium, and Vanadium Ores
The factors evaluated in consideration of Subcategorization
of the uranium, radium, and vanadium mining and ore dressing
industry are: end product, type of processing, ore
mineralogy, waste characteristics, treatability of waste
water, 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 and mineralogy (and, thus,
product). 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
hydrometallurgy. Such point discharges as might occur in
milling processes (i.e., the production of concentrate) are
expected to contain a variety of pollutants that need to be
limited. Mining, for the ores, is expected to lead to a
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smaller set of contaminants. 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-radioactive 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. Th4
agreement states, including the uranium producing states of
Colorado, Texas, New Mexico and Washington, have been
delegated all licensing, record keeping, and inspection
responsibilities for radioactive materials regulated by the
NRC upon establishing regulations regarding radioactive
materials that are compatible with those of the NRC(AEC)
The licensing requirements, as set forth in the code of
Federal Regulations, Title 10, Parts 20 and 40 constitute
present restrictions on the discharge of radionuclides
Uranium mines are regulated by some states for discharge of
radioactive materials but this regulation is not based on
"agreement state" authority since the NRC does not regulate
the uranium mines.
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 ferroalloymetal ore mining and dressing, in
another portion of these guidelines. The vanadium
subcategories are summarized there as members of the mining
and hydrometallurgical process subcategories.
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 mills into fewer subcategories. This
simplification is based on the observations discussed below.
Raw waste waters from mills using acid leaching remain acid
at the process discharge (not to be confused with a point
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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 selective 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 rawwaste 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 waste
water 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 Effluent 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 waste
water characteristics but also (a) differences in the volume
of waste water that must be stored and managed in a zero-
effluent condition and (b) differences 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 considered
relative to present and future exploitation of uranium
reserves in the United States. All exploitable uranium
reserves presently economical to develop are found in arid
climates. Therefore, no point discharges are needed to
manage the raw waste water from most current mining and ore
dressing operations in the uranium industry. In addition,
other milling operations that now discharge waste water plan
to terminate their discharges within a year or two.
Ore characteristics were considered and, within a
subcategory, cause short-term effect on waste water
characteristics that does not justify fur^e*
subcategorization. Waste characteristics 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
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treatability of waste water, since all are interrelated.
Vanadium operations which are not extracting radioactive ore
or covered under government licensing regulations (NPC or
agreement states), are subcategorized in the ferroalloys
section.
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 do not concentrate) ores
of uranium, radium, or vanadium.
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 yield concentrates by alkaline
leaching only.
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 of 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
Mining and milling of ore for primary recovery of antimony
is paracticed at one location 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.
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Final Antimony-Ore Subcategorization
The antimony ore mining and dressing industry has been
separated into two 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.
Beryllium Ores
Beryllium mining and milling in the United States are repre-
sented by one operating facility. Therefore,
Subcategorization 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 beneficiation.
Rare-Earth Ores
Rare-earth ores currently are obtained from two types of
mineralogies: bastnaesite and monazite. Monazite is an ore
both of thorium and of rare-earth elements, such as cerium.
The Subcategorization which follows is based primarily upon
division into mines and mills, as well as on the type of
processing employed for extraction of the rare- earth
elements.
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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. Mill(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).
Tin Ores
Some tin concentrate was produced at dredging operations in
Alaska and placer operations in New Mexico. A single
operating facility currently produces tin as a byproduct of
molybdenum mining and beneficiation. Other placer deposits
of tin may be discovered and could be exploited. Therefore,
a single subcategory for mining and one subcategory for
milling are listed:
I. Mine(s) operating for the primary or byproduct
recovery of tin ores.
II. Mill(s) or mine/mill complex(es) using gravity
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, ilmenite, 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
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methods used in conjunction with electrostatic
and/or magnetic methods for the primary or
byproduct recovery of titanium minerals.
Zirconium Ores
Zirconium is obtained from the mineral zircon in conjunction
with dredging operations. No additional subcategorization
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
electrostatic 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
subcategories, a summary of categories and subcategories
recommended for the ore mining and dressing industry is
presented here in Table IV-1. The discussions in the
following sections, including the recommended effluent
limitations in sections IX, X, and XI, will address the
categories and subcategories presented in Table IV 1.
FINAL SUBCATEGORIZATION
After an analysis of available treatment technologies and
the effluent quality that could be achieved by the
application of the available treatment technologies, and the
fact that many metals occur in conjunction with other
metals, it was determined that the final subcategories
previously discussed could be combined into seven
subcategories based on the product or products. The seven
subcategories can then be further divided into 22
subdivisions for which separate limitations will be set,
based on considerations of type of process and waste water
characteristics and treatability. The other factors
recognized as causing differences in the wastes discharged
do not significantly effect the treatability of the wastes
within a subcategory. Table IV-2 shows the final
subcategorization and the components of each subcategory as
they will be presented in the regulations derived from the
development document.
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TABLE IV-1. SUMMARY OF INDUSTRY SUBCATEGORIZATION RECOMMENDED
CATEGORY
SUBCATEGORIES
MINES
IRON ORES
MILLS
Physical and Chemical Separation,
Physical Separation Only
Magnetic and Physical Separation
MINES
Open-Pit, Underground, Stripping
Hydrometallurgical (Leaching)
COPPER ORES
Vat Leaching
MILLS
Flotation Process
LEAD AND ZINC ORES
GOLD ORES
Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation
Byproduct of Base-Metal Operation
MINES
SILVER ORES
BAUXITE ORE
MILLS
MINES
MINES
Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
Byproduct of Base-Metal Operation
FERROALLOY ORES
< 5,000 metric tons (5,512 short tonsl/year
> 5,000 metric tons/year by Physical Processes
> 5,000 metric tons/year by Flotation
Leaching
MERCURY ORES
MILLS
Gravity Separation
Flotation Process
Byproduct of Base/Precious-Metal Operation
URANIUM. RADIUM,
& VANADIUM ORES
ANTIMONY ORES
MINES
MILLS
MINES
MILLS
Acid or Acid/Alkdline Leaching
Alkaline Leaching
Flotation Process
Byproduct of Basc/Precious-Metal Operation
BERYLLIUM ORES
PLATINUM ORES
MINES
MILLS
MINES OR MINE/MILLS
MINES
RARE EARTH ORES
Flotation 01 Leaching
Dredging or Hydraulic Methods
TIN ORES
MILLS
TITANIUM ORES
MINES
MILLS
Electrostatic/Magnetic and Gravity/Flotation Processes
Physical Processes with Dredge Mining
ZIRCONIUM ORES
MILLS OR MINE/MILLS
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TABLE IV-2. FINAL RECOMMENDED INDUSTRY SUBCATEGORIES
SUBCATEGORY SUBDIVISION
1 ron Ores
1
1
^
1
Bauxite Ore
Ferroalloy Ores
Mercury Ores
Uranium, Radium,
& Vanadium Ores
Antimony Ores
Beryllium Ores
Rare-Earth Ores
Titanium Ores
Mines
Mills
Mines
(Open Pit, Under-
ground, Stripping)
Mines
Mills
Mills
Mills
Mills
Mine or
Mine/Mills
Mines
Mines
Mills &
Mines
Mills
Mills
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mills or
Mine/Mills
Physical and Chemical Separation
Physical Separation Only
Magnetic and Physical Separation
Copper
Lead and zinc
Gold
Silver
Hydrometallurgical (Leaching) (Copper)
Vat Leaching (Copper)
Flotation Process (Copper)
Flotation Process (Silver)
Flotation Process (Lead and zinc)
Flotation Process (Gold)
Cyamdation Process (Gold)
Cyamdation Process (Silver)
Amalgamation Process (Gold)
Amalgamation Process (Silver)
Gravity Separation (Gold)
Gravity Separation (Silver)
Gravity Separation (Platinum)
Gravity Separation (Tin)
>5,000 metric tons
<5,000 metric tons (5,512 short
tons/year
>5,000 metric tons/year by
Physical Processes
>5,000 metric tons/year by Flotation
Leaching
Gravity Separation
Flotation Process
Acid or Acid/ Alkaline Leaching
Alkaline Leaching
Flotation Process
Electrostatic/Magnetic and Gravity/
Flotation Processes
Physical Processes with Dredge Mining
Zirconium Ores
Dredging or Hydraulic Methods
(Monazite)
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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 relationship to mine
water effluent due to infiltration or precipitation. 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 categories 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 waste water 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
industry and various regulatory and research bodies, and
from current data for effluent samples collected and
analyzed during this study. The waste water 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 previously undertaken by industry.
Second, the scope of the water-quality analysis was expanded
to include not only previously monitored parameters, but
also waste parameters which could be present in mine
drainage or mill effluents.
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Mine Water
The waste water 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 segment,
process water is not normally utilized in the actual mining
of ores and is present only in placer operations operating
by gravity methods, in hydraulic mining, and in dust
control. Water is a natural feature that interferes with
mining activities. It enters mines by ground-water
infiltration and surface runoff and comes into contact with
materials in the host rock, ore, and overburden. The mine
water then requires treatment depending on its quality
Before it can be safely discharged into the surface drainage
network. Generally, mining operations control surface
runoff through the use of diversion ditching, and grading to
arevent, as much as possible, excess water from entering the
forking area. The quantity of water from an ore mine thus
LS unrelated, or only indirectly related, to production
quantities. Therefore, raw waste loadings are expressed in
zerms of concentration rather than units of production in
-he ore categories discussed in Section IV.
Cn addition to handling and treating often massive volumes
af mine drainage during active mining operations, metal ore
;nine 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 may increase
.as the mine is expanded and developed and may continue after
ill mining operations have ceased. The long-term drainage
:rom tailing disposal also presents long-term potential
>roblems. Surface mines, on the other hand, are somewhat
tore predictable and less permanent in their production of
line drainage period. Water handling within a surface mine
.s fairly uniform throughout the life of the mine. It is
tighly dependent upon precipitation patterns and
»recautionary methods employed, such as the use of diversion
.itches, burial of toxic materials, and concurrent regrading
nd revegetation.
Because mine drainage does not necessarily cease with mine
losure, a decision must be made as to the point at which a
.ine operator has fulfilled his obligations and
-esponsibilities for a particular mine site. This point
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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
water
(3) Miscellaneous water -
dust control
domestic/sanitary uses
washing and cleaning
drilling fluids
Woncontact 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
beneficiation process, comes into direct contact with any
raw material, intermediate product, byproduct, or product
ased in or resulting from the process.
|Moncontact 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.
Tlash 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
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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 transport
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. Water is consumed by being trapped
in the intersitual voids of the product and tailings and by
evaporation.
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 contaminated 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.
PROCESS WASTE CHARACTERISTICS BY ORE CATEGORY
Iron Ore
The quality and quantity of water discharged 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 dependent on the
deposit mined and the substrata through which the water
flows prior to entry into the mine.
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Sources of Waste. The main sources of waste in iron mining
and ore processing are:
(1) Waste water 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 cases, 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 water 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
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.
177
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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. Constituents which may be
present in the 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) Natural 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
waste water. 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
contri bution s.
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-
tuents of waste water. 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
washrooms, lunchrooms, and other areas is usually disposed
of separately from process and transport wastes through
municipal or drainfield systems. Even when not, it would be
expected to constitute a minor part of the load.
178
-------
The principal characteristics of the waste stream from the
mill operations are:
(1) Loadings of 10 to 50 percent solids (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
flotation 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 indication of the presence of organic
reagents, but no definitive information is related by these
parameters.
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 concentrations of particular parameters tend to
associate with a particular mining area or ore body. For
example, the dissolved 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 beneficiation 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
179
-------
TABLE V-1. HISTORICAL CONSTITUENTS OF IRON-MINE DISCHARGES
PARAMETER
TSS
TDS
COD
pH
Oil and Grease
Al
Ca
Cr
Cu
Fe
Pb
MO
Hfl
Ni
Na
Mn
Zn
Chloride
Cyanide
CONCENTRATION
-------
TABLE V-3. CHEMICAL COMPOSITIONS OF SAMPLED MINE WATERS
PARAMETER
pH
Alkalinity
COD
TSS
TDS
Conductivity
Total Fe
Dissolved Fe
Mn
Sulfate
o
£
5
73"
204
274
2
455
440*
0.04
<002
0.21
85
|
UJ
z
5
7.2"
-
482
2
505
400*
<002
<002
<002
175
M
O
UJ
z
5
7.5'
-
9.2
5
609
700*
<0.02
<002
040
215
CONCENTRATION (mj/i ) in WASTEWATER FROM MINE
*~
8,
1
•&
'
7.2"
176
4.5
30
246
310*
<002
<0.02
<0.02
45
£
3
a
&
£
£
'
7.4"
-
1.0
<1
281
320'
<0.02
<0.02
<0.02
28
CM
01
n
|
(5
£
S
7.4'
_
18.3
21
169
21 5t
<0.02
<0.02
0.059
21
CM
&
IT
1
|
7.6"
211
22.8
20
271
340*
0.18
•C0.02
<0.02
26
r*.
o
Ul
z
1
8.4"
218
18
6
1.302
1,960*
4.50
<0.02
3.20
152
1
i
&
E
I
§
71"
37.4
4.5
10
118
110*
2.80
<002
0.025
11.2
£
I
s
o
i
1
7.2"
118
9.0
2
440
550*
1.30
0.04
0.054
33.2
i
UJ
z
i
8.3"
181
27 .5
12
308
342*
0.30
0.02
0.65
36.7
o
UJ
2
I
7.9"
66.0
<10
48
1790
1.125*
1.10
0.08
<0.02
780
1
1
5
§
|
*
7.54"
151
16.7
13.25
499.5
S66.8*
0.86
0.027
0.39
134
V»luem pH units
Value in micromhos/cm
TABLE V-4. CHEMICAL COMPOSITIONS OF SAMPLED MILL WATERS
PARAMETER
PH
Alkalinity
COD
TSS»*
TDS
Conductance
Total Fe
Dissolved Fe
Mn
Sulfate
1102
Tailing-
Pond
Influent
_
-
9.3
30
533
-
210.0
<0.02
330.0
175
CONCENTRATION (mg/t, ) IN WASTEWATER FROM MILL
1103
Tailing-
Pond
Influent
7.5*
-
<1.0
12
198
350f
90.0
0.06
37.50
40
1105
Tailing-
Pond
Influent
8.2*
-
9.2
50
287
3751
1180.0
0.10
320.0
55
1107
Tailing-
Pond
Influent
*
7.6
-
22.5
15
712
-
0.70
< 0.02
67.0
236
1108
Tailing-
Pond
Influent
7.3*
-
13.5
12
230
1301
8.20
0.16
7.50
19.5
1109
Tailing-
Pond
Influent
9.5*
238
13.5
55
360
262 T
0.04
0.04
16.0
20.7
1110
Tailing-
Pond
Influent
and
Minewater
9.00*
13.4
11.9
22
2,360
1 ,900f
< 0.62
< 0.02
0.032
475
Average
(All
Mills)
8.2*
125.7
11.5
28
669
603f
212.8
0.06
111.1
146
Value in pH units
Value in micromhos
* Expressed in %
181
-------
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 dependent 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 multivalent 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 impoundment systems.
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 is
given in Figure V-l.
Process operations generally recycle high percentages of
their water. Water in the plant process is used to wash and
transport 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
182
-------
Figure V-1. FLOW SCHEME FOR TREATMENT OF MINE WATER
MINE WATER
SEDIMENTATION BASIN
SETTLED
SOLIDS
TO
WASTE
CLARIFIED
EFFLUENT
TO RECEIVING
WATERS
t
TO PROCESS
WATER
Figure V-2. WATER FLOW SCHEME IN A TYPICAL MILLING OPERATION
WATER
TO
STOCKPILE'
PROCESS.
PRODUCT
PROCESS PLANT
COAGULANT
f
TO
WASTE
PROCESS
TAILING
i
RECYCLE (80-97%)
SEDIMENTATION
BASIN
P
SETTLED
SOLIDS
CLARIFIED
EFFLUENT
f
TO RECEIVING
WATERS
183
-------
impoundment structures, loss of water to the tailinas anri
evaporation or water loss during processing. taxllngs' and
Process Descriptions
flow within
Mine and Mill 1105. Mine and mill 1105 is a typical
taconite operation. Open-pit mines associated with ?he
Crude magnetic taconxte 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
i^Tar^' iJ?6 ^^ magnetite fro<« the mill is agglomerate!
me^rif tons oT "I v ^^ *W™*^eIy 2?6H million
metric tons (2.6 million long tons) of oxide oellPt«?
annually for blast-furnace feed. pellets
The mine mill and pelletizing plant are located on a large
(20 oS^acr^f £y the operating company, with 8094 hectares
^ an? £ } utxllzed at Present. An initial tailing pond
of 405 hectares (1000 acres) has beeji filled. A second
l,619hectare (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-l Settling
*° C°ntaln ^ "
to wo akes ±S
The mill water system is a closed loop. Plant processes use
oer nn mgd) ' with 189 cubic mers
per minute (72 mgd) returned from the 91.4-meter (300-foot)
diameter tailing thickener overflow and 15.1 cubic me?2rs
per minute (5.7 mgd) returned from the tailing ponTo?
basin. The tailing thickener receives waste or tailings in
a slurry from the concentrate pellet plant. A non?oxic
anionic polyacrylamide flocculant is added to the ?h7cken"er
to assist in settling out solids. Tailing thickener unde?-
flow is pumped to the tailing basin.
Rotary drilling machines are used in the mine to prepare
blast holes for the ammonium nitrate- fuel oil (ANFO) and
used ?o 1oT b^St±n2 agentS- Electric: shoved are
S f i, ? . ? the broken ore into 100-ton-capacitv
diesel/electric trucks for haulage to the primary crusher.
184
-------
Figure V-3. WATER BALANCE FOR MINE/MILL 1105 (SEPTEMBER 1974)
3.4 m3/mm (900 gpm)
(INTERMITTENT)
SETTLING BASINS
CREEK
J)
MINE DEWATERING
17 to 32 m3/min (4,500 to 8,500 gpm)
MAX. 8.23 m3/mm (2,200 gpm)
(INTERMITTENT)
Q LAKE 1 J
2 PUMPS
@ 11.4 m3/mm
(3,000 gpm) EA.
(INTERMITTENT)
PLANT STORAGE TANK
15.1 m3/min (4,000 gpm)
PLANT PROCESSES
204 m3/min (54,000 gpm)
TAILINGS THICKENER
189 m3/min
(50,000 gpm)
15.1 m3/min (4,000 gpm)
s*— "x 15-
( TAILING POND %—
15.1 m3/min (4,000 gpm)
SETTLING BASINS
17to32m3/mm
(4,500 to 8.500 gpm)
C LAKE 2 ^
185
-------
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
m-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
in closed circuit with screens to return undersize material
to the drum and to control the green ball size.
186
-------
Figure V-4. CONCENTRATOR FLOWSHEET FOR MILL 1105
FEED
1
DRY SEMIAUTOGENOUS GRINDING MILLS
I
VERTICAL DRY CLASSIFIER
OVERSIZE
UNDERSIZE
I
CYCLONE CLASSIFIER
OVERSIZE
UNDERSIZE
t
WET MAGNETIC COBBING
OVERSIZE UNDERSIZE
I t
CONCENTRATE
TAIL
DRY MAGNETIC ROUGHER
TAIL
CONCENTRATE
I
DRY MAGNETIC SCREENING
I
MIDDLING
WET SECONDARY
GRINDING MILLS
HYDROCYCLONES
UNDERSIZE OVERSIZE
TAIL
i
HYDROSEPARATION
CONCENTRATE
TAIL
WET FINISHER MAGNETIC SEPARATION
CONCENTRATE
TAIL
TO
PELLET
PLANT
TAILING THICKENER
I
UNDERFLOW
OVERFLOW
TO
WASTE
TO
TAILING POND
TO
REUSE WATER
187
-------
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 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 (one not requiring fine grinding for
concentration) 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
a maximum rate of 7,086 cubic meters per day (1,872,000 gpd)
and at an average 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 U5 cubic meters (12,000 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.
Mining is carried out by conventional open-pit methods.
Ammonium nitrate explosives are used in blasting. 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.
188
-------
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/£ ) 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
189
-------
Figure V-5. FLOWSHEET FOR MILL 1104 (HEAVY-MEDIA PLANT)
CRUDE ORE
(37% IRON)
DOUBLE-DECK
SCREEN
-< 0.64 cm (0.26 in.)-
> 15.2 cm
10.2 to 15.2 cm
(4 to 6 in.)
. I
DOUBLE-DECK
SCREEN
J<24%F«
TO ROCK-REJECTS
STOCKPILE
j3C%Fi
42% p.
CONE
CRUSHER
SCREEN
> 0.64 em
O0.2S in.)
< 0.64 cm
K0.2B in.)
!
HEAVY-MEDIA
SURGE PILE
0.64 to 3.2 cm (0.25,10 1.25 m.l-
< 0.64 cm
K 0.26 in.)
<0.64cm(<0.2Sin.l-
RAKE CLASSIFIER
48m.it. to
(0.64 cm (0.25 in.)
OVERFLOW
I 31% Fe
TO
TAILING POND
HEAVY-
MEDIA
SEPARATOR
(SP.GR. - 3.10)
14% Fi
HEAVY-
MEDIA
SEPARATOR
(SP.GR. - 2.90)
58% Ft
16% F. 60% Ft
i '
1
FLOAT REJECTS FLOAT REJECTS
» i<
TO FLOAT-REJECTS
STOCKPILE
190
-------
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
rem
depending on the quality of the oversize material. P^nt
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
heavymedia surge pile. The fraction which is ^ss 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 suspension treatment are trucked to a stockpile.
Concentrates go directly to a railroad loading pocket. 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 Great Lakes. The 5 8-percent- iron heavy-
media concentrate serves as blast-furnace feed. The 58-
percentiron jig concentrate is later sintered at the steel
plant before entering the blast furnace.
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
constituents 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.
191
-------
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 slime materials removed by wet cycloning. A simplified
flow scheme is included in Figure V-6. Subsequently 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 waste water 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
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.
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
192
-------
TABLE V-6. CHEMICAL CHARACTERISTICS OF DISCHARGE WATER
FROM MINE 1108
PARAMETER
pH
Alkalinity
COD
TSS
TDS
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION
(rna/H )
7.2»
118
9.0
2
440
1.3
0.04
0.054
33.2
•Value in pH units
193
-------
Figure V-6. SIMPLIFIED CONCENTRATION FLOWSHEET FOR MINE/MILL 1108
MINING
I
CRUDE ORE
9.3 m3/min (5.5 cfs
24,500 metric tons
(20,700 long tons)
per day
REGRINDING,
FLOTATION, THICKENING,
AND FILTRATION
I
SECONDARY
CONCENTRATE
i
PELLETIZING
OPERATION
PELLETS
t
TO STOCKPILE
17 m3/min (10 cfs)
41.6 m-Vmin (24.5 cfs)
I ' 14.4 m3/min (8.5 cfs)
17 m3/min (10 cfs)
0.85 m3/min (0.5 cfs)
8.1% SOLIDS
100 m3/min (59 cfs)
I
TO TAILINGS
194
-------
TABLE V-7. CHARACTERISTICS OF MILL 1108 DISCHARGE WATER
PARAMETER
pH
Alkalinity
COD
TSS
TDS
Tout F.
Dtaolv«lF>
Mn
Sulfttl
PROGRAM SAMPLE
CONCENTRATION
W8.HN
WASTE WATER
7.1*
62.0
22.5
10
160
2.06
0.93
0.06
5
WASTE LOAD
in g/nwtric ton
(Ib/fhort ton)
PRODUCT
-
213 10,42)
77.4 (0.1S)
3.4 (0.0071
560 11.08)
7.06 (0.013)
3.2 (0.0061
0.17 (0.0003)
17.2 (0.034)
104MMNTH AVERAGES
AVERAGE
CONCENTRATION
(mg/JU
7.0-
-
8.6
-
0.76
0.66
~
WASTE LOAD
in g/nwtrk: ton
(Ib/lhort ton)
PRODUCT
-
-
20.7 (0.040)
-
1.83 (0.0036)
1.68 10.0031)
~
HIGH
CONCENTRATION
(mg/fc)
7.8*
-
S3
-
3.60
5.80
LOW
CONCENTRATION
(mj/m
6.5'
~
1
~
0.01
0.01
•Vlhn in pH unitl
195
-------
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
mineralization 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 blockcaving or open-pit methods. The choice
of method is determined 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 overburden 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.
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 acid.
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
196
-------
TABLE V-8. PRINCIPAL COPPER MINERALS USED
IN THE UNITED STATES
MINERAL
Chalcocite
Chalcopyrite
Bornite
Covellite
Enargite
COMPOSITION
SULFIDES
Cu2S
CuFeS2
Cu5FeS4
CuS
Cu3AsS4
OCCURRENCE*
SW, NW, NC,**
SW, NW, **
NW, SW
NW, SW
NW
OXIDES
Chrysocolla
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
Cu
NC, SW**
*SW = Southwest U.S.
NW= Northwest U.S.
NC = Northcentral U.S.
**Major minerals
197
-------
or as oxide ore, which does not concentrate economically bv
flotation. J
Water Sources and Usage. in the mining of copper ores,
water collected from the mines may originate from subsurface
drainage or infiltration 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 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 waste water 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.
Air quality control within open-pit mines 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.
Waste Water 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 underground 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.
198
-------
Figure V-7. WASTEWATER FLOWSHEET FOR PLANT 2120-B PIT
NATURAL DRAINAGE,
SEEPAGE, AND
RUNOFF
MINING
(DRILLING,
BLASTING,
AND
LOADING)
ORE-
16,560,000 metric tons/year
(18,250,000 short tons/year)
TO
MILL
EXCESS
MINEWATER
«J
0.06 m /metric ton
(14.4 gal/short ton)
LIME PRECIPITATION
0.06 m3/metric ton
(14.4 gal/short ton)
DISCHARGED
199
-------
TABLE V-S. MINE-WATER PRODUCTION FROM SELECTED MAJOR COPPER-PRODUCING
MINES AND FATE(S) OF EFFLUENT
MINE
2101
2102
2103
2104
2107
2108
2109
2110
2111
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
0.270
0.008
N.E.
0.086
N/A
N.E.
N.E.
N.E.
N.E.
0.015
40.5 (avg)f
1.769
0.030
0.886
0.014
0.654
0.486
0.170
0.034
0.075
N.E.
gal/short ton
ore produced
64.7
1.85
N.E.
20.6
N/A
N.E.
N.E.
N.E.
N.E.
3.5
gjIS.OIavg)*
424.0
7.1
212.3
3.4
156.7
116.4
40.85
8.1
18.0
N.E.
EFFLUENT FATE(S)
Reuse in Dump Leach
Reuse in Mill and Leach
Mine above Water Table
Reuse in Dump Leach
Reuse in Mill
Evaporation and Seepage in Mine
Evaporation and Seepage in Mine
Evaporation and Seepage in Mine
Evaporation and Seepage in Mine
Reuse in Mill
Discharged
Reuse in Mill
Reuse in Leaching
Discharged
Reuse in Dump Leach
Reuse in Mill
Discharged
Discharged
Reuse in Dump Leach
Reuse in Mill
Evaporation and Seepage in Mine
OP = open pit; UG = underground.
0 to 81.1 m3/metr,c ton (0 to 19.432 gal/short ton) ore produced; variable due to seasonal rainfall and
open-pit operations; average calculated assuming six dry (0) and six wet (81.1-m3/19.432-gal) months.
N/A = not available
N.E. = no effluent
200
-------
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
34,765,038*
19,534,193*
51,903,633
20,075,681*
0(UG)
11,400,238*
24,222,246
104,328
8,545,824
45,360 (UG)
17,938,604
10,886,400*
18,144 (UG)
32,257,310*
0(UG)
33,623353*
82,737 (UG)
33,112,800*
O(UG)
88,452,000*
10386,400 *
15,339,844
short tons
38,321,250*
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
12,000,000t
20,000 (UG)
35,557,000*
0(UG)
37,063,000*
91,200(UG)
36,500,000*
0(UG)
97,500,000*
12,000,000*
16,909,000
MILL ORE (1973)
metric tons
7,198,015
7,967,575
13,977,230
7,349,938
N/A
3362^74
1,567,460
3,712,262
1,480^50
635,040
9,383,475
130,386
471,375
11,465,193
1,211,680
16,656,192
19,935,266
23,342,256
8,059,688
34,745,760
1,970,438
7,912398
short tons
7,934,320
8,782,600
15,407,000
8,101,784
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
8,884,136
38,300,000
2,172,000
8,722,000
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
201
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 1 of 4)
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
CONCENTRATION
(mg/K,l
42,01 3.5m3/day
9.64*
544
8
1
5
<10
0.2
0.5
< 0.05
< 0.003
< 0.50
< 0.07
< 0.05
< 0.2
3.80
< 0.05
< 0.05
< 0.10
< 0.2
0.13
0.0008
< 0.05
MINE 2119
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
752 m3/1000 metric tons
9.64*
418.5
6.2
0.77
3.85
<7.69
0.154
0.385
< 0.038
< 0.002
< 0.385
< 0.054
< 0.038
< 0.154
2.923
< 0.0385
< 0.0385
< 0.077
< 0.154
0.10
0.00062
< 0.038
lb/1000 short tons
180,332 gal/1000 short tons
9.64*
837.0
12.4
1.54
7.70
< 15.38
0.308
0.770
< 0.076
< 0.004
< 0.770
< 0.108
< 0.076
< 0.308
5.846
< 0.0770
< 0.0770
< 0.154
< 0.308
0.20
0.00124
< 0.076
MINE 2120-K
CONCENTRATION
(mg/U
27,524.5m3/day
3.49*
4,590
4
< 1.0
31
20
0.10
92.0
0.32
N/A
< 0.02
< 0.07
172.0
< 0.5
2.000.0
100
0.33
0.24
< 0.5
1.35
0.0784
< 0.1
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
15,173 m3/1000 metric tons
3.49*
69,630.3
60.7
O5.17
7.13
303.4
1.52
1 ,395.6
4.85
N/A
< 3.03
< 1.06
2,609.2
< 7.59
30,340
1,517
5.01
3.64
< 7.59
20.48
1.19
< 1.52
lb/1000 short tons
3,635,997 gal/1000 short tons
3.49*
139,260.6
121.4
< 30.34
14.26
606.8
3.04
2,791.2
9.7
N/A
< 6.06
< 2.12
5,218.4
< 15.17
60,680
3,034
10.02
7.28
< 15.17
40.96
2.38
< 3.04
o
to
*Value in pH units
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 2 of 4)
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pta
MINE 2120 B
CONCENTRATION
ImgU)
2,725.2m3/day
6.1*
2,152
40
< 1.0
3.2
<10
0.04
5.30
0 1
0.007
<0.2
< 0.07
31.25
< 0.5
6.00
26.5
1.3
0.13
<0.5
1.55
0.0005
< 0.1
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
60.08 m3/1000 metric tons
6.1*
129.3
2.4
< 0.060
0.192
< 0.601
0.002
0.318
0.006
0.0004
< 0.01 2
< 0.004
1.88
<0.03
0.361
1.592
0.781
0.008
<0.03
0.093
0.00003
< 0.006
lb/1000 short tons
14,400 gal/1000 short tons
6.1*
258.6
4.8
<0 12
0.384
< 1.202
0.004
0.636
0.012
0.0008
< 0.024
< 0.008
3.76
< 0.06
0.722
3.184
1.562
0.016
< 0.06
0.186
0.00006
< 0.01 2
MINE 2120 CE
CONCENTRATION
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 3 of 4)
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
HS
Pb
CONCENTRATION
-------
TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
SELECTED COPPER MINES (Sheet 4 of 4)
O
Ul
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
MINE 2123
CONCENTRATION
(mg/i'l
409m3/day
6.96'
1.350
2
7
10
4
0.07
1.05
< 0.06
0.096
< 0.2
< 0.01
0.1
< 0.5
< 0.1
0.9
< 0.03
< 0.05
< 0.2
0.8
< 0.0001
< 0.5
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
75 m3/10OO metric tons
6.96"
101
0.2
0.5
0.75
0.3
0.005
0.08
< 0.005
0.007
< 0.02
< 0.0008
0.008
< 0.04
< 0.008
0.07
< 0.002
< 0.004
< 0.02
0.06
< O.OOOOO8
< 0.04
lb/1000 short tons
18,000 gal/1000 short tons
6.96"
202
0.4
1.0
1.5
0.6
0.01
0.16
< 0.01
0.014
< 0.04
< O.O016
0.016
< 0.08
< 0.016
0.14
< 0.004
< 0.008
< O.O4
0.12
< O.O00016
< 0.08
*Valu^in pH units
-------
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 (H) 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 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
surfaces 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
206
-------
Figure V-8. FLOWSHEET OF HYDROMETALLURGICAL PROCESSES USED IN
ACID LEACHING AT MINE 2122
TO ATMOSPHERE
TO ATMOSPHERE
43 nr/nxtric ton
(10,282 gri/ihort ton)
241 m3/mttric ton
(57,834 galMiort ton)
EVAPORATION
EVAPORATION
SEWAGE,
SEEPAGE,
RUNOFF,
MINE WATER,
AND WELL WATER
MAI
WA
{
296m3
(70,879 pi
EUP
TER
\
— ^""RESERVOIR
27 m /mttric ton _
(6,426 pl/ihort ton)
2,294 m3/m«tric to
'm.tric ton (549,873 gd/ihort tor
/then ton)
BA
360 m3/RMtric ton SOL
(86,303 j«l/«hort ton) MAKEUP
WATER '
107 m3/nwtric ton r—
U. (25,704 gri/thort ton)
DUMP
LEACH
i ,
PREG
", SOLI
RREN
UTION
L
r
NANT
TION
2,386 m3/m«tric ton
(571.914 ^l/ihort ton)
IK m3/mMric t™
PRECIPITATION ^ (3,727 jrt/ihort ton! poTABLE
PLANT
^ WATER
1
CEMENT
COPPER
03 m3/metric ton
(64 |il/ihort ton)
TO
STOCKPILE
207
-------
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.
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 leaching of surface
ore bodies is similar to a heap or dump leach.
Recovery of Copper From Leach Solutions. Copper dissolved
in leach solutions may be recovered by iron precipitation,
electrowinning, 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.
208
-------
Fiflure V-9. REACTIONS BY WHICH COPPER MINERALS ARE DISSOLVED IN
DUMP, HEAP, OR IN-SITU LEACHING
AZURITE
Cu3(OH)2-(C03)2 + 3H2S04 ^ * 3CuSO4 + 2CO2 + 4H2O
MALACHITE
Cu2(OH)2-C03 + 2H2S04 ~ — ^ 2CuSO4 + CO2 + 3H2O
CHRYSOCOLLA
CuSi03'2H20 + H2S04 ^Z± CuSO4 + SiO2 + 3H2O
CUPRITE
Cu20 + H2S04 ~ ^ CuS04 + Cu + H20
Cu20 + H2S04 + Fe2(S04)3 ^Z± 2CuSO4 + H2O + 2FeSO4
NATIVE COPPER
Cu + Fe2(SO4)3 ^ ^ CuSO4 + 2FeSO4
TENORITE
CuO + H2SO4 v ^ CuSO4 + H2O
3CuO + F«2(S04)3 + 3H20 ^Z± 3 CuSO4 + 2Fe(OH»3
4CuO + 4FeS04 + 6H2O + O2 ^— ^ 4CuSO4 + 4Fe(OH)3
CHALCOCITE
Cu2S + Fe2(SO4)3 ~ ^ CuS + CuSO4 + 2FeSO4
Cu2S + 2Fe2(S04)3 ^ ^ 2CuSO4 + 4FeSO4 + S
COVELLITE
CuS + Fe2(SO4)3 ~ ^ CuSO4 + 2FeSO4 + S
Chalcopyrite will slowly dissolve in acid ferric sulfate solutions and also
oxidize according to:
CuFeS2 + 2O2 7~^ CuS + FeSO4;
CuS + 2O2 ~ V CuS04.
Pyrite oxidizes according to:
2FeS2 + 2H20 + 7O2 ^Z± 2FeSO4 + 2H2SO4.
-------
Figure V-10. TYPICAL DESIGN OF GRAVITY LAUNDER/PRECIPITATION PLANT
DRAINS
SIDE VIEW
CELL
SOLUTION
CELL FLOW *ft
•z-
o
TOP VIEW
CANS
SOLUTION
FLOW
PERFORATED
SCREEN
END VIEW
SOURCE: REFERENCE 23
-------
Figure V-11. CUTAWAY DIAGRAM OF CONE PRECIPITATOR
BARREN
SOLUTION
COPPER
SETTLING AND
COLLECTION
ZONE
COPPER DISCHARGE
SCRAP IRON
DYNAMIC
ACTION
ZONE
COPPER-BEARING
SOLUTION
SOURCE: REFERENCE 23
211
-------
•=*.-*
CuS04 + Fe -r-> cu + FeS04
Scrap iron of other forms and sponge iron may be employed.
: - •—•
cone, into the shredded iron scrap. This
percent The resulting cement copper is 85 to 99 percent
pure and 1S sent to the smelter for further purification?
^lbarre£ ^lution from * Precipitation plant is recycled
from a holding pond to the top of the ore body after
sulfunc acid and makeup water are added, if necessary.
Leach solutions containing greater than 25 to 30 grams per
ties T^e " U8ua11* Sent to electrowinning
Solvent extraction of copper from acid leach solutions by
"6 ±S raPidly Becoming an important methoS of
liquors contain less than 30 grams
1S m°St
.extra?tiQn, a reagent with high affinity for
ainitv fn°\Sn W6ak 3Cid solutionsr and with7 low
?t is DifSd t • ions; is carried in an or^anic medi™-
It is placed in intimate contact with copper leach
solutions, where H+ ions are exchanged for Cu(++) ion?
This regenerates the acid, which is recycled lo the dump"
The organic medium, together with copper, is sent to -
stripping cell, where acidic copper sulfate
212
-------
Figure V-12. DIAGRAM OF SOLVENT EXTRACTION PROCESS FOR RECOVERY OF
COPPER BY LEACHING OF ORE AND WASTE
MINE
DUMP
WEAK CuSO4
LEACH SOLUTION
RAFFINATE_
RECYCLED
SOLVENT
EXTRACTION
PLANT
I
Cu++ ON
ORGANIC CARRIER
RECYCLED
ORGANIC
CARRIER (H+)
i
STRIPPING
AREA
RECYCLED
ACIDIC ELECTROLYTE
(H2S04)
I
CuS04
ELECTROLYTE
i
ELECTROLYTIC
RECOVERY
PLANT
CATHODE
COPPER
I
TO
STOCKPILE
213
-------
exchange H+ ion for Cu(++). This regenerates the organic/m-
media and passes copper to the electrolytic cells, where
impurity-free copper (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
recycling.
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 technigues 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
with sulfuric 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.
214
-------
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
215
-------
TABLE V-13. CHEMICAL CHARACTERISTICS OF BARREN HEAP,
DUMP, OR IN-SITU ACID LEACH SOLUTIONS
(RECYCLED: NO WASTE LOAD)
PARAMETER
pH
TS
TSS
COD
TOC
Oil and Grease
S
As
B
Cd
Cu
Fe
Pb
Mn
Hg
Ni
Tl
Se
Ag
Te
Zn
Sb
Au
Co
Mo
Sn
Cyanide
CONCENTRATION (mg/H) IN LEACH SOLUTION FROM MINE
2120
3.56*
28,148
14
515.8
1.3
<1.0
<0.5
<0.07
0.11
7.74
36.0
2,880.0
0.1
260.0
0.0009
2.40
<1.0
< 0.003
<0.1
1.0
940.0
<0.5
<0.05
3.30
-
-
<0.01
2124
2.82*
47,764
186
1,172
28.0
6.0
<0.5
0.23
0.31
0.092
145.0
6,300.0
<0.1
94.0
0.0012
7.20
<0.1
< 0.040
<0.1
1.0
28.5
<0.5
<0.05
3.80
0.75
-
<0.01
2123
3.56"
44,368
162
80
27.5
2.0
<0.5
0.07
<0.01
5.55
97.0
650.0
0.1
123.5
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
2.49*
83,226
34
385.1
46.0
5.0
<0.5
<0.01
0.08
4.50
72.0
3,500.0
1.14
190.0
0.0003
31.1
<0.1
< 0.003
0.038
2.5
74.5
<2.0
<0.05
72.0
0.35
2.40
•C0.01
2125
4.24*
29,494
218
440.0
11.0
<1.0
<0.5
<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
<0.5
<0.05
13.70
0.5
-
<0.01
2104
3.39*
-
-
-
.
-
-
0.04 to 0.60
-
0.56
52.25
-
0.68
-
0.0003
-
-
0.13
-
-
-
-
-
-
-
-
-
•Value in pH units
216
-------
Figure V-13. VAT LEACH FLOW DIAGRAM (MILL 2124)
TO ATMOSPHERE
ORE WASH WATER
16i m3/nwtric ton
(38,463 (ri/ihort ton)
TO ATMOSPHERE
I m /metric ton
(8,166a«l/thort tonl
EVAPORATION
CRUSHER
H2S04
1.1 rrr/rmtric ton
(2«4 wl/ihon ton
ORE MAKEUP _
280 m /nwtric ton
i
\
LEACH
TANKS
1
l<
2
8.506 gal/short ton)
SLIMfS
TAILS
BARREN SOLUTION-
176 m3/m»tric ton
142.181 oil/ikoit ton)
COPPER-RICH
ELECTROLYTE
287 m3/metric ton
(48,578 gil/ihort ton)
31 m3/mrtric ton
(7,397 ad/short ton)
ELECTROWINNING
W§ m /iMtric ton
(38,463 (ri/ihort ton)
28 m3/mttric ton
(5,620 o>l/thort ton)
10 m3/nwtric ton
(2,411 o>l/ihort ton)
TO MILL
WASH
WATER
TO MILL
TO WASTE
188 m3/rr»tric ton
(45,176 9.1/ihort ton)
TO
STOCKPILE
TO LEACH DUMPS
AND PRECIPITATION
PLANT
217
-------
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
solutions 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.
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 flotation at the concentrator. The
process includes crushing, grinding, classification,
flotation, thickening, and filtration. (See Figure V-11.)
Typically, coarse ore is delivered to the mill for two- or
three-stage reduction by truck, rail or conveyor and is then
218
-------
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
1 2,568 t
49,578
METHOD OF RECOVERY
Solvent Extraction/Iron
Precipitation*
Electrowinning**
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
219
-------
TABLE V-16. CHEMICAL CHARACTERISTICS OF VAT-LEACH BARREN ACID
SOLUTION (RECYCLED: NO WASTE LOAD, MILL 2124)
CONCENTRATION (mg/e )
*Value in pH units
220
-------
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
Hg
Ni
Se
Ag
Ti
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/£)
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
221
-------
Figure V-14. FLOW DIAGRAM FOR FLOTATION OF COPPER (MILL 2120)
MINING I
ORE
TO ATMOSPHERE
195 m3/metric ton
(46.735 9»l/»hort ton)
27 m /metric ton
(6,491 Hl/ihort ton)
15 m3/nwtric ton
(3,709 pl/ihort ton)
EVAPORATION
(3,709 gal/thort ton)
RECYCLES
©*©'
OVERFLOW
(IF ANY)
DISCHARGE
118 m'/mctric ton
(28,189 gil/ihort ton)
222
-------
fed to a vibrating grizzly feeder, which passes its oversize
material to a jaw crusher. The ore then travels by conveyor
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
type
pH control
collector
collector
frother
Example of
Reagent
lime
Xanthate
'Minerec
compounds
MIBC
Ib/short ton
mill feed
10.0
0.01
0.03
0.02
kg/metric ton
mill feed
5.0
0.005
0.015
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.
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
223
-------
Figure V-15. ADDITION OF FLOTATION AGENTS TO MODIFY MINERAL SURFACE
REAGENTS TO
ADJUST pH
COLLECTOR
PULP FROM GRINDING CIRCUIT
(25-45% SOLIDS)
CONDITIONER
i
WETTING AGENT
DISPERSANT
CONDITIONER
•FROTH-
TO
ADDITIONAL
PROCESSING
ACTIVATOR
(OR DEPRESSANT)
CONDITIONER
FLOTATION
CELLS
-TAILINGS-
TO
WASTE
224
-------
TABLE V-17. EXAMPLES OF CHEMICAL AGENTS WHICH MAY BE
EMPLOYED IN COPPER FLOTATION
MINERAL
Bornite
Chalcocite
Ghalaopvrite
Native Copper
Azurite
Cuprite
Malachite
PRECIPITATION AGCNT
—
—
—
—
Sodium monotulfide
Sodium monoiulfide
Bedium menatulfide
PH
REGULATION
Lime
Lime
Lime
Lime
Sodium carbonate
Sodium carbonate
Sodium carbonate
DtSPERCAMT
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
DEPRISSANT
Sodium cyanide
Sodium cyanide
Sodium cyanide
Sodium cyanide
Quebracho
Quebracho
Tannie acid
ACTIVATOR
—
—
—
—
PotytuMide
PotytuMide
Polyiulfio.
COLLECTOR
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats.
Fatty acids
and salts
Fatty acids
and salts,
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
225
-------
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) may separate molybdenum
from copper concentrate, copper sulfide from pyrite, and
copper sulfide from copper/lead/zinc ore. silver may Se
floated from copper flotation feed; gold and silver may be
leached by cyanide from the copper concentrate, with
precipitation by zinc dust.
W^ter Usage in 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 lagoon at this time.
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 aguatic biota, although their
constituents 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.
226
-------
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
227
-------
TA*tE V-U. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAILING PONDS (Sh«t 1 of 4)
II
pH
•M'
TM
QMandGnMt
CQNC€MT RATION
213.080 m3/d.y
0.1J-
1.4N
4
<1.0
Al < 1.0
At
Cd
<0.07
<0.06
0.28
0,06
< 0.6
I
Ag
Sr
Zn
< 0.803
< 0.1
0.83
O»WMinf
< 0.2
< 0.01
MILL 2110 || |ortMM
48,200 »H/*ort ton
0.12*
—
478.883
II CONCENTRATION
JGacMnaiGaB
146,080 m3/d.y
(28,008.000 1.1/d.y)
11.00'
36%
2,082
1 4«K fl
< 336.2
< 336.2
< 23.63
< 16.81
04.13
310.38
V00.1
-------
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAILING PONDS (Sheet 2 of 4)
PARAMETER
Flow
pH
SES*
TDS
TSS
Oil and Grease
SiO2
Al
A<
Cd
Cu
ft
Pta
Mn
Hg
Ni
S>
Ag
Sr
Zn
Sb
Co
Au
Mo
Phosphate
Cyanide
Operating
dayi/yaar
Annual
Production
MILL 2121 (Slime Tails)
CONCENTRATION
(mg/il
64.964 m3/day
17,161 ,000 gal/day)
9.3-
30%
438
202
1.0
4.75
5.9
<0.07
<0.02
3.50
10.05
0.22
0.25
0.0098
<0.05
0.022
<0.1
0.07
0.9
<0.5
<0.04
<0.05
<0.5
0.24
<0.01
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
104.7 m3/metric ton
9.3'
-
45.863
21,151
104.7
49,737
617.8
<7,33
<2.09
366.48
1,052.23
23.04
26.18
1.0261
<5.24
2.304
<10.5
7.33
94.24
<52.4
<4.19
< 524
<52.4
25.13
<1 OS
lb/1000 short tons
25.097 gal/short ton
9.3"
-
91.725
42.302
2O9.4
99.474
1.236
< 14.66
< 4.19
732.96
2.104.65
46.07
52.35
2.0523
< 10.47
4.607
<20.9
14.66
188.48
< 104.7
< 8.38
< 10.47
< 104.7
50.26
< 2.O9
•utn
223,318 metric torn (246,162 short tons)
MILL 2121 (lands)
CONCENTRATION
Img/JU
32,825 m3/day
18.672,400 gal/day)
9.28*
10%
310
6
<1
-
900
<0.07
0.03
46
1.216
0.40
48
O.O001
1.72
< O.O03
<0.1
0.06
8.50
<0.5
1.1
<005
<0.5
-
<0.01
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
52.9 m^/metnc ton
9.28*
—
16,404
317
< 52.9
—
47.624
<3.70
1.59
2,434.1
64.345.3
21.17
2.539.9
0.0053
91.01
< 0.159
< 5.3
317
449.78
<265
58.21
< 2.65
<26.5
-
-------
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAILING PONDS (Sheet 3 of 4)
PARAMETER
Flow
pH
SES*
TDS
TSS
Oil end Grin*
&02
Al
As
Cd
Cu
ft,
Pta
Mn
Hg
Ni
Se
Ag
Sr
Zn
Sb
Co
Au
Mo
Phosphete
Cyanide
Operating
days/year
Annual
Production
of Concentrate
CONCENTRATION
Img/il
278.084 m3/d.v
(73,470,000 gal/day)
8.54'
IS*
4,276
24
3
12.26
<1
<0.07
<0.05
0.08
<0.1
2.79
0.047
0.0002
<0.1
0.022
<0.1
1.81
<0.05
<1.0
0.08
<0.05
<0.2
0.15
<0.01
MILL 2122
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
134 m3/metnc ton
8.54-
-
573,188
3,217
402.1
1,675.60
<134
<9.38
<6.7
10.72
<13.4
373.99
6.3
0.0268
<134
2.949
< 13.4
242.63
< 6.7
<134
10.72
< 67
< 26.8
20.11
< 1.34
lb/1000 ihort tons
32,079 gel/short ton
8.54*
-
1,146,376
6,434
804.3
3,351 19
< 268.1
< 18.77
< 13.4
21.45
< 26.8
747.99
12.6
0.0536
<26.8
5.898
< 26.8
485.25
<13.4
< 268.1
21.45
<13.4
< 53.6
4021
< 268
357
740.602 metric tons (81 7,636 short tonl)
MILL 2123
CONCENTRATION
(mg/£)
11,446 m3/day
(3,024,000 gal/dayl
13.00*
30%
2,494
20
10
27
1
<0.07
<0.03
0.77
0.15
<0.1
< 0.06
0.0019
< 0.05
0.07
<0.1
2.26
-------
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
TAI LING PONDS (Sheet 4 of 4)
PARAMETER
Flow
pH
SES'
TDS
TSS
Oil and Grease
Si02
Al
As
Cd
Cu
ft
Pb
Mn
Hg
Ni
Sa
Ag
Sr
Zn
Sb
Co
Au
Mo
Phosphate
Cyanide
Operating
days/year
Annual
Production
of Concentrat
MILL 21 24
CONCENTRATION
-------
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 discharged.
Handling or Treatment of Decanted Water From Mi11 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 guality of
the discharge.
Three of the copper mills surveyed , all of which discharge
water from the tailing pond, are compared in Table V-20 as
to the quality of, and the amount of loading in, the
discharged 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.
232
-------
TABLE V-20. WASTEWATER CONSTITUENTS AND WASTE LOADS RESULTING
FROM DISCHARGE OF HILL PROCESS WATERS
PARAMETER
TUB
MmriGrMM
M»
CyMM*
CONTROL
TREATMENT
CONCENTRATION
Inn/ « ) IN WASTEWATER
9.6**'
3.3B*
15.0
<0.07
< 0.01
< 0.0*6
0.06
< 0.10
< 0.1
0.03
0.6*11
< 8.06
0.043
2.40
Md tonh soluMn
"•In
ttf
**tt Mn>
\W«. in pH urn*
-------
Figure V-16. FLOWSHEET FOR MISCELLANEOUS HANDLING OF FLOTATION
TAILS (MILL 2124)
SULFIDE
FLOTATION CELLS
TAILINGS
i
HYDROSEPARATOR
SLIMES
i
ACID
LEACH
(THICKENER)
PREGNANT
SOLUTION
BARREN
SOLUTION
PRECIPITATION
PLANT
I
CEMENT
COPPER
CONCENTRATE
•SANDS-
UNDERFLOW
(SLIMES) "
TO
STOCKPILE
TO
TAILING
POND
TO
STOCKPILE
234
-------
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
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.
Leach/Precipitation/Flotation (LPF) Process. Mixed sulfide
and oxide mineralization may also be handled by the
leach/precipitation/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 + FeSO^J
(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 reagents, 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:
235
-------
Figur. V-17. DUAL PROCESSING OF ORE {MILL 2124)
MINING
ORE
CRUSHERS
i
LEACH
RECYCLED
WATER ORE
LEACH
TAILS
_
. ACID
SOLUTION"
RECYCLED AGIO-
DIRECT
SULFIDE
MILL
FEED
i
CONCENTRATOR
RECYCLED
'OVEftFLOW'
.RECYCLED
DECANT
ELECTROWINNING
TAILING
THICKENERS
TO
STOCKPILE
236
-------
Figure V-18. LEACH/PRECIPITATION/FLOTATION PROCESS
COPPER SULFIDE CONCENTRATE
AND
SPONGE COPPER
TO STOCKPILE
237
-------
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
Lead and Zinc Ores
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 million gallons per
day) .
The chemical characteristic of raw waste water from the
milling operation appear to be considerably less variable
from facility to facility than mine waste water. 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 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 infiltration.
(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 in situ before any metals
238
-------
are solubilized. Therefore, the extent of heavy metals in
solution is minimal. The principal contaminants of such
mine waters are:
(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
waste water flow for a mining operation exhibiting no
solubilization 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 three 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 from mines with solubilization
potential are the same as those for mines with no
solubilization potential. The key difference in this
situation 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 waste
water pumped from the mine contains the same waste
parameters as that from the preceding subcategory but also
contains substantial soluble metals. Table V-22 shows the
239
-------
Figure V-19. WATER FLOW DIAGRAM FOR MINE 3105
SEEPAGE
DRILL
COOLING'
WATER 270 m^/day
(72,000 gpd)
MINE
PUMPING
7,600 m3/day
(2,000,000 gpd)
(Ml
MILL FEED-WATER
RESERVOIR
FUEL AND LUBRICANT
SPILLAGE AND
LEAKAGE
EXPLOSIVE
WASTE PRODUCTS
240
-------
TABLE V-21. RANGE OF CHEMICAL CHARACTERISTICS OF SAMPLED
RAW MINE WATER FROM LEAD/ZINC MINES 3102, 3103,
AND 3104 SHOWING LOW SOLUBILIZATION
PARAMETER
PH
Alkalinity
Hardness
TSS
IDS
COD
TOC
Oil and Grease
P
NH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£ )
7. 4 to 8.1*
180 to 196
200 to 330
2 to 138
326 to 510
< 10 to 631
<1 to 4
3 to 29
0.03 to 0.15
< 0.05 to 1.0
< 0.0001 to 0.0001
< 0.2 to 4.9 1
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
Data may reflect influence of acid stabilization on sediment
241
-------
TABLE V-22. RANGE OF CHEMICAL CHARACTERISTICS OF RAW MINE WATERS
FROM FOUR OPERATIONS INDICATING HIGH SOLUBILIZATION
POTENTIAL)
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
IMH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
i
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
1 to 11
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
242
-------
range of chemical constituents from four mines exhibiting
solubilization potential.
The following reactions are the basic chemical reactions
that describe an acid mine-drainage situation:
Reaction 1—Oxidation of Sulfide 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 (watersoluble) ferrous
sulfate:
FeS_2 + 302 > FeS04 + 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:
2F6S2. + 2H20 + 70J2 > 2FeS04 + 2H2SO<4
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:
4FeS04 + 2H2S04 + 02 —> 2Fe2(S04)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
ferrobacillusf Thiobacillus sulfooxidans)^ are involved~IH
the above reaction and, at least, are responsible for
accelerating the oxidation of ferrous iron to the ferric
state.
Reaction 3_—Precipitation of Iron
243
-------
The ferric iron associated with the sulfate ion commonly
combines with the hydroxyl ion of water to form ferric
hydroxide. In an acid environment, ferric hydroxide is
largely insoluble and precipitates:
Fe2_(SOU)J + 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)jl + FeS2 + H2O --- > 3FeSOU + 2S
S + 30 + HO --- > R2SQ
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 O.U5 micrometer (0.000018 inch)
in size and, thus, appear as soluble metals under current
analytical practice.
Conditions compatible with solubilization of certain metal s-
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.
When conditions exist which provide a potential for
solubilization, 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-
244
-------
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 waste water 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.
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.
245
-------
Figure V-20. WATER FLOW DIAGRAM FOR MINE 3104
SEEPAGE-
MINE
(ALL WATER REQUIRED
FOR DRILLING FROM
SEEPAGE)
I
PUMPING
3,460 m3/day
(915,000 gpd)
FUEL AND LUBRICANT
SPILLS AND LEAKAGE
EXPLOSIVE WASTE
PRODUCTS
246
-------
Figure V-21. FLOW DIAGRAM FOR MILL 3103
WATER
FROM MILL
FEED RESERVOIR
9,500 m3/diy
(2,500.000 gpdl
Zn SCAVENGERS Zn ROUGHERS < Zn CONDITIONER
TO TAILING-
POND SYSTEM
TO STOCKPILES
(b) MILL PROCESS
247
-------
TABLE V-23. RANGES OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE
LOADS FOR MILLS 3102, 3103, 3104, 3105, AND 3106
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
MBAS Surfactants
P
Ammonia
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Cyanide
Sulfate
Chloride
Fluoride
RANGE OF
CONCENTRATION
lmg/2 I
IN WASTEWATER
lower limit
7.9-
26
310
<2
670
71.4
11
0
0 18
0042
<005
< 00001
<0 1
0 12
<002
0005
<002
<002
005
< 001
295
21
0 13
upper limit
88-
609
1.760
108
2.834
1.535
35
8
3.7
0 150
14
0 1
1 9
046
036
0.011
067
008
0 53
003
1 825
395
0 26
RANGE OF RAW WASTE LOAD
per unit ore milled
kg/1000 metric tons
lower limit
_
410
460
7
940
6
635
5
0236
0 108
0064
<0 00013
<0127
0089
<0026
0008
<0026
<0026
0064
< 0013
130
20
0370
upper limit
_
1.600
4.700
285
8,500
4.800
130
21
13
0876
26.4
00026
69
17 2
0158
0018
1 77
0290
1 16
0109
4.800
870
0944
lb/1000 short tons
lower limit
_
820
920
14
1.840
12
13
10
047
021
0 125
< 0.00026
<0.25
019
<0052
0016
< 0.052
<0052
0 000129
< 0026
260
40
074
upper limit
_
3.200
9.400
570
17.000
9.600
260
42
26
1 76
528
00052
138
344
0316
0036
344
0580
232
0218
9.600
1,740
1 88
per unit concentrate produced
kg/1000 metric tons
lower limit
_
1.450
2.290
30
4,800
30
30
30
205
054
032
< 000168
<0900
062
< 0 18
<0 18
< 0 18
< 045
0012
0.091
1,260
210
203
upper limit
_
10.200
32,500
2,000
50,900
50,000
580
130
607
254
185
0130
322
860
1 96
885
1 36
100
0198
0509
33,700
4.070
545
lb/1000 short tons
lower limit
_
2,900
4,580
60
9,600
60
60
60
570
1.08
064
< 000336
< 1 8
1 24
< 036
< 036
<036
< 0.90
< 0.024
0 182
2,520
420
4.06
upper limit
_
20,400
65,000
4.000
101,800
100,000
1,160
260
121 4
508
370
0.260
644
172
392
17 7
272
20
0396
1 18
67,400
8,140
10.9
•Value in pH untts
248
-------
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.
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 out
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 and dredgind 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.
249
-------
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 infiltration
into the pit. Ground-water infiltration 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 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.
Waste water 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 a 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 waste water constituents
from placer operations are high suspended solids.
Waste water 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
250
-------
Figure V-22. WATER FLOW IN FOUR SELECTED GOLD
MINING AND MILLING OPERATIONS
(a) MINE/MILL 4101
UNDERGROUND
MINE
3317 m3/d«Y
DISCHARGE TO STREAM
' 2.290 m3/d.y ~"
(600,000 gpd)
AMALGAMATION
MILL*
DISCHARGE TO STREAM
•AMALGAMATION OF GRAVITY-SEPARATED SANDS; FINES AND GOLD-EXTRACTED SANDS
ARC FLOATED FOR RECOVERY OF BASE METALS.
(b) MINE/MILL 4102
4,947 m3/d.¥
11,296.000 gpd)
'GOLD VALUES PRESENT IN BASE-METAL CONCENTRATES. RECOVERED AT SMELTER OR REFINERY.
(c) MINE/MILL 4103
UNDERGROUND
MINE
DISCHARGE
IMPOUNDED
964 m3/div
(250.000 gpd)
'824 m3/d«v
(210,000 gpd)
1,908 m3/d.y
1600.000 tpd)
INTERMITTENT DISCHARGE
DURING APRIL AND MAY
(d) MINE/MILL 4104
251
-------
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, SO4—, Fe, Mn, and Zn).
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.
252
-------
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
Ba
B
Cd
Ca
Cr
Cu
Total Fa
Pb
CONCENTRATION (mg/£)
MINE 4105
275
34f
2.40
1,190
1,176
14
733
35.01
12.0
1
0.095
<0.2
0.03
< 0.002
-------
In general, the cyanidation process involves solubilization
of gold with cyanide solution, followed by precipitation of
gold from solution with zinc dust. (See Figure m-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.
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.
254
-------
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 1104 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 upgraded by filtering and thickening.
Tailings from the flotation process of mill 410U 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 pollutant 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 concentrations 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
255
-------
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
_
_
_
0.002
TABLE V-26. MINERALS COMMONLY ASSOCIATED
WITH GOLD ORE
MINERAL
Arsenopyrite
Pyrite
Chalcopyrite
Galena
Sphalerite
Greenockite
Cinnabar
Pentlandite
Calverite
Sylvanite
Native Gold
Selenium
COMPOSITION
FeAsS
FeS
Cu FeS
PbS
ZnS
CdS
HgS
(Fe, Ni)gS8
Au Te2
(Au, Ag) Te2
Au
Se*
'Accompanies sulfur in sulfide minerals
256
-------
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.
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.
257
-------
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
MILLING OPERATIONS (Sheet 1 of 2)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Leach)
4105
(Vat Leach)
4105 (Carbon
in-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Leach)
(Vat Leach)
4105 (Carbon
m-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laach)
4105
(Vat Laaeh)
tn-Pulp)
MINE/MILL
4102
(Amalgamation)
(Agitation Laach)
(Vat Laach)
in-Pulp)
CONCEN-
TRATION
(mg/i)
495,000
545,000
485,000
CONCEN-
TRATION
(mg/2)
34.3
500
97.0
CONCEN-
TRATION
(ma/£l
0.03
0.17
2.0
CONCEN-
TRATION
(ma/Hi
1.5
<0.5
77.0
TSS
WASTE LOAD
in kg/IOOO metric tons
(lb/10OO short tons)
of concentrate produced
61.695,315,000
(123,390,630)
11,541,465.000
(23.082,930.000)
-
4.7 x ID!]
9.4 x 1011
in kg/1000 metric tons
(to/1000 short tons)
of ore milled
2.871,000
(5,742,000)
436,000
(872,000)
_
4. 17 1.OOO
(8.342,0001
TOC '
WASTE LOAD
in kg/1000 metric tons
lib/1000 abort tonil
of concentrate produced
4,275,000
18,550,000)
1,059,000
(2,1 18.000)
-
94,100,000
1188.200,000)
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
199
1398)
40
(80)
-
830
11,660)
Cu
WASTE LOAD
m kg/1000 metric tons
(lb/1000 abort tonil
of concentrate produced
3,740
(7,480)
3.600
(7,200)
~
1,941,000
(3,882,000)
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
0.2
(0.4)
01
(02)
-
17
134)
Fe
WASTE LOAD
in kg/IOOO metric tons
(lb/1000 short tons)
of concentrate produced
187,000
(374,000)
< 10,600
K2 1.2OOI
-
74,700,000
(149.400.000)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
8.7
(17.4)
•>0.4
K0.8)
-
660
(1,320)
TDS
CONCEN-
TRATION
(mg/£l
462
4.536
-
886
WASTE LOAD
in kg/1000 metric tons
(lb/1000 short tons)
of concentrate produced
19,942.000
(39,884,000)
96,060,000
1192,120,000)
-
859300.0OO
(1,719,800,000)
in kg/1000 metric tons
(lta/1 000 short tons)
of ore milled
930
(1.860)
3,600
(7,200)
7.600
(15.200)
COD
CONCEN-
TRATION
Img/dl
1142
43
-
17894
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
1,423,000
(2,847 000)
911,000
11.822,000)
_
173,700,000
(347,400.000)
in kg/1000 metric tons
(Ib/IOOO short tons)
of ore milled
66
11321
34
1681
-
1,540
1 3.080)
As
CONCEN
TRATION
(mg/)l)
<0.07
005
3.5
-
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
< 8.700
107,400)
106
1212)
1,510.000
(3,020,000)
-
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
<0.4
K0.8I
0.04
10.081
14
(281
-
Zn
CONCEN-
TRATION
(mg/dl
1.3
3.1
-
0.22
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
162,000
1324,000)
65.600
(131,200)
-
213,000
(426,000)
in kg/1000 metric tons
(Ib/IOOO short tons)
of ora milled
7.5
(15.1)
2.5
(51
2
(41
258
-------
• ABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
MILLING OPERATIONS (Sheet 2 of 2)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation LMCh]
4105
IV.t LMCh}
4105 (Carbon
m-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laach)
4105
(Vat Laach)
4105 (Carbon-
in-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laachl
4105
(Vat Laach)
4105 (Carbon
in-Pulp)
Pb
CONCEN-
TRATION
(mgU)
<0.1
<0.1
—
<0.1
WASTE LOAD
in kg/1000 matric torn
lib/1000 Ihort tonil
of concantrata producad
<12,500
K 25.000!
< 2.1 00
K 4,2001
—
< 97.000
«1 94,000)
in kg/1000 matric toni
(t>/1000 short tons)
of ora millad
<0.6
kl.21
<0.08
K 0.1 61
-
<0.9
K1.8I
Hg
CONCEN-
TRATION
Img/i)
0.0011
—
0.004
0.0042
WASTE LOAD
in kg/1000 matrk tons
(Ib/IOOOahorttons)
of concantrata producad
137
(274)
_
1,700
(3.400)
4,070
(8.140)
in kg/IOOO matric toni
lfc/1000 ihort torn)
of on millad
0.0064
(0.0128)
-
0.016
10.032)
0.036
10.072)
SULFIDE
CONCEN-
TRATION
(mg/il
<0.5
<0.5
0.2
1.7
WASTE LOAD
in kg/1000 matric tom
(Ib/IOOOihorttoiw)
of concantrata producad
< 62.000
K 124.000)
< 10.600
K21.200)
86.000
(172,000)
1.650.000
(3.300.000)
in kg/1000 matric ton.
(fc/1 000 short tons)
of ora millad
<2.9
K5.8I
•c-0.4
K0.8)
0.8
(1.6)
15
(30)
CONCEN-
TRATION
<0.02
0.10
<0.01
<0.02
Cd
WASTE LOAD
in kg/1000 matric torn
(lb/1000 ihort torn)
of concantrata producad
< 2,500
K 5.000)
2.100
(4700)
< 4.300
K8.600I
< 19.400
« 38,800)
in kg/1000 matric tons
lib/1000 short toni)
of ora millad
<0.1
K0.2)
0.08
(0.16)
•C0.04
X0.08)
<0.17
K0.34)
CYANIDE
CONCEN-
TRATION
(mg/fcl
<0.01
5.06
-
0.06
WASTE LOAD
in kg/1000 matric tons
I to/1000 ihort torn!
of concantrata producad
< 1.250
K2.500)
107,000
(214.0001
_
58,000
(116,000)
in kg/1000 matric tons
(lb/1000 short tons)
of ora millad
<0.06
K0.12)
4
18)
-
0.52
(1.04)
259
-------
Figure V-23. WATER FLOW IN SILVER MINES AND MILLS
1.132 mj (299.060 pl)/d«y
(a) MINE/MILL 4401
UNDERGROUND
MINE
DISCHARGE
2,933 m3/day
(775,000 gpd)
( r-.acfK \
V ./ 545 m3/
^ (144.00C
^ FinTATinM
day "~ MILL
L_
RAIN
12m3/day
1 ' (3,340 gpd)
<-^ ^--^ _ __.
1.500 m3/d.y *^^ POND ) , " , »• EVAPOflATION
(390.000 ypd) (715 to 914 gpd)
954 n.3/day
(252.000 gpd)
(b) MINE/MILL 4402
260
-------
Sources of Wastes
There are two basic sources of effluents: 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 infiltration into the pit. Ground-
water infiltration 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 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 of silver is obtained from placer
deposits as a byproduct of gold recovery. Waste water from
placer operations is primarily the water which was used in
the gravity separation processing of the ore and/or
hydraulic mining of a deposit. 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 barge pond discharge, and this comprises
the final discharge. The principal waste water constituent
from any placer operations, whether silver, gold, or other
materials, is high loadings of suspended solids.
Waste water 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 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
mineralization of the 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
261
-------
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.
Process Description - Mining
As discussed previously, very little water use is
encountered in 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 and total manganese.
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) ama Igamat ion.
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
262
-------
TABLE V-28. RAW WASTE CHARACTERISTICS OF SILVER
MINING OPERATIONS
PARAMETER
pH
Acidity
Alkalinity
Color
Turbidity (JTUI
TOS
TDS
TSS
Hardness
COO
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
240.8
11.9
17
4
0.085
<0.2
<0.07
< 0.002
<0.6
0.11
<0.02
46.0
<0.1
<0.02
0.33
<0.1
27.5
MINE 4403
.
4.2
76.2
<5t
2.2
622
622
<2
424.8
19.8
16
2
0.030
<0.2
<0.07
< 0.002
<0.5
0.09
<0.02
44.5
<0.1
<0.02
2.05
0.18
32.0
PARAMETER
Mn
Hfl
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Sb
Mo
Chloride
Sulfate
Nitrate
Phosphate
Cyanide
Phenol
Fluoride
Kjeldahl N
Sulfide
SiO2
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
14.5
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
263
-------
bottom of the cells. The collector and activating agents
cause the 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 settled, the very fine
particles of ground ore (slimes) are normally suspended to
264
-------
some extent in 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 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.
Bauxite Ores
Water handling and quantity of waste water flow within
surface bauxite mines are largely dependent upon
precipitation patterns and local topography. Topographic
conditions are often modified by precautionary measures,
such as diversion ditching, disposal of undesirable
materials, regrading, and revegetation. in contrast,
265
-------
TABLE V-29. MAJOR MINERALS FOUND ASSOCIATED
WITH SILVER ORES
MINERAL
Tetrahedrite
Tennantite
Galena
Sphalerite
Chalcopyrite
Pyrite
Naumannite
Greenockite/
Xanthochroite
Garnierite
Pentlandite
Native Bismuth
Argenite
Stephanite
Stibnite
COMPOSITION
(Cu,Fe, Ag)i2Sb4S13
(Cu, Fe,Ag)12As4S13
PbS
ZnS
Cu FeS2
FeS
Ag2S
CdS
(Mg, Ni) O- Si O2 • x H2O
(FefNi)9S8
Bi
Ag2S
Ag5 Sb 84
Sb2S3
266
-------
TABLE V-30. FLOTATION REAGENTS 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)
D-52
2-200 (Isopropl ethylthiocarbamate)
Lime (Calcium oxide)
Sodium cyanide
PURPOSE
CONSUMPTION
g/metric ton
ore milled
MILL 4401
Frother
Frother
Collector
pH Modifier
and Depressant
Depressant
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
Cresylicacid
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
2.50
25
150
200
0.0025
0.0025
0.0075
0.015
0.01
0.005
0.05
0.3
0.4
267
-------
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 1 of 2)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
41 OS
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
TSS
CONCEN-
TRATION
(ing/ HI
550,000
203.000
90.000
WASTE LOAD
in kg/1000 matric tons
(lfa/1000ihorttoiul
of concantrata producad
99.000.000
(198.000.000)
33.901.000
(62.802.000)
9.720.000
(19,440.000)
-
HI kg/1000 matric tona
Id/1000 Ihort torn)
of or. millad
2.475.000
(4.950.OOOI
1.543.000
(3.086,000)
990.000
(1.980.000)
-
TOC
CONCEN-
TRATION
(mo/KI
22.0
24.0
29.0
WASTE LOAD
in kg/1 000 matric toni
lib/1000 short tons)
of concantrata producad
4.000
(8.000)
4.000
(8,000)
3,130
16.260)
-
m kg/1000 matric tona
(*>/1000 short tons)
of ora millad
100
(2001
180
(360)
320
(6401
-
Cu
CONCEN
TRATION
-------
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 2 of 2)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
Hg
CONCEN
TRATION
<2
K4I
-
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
<009
K018I
<015
K030I
<0.2
(CO 4)
-
Mo
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
<36
«72I
<30
K60I
58
11161
< 32,400
K64.800)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
<09
«1 8)
<4
«8I
6
(121
<0.3
(<06)
Cd
CONCEN-
TRATION
lm«/£)
<002
<002
<002
<0.01
WASTE LOAD
in kg/1000 metric tons
(lb/1000 short tons)
of concentrate produced
<3.6
K7.2)
<3.3
(<6.6)
<2.2
K44)
< 6.500
K 13,000)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
<2
K4I
269
-------
underground mine infiltration 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
accumulates 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 of Waste water
Open-Pit Mining. The sequence of operations that occurs in
a typical open-pit mining operation is that the mine site is
cleared of 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
explosive 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.
Regardless of the method of mining, water use at the two
existing operations is generally limited to dust
suppression. Water removal is required because drainage is
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
270
-------
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 segregation 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
wastewater 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 wastewater at times when pumping of
mine water for discharge is required.
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 underground—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
271
-------
Figure V-24. PROCESS AND WASTEWATER FLOW DIAGRAM FOR OPEN-PIT BAUXITE
MINE 5101
EXPLORATION AND ORE-BODY EVALUATION:
GEOLOGICAL SURVEY
TEST DRILLING
SITE PREPARATION:
CLEARING
STRIPPING
RUNOFF
AND
GROUND WATER
MINING:
BLASTING
LOADING
HAULING
RUNOFF
AND
GROUND WATER
MILLING:
CRUSHING AND GRINDING
STORAGE
BLENDING
2.76 m3/metric ton
(664 gal/short ton)
BAUXITE
Q SETTLING PONDS )
REFINING:
COMBINATION PROCESS
2.76 m3/metric ton
i ' (664 gal/short ton) BAUXITE
WATER TREATMENT
PLANT
2.76 m3/metric ton
(664 gal/short ton) BAUXITE
Q SLUDGE POND J
PRODUCTION = 2,594 metric tons (2,860 short tons) per day
WATER TREATED DAILY = 7,165 m3 (1,900.000 gal)
2.76 m3/metric ton
(664 gal/short tort) BAUXITE
DISCHARGE
-------
mined. A room-and-piliar technique is then used to support
the mxne 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 mine
water 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
unconsolldated sands and clays, whereas open-pit drainage is
exposed 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
subcategoriz ation.
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 mg/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.
273
-------
The class of raw mine drainage corresponds closely with
raining 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.
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.
274
-------
TABLE V-32. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID RAW
MINE DRAINAGE FROM OPEN-PIT MINE 5101
PARAMETER
PH
Specific Conductance
Acidity
Alkalinity
IDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION (mg/£)
THIS STUDY
2.8*
1,000*
397
0
560
<2
7.2
3.5
23.8
0.82
0.3
500
0.29
INDUSTRY DATA
* *#
3.4 •
250
0
617
2
15.4**
59.6**
5.9**
25.3**
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
''Averages of five samples
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*
1,580*
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*
2,000t
—
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
275
-------
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 (mg/£)
THIS STUDY
7.2»
1,260f
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 f
150
550
300
5.0
5.0
2.0
1.6
0.01
50
2.5
'Value in pH units
Value in micromhos
276
-------
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-lake 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 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 introduction 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
277
-------
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
(mg/ 2, )
IN WASTEWATER
560 to 1290
< 2 to 42
7.0 to 129.1
2.83 to 9.75
2.76 to 52.3
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 0.1 2
0.02 to 0.36
0.008 to 0.027
0.008 to 0.14
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
< 0.01 2 to 0.24
0.04 to 0.72
0.016 to 0.054
0.016 to 0.28
0.004 to 0.006
0.001 6 to 0.002
2.70 to 3.86
0.0002 to 0.008
Daily flow of waste water = 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 m3 (22,000 gal)
Daily mine production = 390 metric tons (430 short tons)
278
-------
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(1)»
V(2)
V(1)
MILL
Category 1
(< 5,000 metric tons
15,51 2 short tons] per year)
X
Category 2
(Physical
Concentration)
0
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
279
-------
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 waste water streams are
generally characterized by:
(1) High suspended-solid loads
(2) High volume
(3) Low concentrations of most dissolved pollutants.
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 discharge. In addition, the waste
stream is generally contaminated 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++, Na+, K+,
Mg++, C0_3—, 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
intercepted 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
result from direct precipitation on the pond, from runoff
from surrounding areas or even from seepage. and are only
partially amenable to elimination or control.
280
-------
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 of oil and grease from machinery in
the waste water 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
waste water generated 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.
281
-------
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 TOC. 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.
(2) Oxidation of reduced (especially, sulfide) ores,
producing acid and increased soluble material.
(3) Blasting decomposition products, resulting in NH_3,
N0.3, 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
Mri 5.5
282
-------
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 /min (gpml)
2.66 (700)
6.43(1,700)
34.06 (9,000)
11.36 13,000)
pH
4.5.
7.0
6.5
7.3
CONCENTRATION (ms/£)
Oil and
Gr*aw
14
1.0
2.0
Nitrate
0.15
0.12
Fluoride
44.5
4.5
0.52
As
<0.01
<0.01
<0.07
<0.07
Cd
0.07
<0.01
<0.01
<0.005
Cu
3.8
0.06
<0.02
<0.02
Mn
5.3
5.5
021
6.8
Mo
0.5
<0.1
<0.1
<0.1
Pb
O.CX
0.19
0.14
-
V
<0.5
<0.5
-------
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 of Wastes - Mill Effluents
Physical Processing Mill Effluents. In general, mills
practicing purely physical ore beneficiation yield a minimal
set of 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 operations 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.
284
-------
Flotation reagents pose particular problems in effluent
limitation 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 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-
285
-------
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 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 comparison.
Process Description and Raw-Waste Characterization For
Specific Mines and Mills Visited
Mine/Mi11 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.
286
-------
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 MoS2:.
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.
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 MoS^. Tungsten
287
-------
Figure V-25. MILL 6601 FLOWSHEET
22.000 m3/d§y
(6.000.000 gpd)
-UNDERFLOW -1
OVERFLOW
-TAILS
, - ,.
—MIDDLINGS
J
ROUGHER FLOAT
MIDDLINGS
SCAVENGER
FLOAT
-CONCENTRATE-
(4 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
TAILS
-CONCENTRATE—I
MIDDLINGS-
CLEANER
FLOAT
(6 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
-TAILS-
CONCENTRATE
UNDERFLOW
M+UNDERFLOW
•^•UNDERFLOW
22.000 m3/*y
(•,000,000 gpd)
2,200 m3/
-------
TABLE V-39. REAGENT USE IN MOLYBDENUM MILL 6101
REAGENT
Lime
Vapor Oil
Pine Oil
Hypo (Sodium Thiosulfate)
(Na2 $203 • 5H2O)
Phosphorus Pentasulfide (?2 85)
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 Green
COD
At
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Total Cyanide
Fluoride
CONCENTRATION
(mo/Jt )
IN WASTEWATER
600JOOO
2,598
2.0
136
0.01
0.74
51
56.5
5.3
9.8
76.9
1,305
0.02
6.2
TOTAL WASTE
ke/day
14,000,000
42,000
32
2.200
0.16
12
820
900
85
160
1^00
21,000
0.32
99
lb/d.y
32,000.000
92,000
70
4300
0.35
26
1300
2,000
190
350
2,600
46,000
0.70
220
RAW WASTE LOAD
per unit ore milled
kg/metric ton
995
3.0
O.O023
0.16
0.000012
0.00086
0.059
0.064
0.0061
0.011
0.086
1.5
0.000023
0.0071
Ib/short ton
1390
6.0
0.0046
0.32
0.000023
0.0017
0.11
0.13
0.012
0.023
0.17
3.0
0.000046
0.014
per unit concentrate produced
kg/metric ton
610.000
1330
1.4
96
0.0070
0.52
36
39
3.7
7.0
52.4
915
0.014
4.3
Ib/ihort ton
1,200,000
3,670
2.8
190
0.014
1.0
72
79
7.4
14.0
105
1330
0.028
8.7
289
-------
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 crushing 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 MoS^,
is fed to four stages of further flotation. Reagents used
in the primary flotation step are summarized in Table V-m.
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 MoS2
content. Reagent use in the cleaner grinding and flotation
circuit is summarized in Table V-42.
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
temperature raised to 70 degrees Celsius (158 degrees
Fahrenheit), and monazite is removed by flotation. The
tailings from this 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,
290
-------
Figure V-26. SIMPLIFIED MILL FLOW DIAGRAM FOR MILL 6102
CRUSHING
(3 STAGES)
28% ••• 3 MESH
\
GRINDING IN
BALL MILLS
1
36% + 100 MESH
*|
t
FLOTATION
* *
FLOTATION
1
96% OF MILL FEED
TO «
TAILINGS | /~^
V9)
MONA7ITF _--
^ CONCENTRATE
V
RAVITY SEPARATK
lUMPHREY'SSPIRAI
\
PYRITE
FLOTATION
CLEANER Tn
FLOTATION ^- '" ...-.«.
(4 STAGES) TAILINGS
1
DRYING
,
Y
CONCENTRATE
(93% + MO^J
TAILS
TABLES
i^ /O\ INDICATES 1
MONAZITE
FLOTATION
"*~V°)
MAGNETIC
SEPARATION
\< V
NONMAGNETIC MAGNETIC
TIN TUNGSTEN
CONCENTRATE CONCENTRATE
291
-------
TABLE V-41. REAGENT USE FOR ROUGHER AND SCAVENGER
FLOTATION AT IV ILL 8102
REAGENT
Pin* oil
Vapor oH
Syntax
Lime (Calcium oxide)
tedium silicate
Nokes reagent
PURPOSE
Frother
Collector
Surfactant and Frother
Adjustment of pH to 8.0
Slime Dispenent
Lead Depfeasant
CONSUMPTION
kg/metric ton
ore milled
0.18
0.34
0.017
0.15
0.25
0.01S
Ib/ihortton
OM miMed
0.35
0.«7
0.034
0.30
0.50
0.03
TABLE V-42. REAGENT USE FOR CLEANER FLOTATION AT MILL 6102
RfAQfNT
Vapor oil
Sodium cyanide
Nokas reafurt
Dowfroth250
Valco 1861
PURPOSE
Collector
Pyrite and Chatco-
pyrite Depreuawt
Lead Deprewant
Frother
Fteocuiant
CONSUMPTION
kg/metric ton
oremiMed
0.45
0.13
0.45
0.015
0.003
Ib/ihortton
nm ijiiHari
vw IVIICT^V
0.90
OJ5
O.M
0.03
0.006
292
-------
TABLE V-43. REAGENT USE AT BYPRODUCT PLANT OF MILL 6102
(Based on total byproduct plant feed)
REAGENT PURPOSE
CONSUMPTION
kg/metric ton
ore milled
Ib/short ton
ore milled
PYRITE FLOTATION
Sulfuric acid
Z-3 Xanthate
Dowfroth 250
pH Regulation
Collector
Frother
0.018
0.0005
0.0005
0.036
0.001
0.001
MONAZITE FLOTATION
ARMAC C
Starch
Sulfuric acid
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-TAI LINGS SAMPLE)
PARAMETER
TSS
TDS
Oil and Grease
COD
As
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION
(mg/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/.00
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/short ton
1996
9.3
0.016
0.098
<0.0004
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
293
-------
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
significantly 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.
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
294
-------
TABLE V-45. CHEMICAL CHARACTERISTICS OF ACID-FLOTATION STEP
PARAMETER
PH
Cd
Cu
Fe
Mn
Mo
Pb
CONCENTRATION (mg/ A ) AT INDICATED POINTS OF FIGURE V-26
PYRITE FLOAT (19)
4.5«
0.01
02
4.2
4.0
3.0
0.3
MONAZITE FLOAT (20)
1.5
0.042
05
490
53.3
4.0
1.34
•Value in pH units
295
-------
Figure V-27. INTERNAL WATER FLOW FOR MILL 6104 THROUGH
MOLYBDENUM SEPARATION
WET
ORE
WATER
FROM
CREEK
WA
FR
Ml
121 m3/di
132400*
ren
NE
461 m3/*y
(127.000 ffd
i
«" CRIM
)l— *• Al
ORIN
(782400 gpd)
MING (614000 91
EN NO
4H m3/*v (121,000 fpd)
'
16-MtTER
(60-FT)
THICKENER
i i
136 m3/OWN
23 m3Afcy
MoO3
STO
E J2.6 m3/dw
|(700gpd) '
YING SO,
ISTING ^ SCRUBBER
|
) '
i 6,076 m
(1.606,01
T
TO TO
CKPILE TAILING POND
1.336 m3Ahy
(363.000 gpd)
•/diy
**' to SCHEELITE
FLOTATION
UNDE
IFLOW
492m3/.
CONCEI
THICK
PER
WRATE
ENER
4 MO m /ifav
0,162,000 |pd)
toy (130400 gpd)
96 m3/diy
(26400 gpd)
Cu CONCENTRATE PRODUCT
TO STOCKPILE
114 m3/d«y
130400 gpd)
1
FILTI
WAS
299m3/
(79,000
RING
4D
KING
toy
H>d)
1,106 m3/d«y
(292.000 gpd)
2*9m3/dw
(79400 gpd)
MOLYBDENUM
P SEPARATION
3/diy
Mgpd)
409m3/d>y
(106,000 gpd)
TO
» SOLVENT
EXTRACTION
(FIGURE V 28)
296
-------
Figure V-28. INTERNAL WATER FLOW FOR MILL 6104
FOLLOWING MOLYBDENUM SEPARATION
TO ATMOSPHERE
FROM MOLYBDENUM
SEPARATION
(FIGURE V-27)
297,
-------
grinding, sulfides of copper and molybdenum are floated from
the ore, employing xanthate collectors and soda ash for pH
modification. 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
concentration, 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-U6, 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.
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 aguifer producing mine
water, which must be intermittently pumped out (for approxi-
mately hour every 12 hours). Total effluent volume is less
298
-------
TABLE V-46. COMPOSITE WASTE CHARACTERISTICS FOR BENEFICIATiON
AT MILL 6104 (SAMPLES 6, 8,9, AND 11)
PARAMETER
pH
COD
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
Cyanide
CONCENTRATION
-------
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 milled
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/£) 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
300
-------
TABLE V-50. WASTE CHARACTERISTICS OF COMBINED-TAILING
DISCHARGE FOR MILL 6104 (SAMPLES 15,16, AND 17)
PARAMETER
IDS
Oil and Grease
COD
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Total Cyanide
CONCENTRATION
(mg/£) 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
to/short ton
340
2.2
26
-------
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/i) 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
O.OS2
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
<0.0001
<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 8106
PARAMETER
pH
TSS
TDS
Oil and pease
A*
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Ni
CONCENTRATION
Una/ JO
IN WASTEWATER
8.6*
226.9
212
3.4
< 0.07
< 0.005
<0.03
0.53
0.5
<0.1
0.05
24
0.4
TOTAL WASTE
kg/day
-
3,600
3,300
54
<1
<0.08
<0.5
8.3
7.9
<2
0.79
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
par unit or* milled
ko/1000
metric torn
—
790
730
12
<0.2
<0.02
<0.1
1.8
1.7
<0.4
0.17
84
1.4
lb/1000 short toni
—
1,600
1,500
24
< 0.4
<0.04
< 0.2
3.7
3.5
<0.9
0.35
170
2.8
par unit concentrate preduced
kg/1000
metric tons
43,000
39.000
640
<10
< 1
< 6
99
94
<20
9.4
4.500
75
lb/1000 short tons
86.000
79,000
1.300
<20
< 2
<10
200
190
<50
19
9.000
150
•Value in pH units.
302
-------
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 (garnierite) 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. Beneficiation
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.
Mine/Mill 6107
At this operation, vanadium pentoxide, V205, is produced
from an open-pit mine by a complex hydr©metallurgical
process involving roasting, leaching, solvent extraction,
and precipitation. The process is illustrated in Figure
III-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
pelletized. 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 (Na^VKK)^) is concentrated by solvent
extraction, and ammonia is added to precipitate ammonium
vanadate, which is dried and calcined to yield a V205
product.
303
-------
Figure V-29. WATER USE AND WASTE SOURCES FOR VANADIUM MILL 6107
U)
o
NOTE
RUNOFF FROM RAIN
IS NOT CONSIDERED
EXCEPT WHERE IT
ENTERS THE PROCESS
©
-- SAMPLE NUMBER
SAMPLES (7l;AND (72JARE MINE-WATER SAMPLES
-------
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-5U, 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 beneficiation are described in this
section.
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 classification for sizing and feed
into the concentration process.
305
-------
TABLE V-53. WASTE CHARACTERIZATION AND RAW WASTE LOAD FOR MILL 6107
LEACH AND SOLVENT-EXTRACTION EFFLUENT (SAMPLE 80)
PARAMETER
pH
TDS
Otl and grease
COD
NH3
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Ca
Chloride
Fluoride
Sulfate
CONCENTRATION
(mg/SL)
IN WASTEWATER
3.5*
39,350
94
475
0.16
0.35
0.037
1.15
0.15
54
< 0.1
< 0.05
31
0.52
0.26
206
7.900
4.6
26,000
TOTAL WASTE
kg/day
-
83,000
200
1.000
0.34
0.74
0.078
2.4
0.32
110
< 0.2
< 0.1
65
1.1
0.55
430
17,000
9.7
55,000
Ib/day
-
180.000
440
2.200
0.75
1.6
0.17
5.3
0.7
240
< 0.4
< 0.2
140
2.4
1.2
950
37.000
21
120,000
RAW WASTE LOAD
per unit ore milled
kg/metric ton
-
73
0.18
0.88
0.0003
0.00065
0.000068
0.0021
0.00028
0.096
< 0.0002
< 0.0001
0.057
0.00096
0.0005
0.38
15
0.0085
48
Ib/short ton
-
146
0.35
1.76
0.0006
0.0013
0.00014
0.0042
0.00056
0.19
< 0.0004
< 0.0002
0.11
0.0019
0.001
0.75
30
0.017
96
*Value in pH units
306
-------
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/£)
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
<0.1
<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
C0.00004
< 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
3<"»-i
u/
-------
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
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Chloride
Fluoride
Sulfate
CONCENTRATION
(mg/£)
INWASTEWATER
2.3*
2,000
80.768
5
1,844
0.04
0.08
< 0.005
0.9
< 0.03
5.5
-
< 0.05
-
< 0.003
59,500
7.5
780
TOTAL WASTE
kg/day
-
1,900
76,000
4.7
1,700
0.039
0.075
< 0.005
0.86
<0.03
5.2
—
<0.05
—
< 0.003
51,000
7.0
740
Ib/day
—
4,100
160,000
10
3,800
0.086
0.15
<0.01
1.9
<0.07
12
_
<0.1
—
< 0.007
110,000
16
1,600
RAW WASTE LOAD
per unit ore milled
kg/metric ton
—
1.6
67
0.0041
1.5
0.000031
0.000063
< 0.000004
0.00075
< 0.00003
0.0045
—
< 0.00004
—
< 0.000003
45
0.0062
0.64
Ib/short ton
—
3.3
130
0.0085
3.1
0.000063
0.00013
< 0.000008
0.0015
< 0.00006
0.0094
—
< 0.00008
_
< 0.000006
89
0.012
1.3
'Value in pH units
308
-------
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 infiltration 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: 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 infiltration into the pit. Ground-
water infiltration 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.
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
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.
Waste water 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 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. 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
309
-------
Figure V-30. WATER FLOW IN MERCURY MILLS 9101 AND 9102
(NO DISCHARGE)
^S
MILL
RESERVOIR
— -^
' 16.4 m3/day
(4,320 gpd]
GRAVITY-
SEPARATION
MILL
t
fc ( TAILINfi
" \ POND
^T
(NO DISCHARGE)
1,649 m3 day (432,000 gpd)
(a) MINE/MILL 9201
MILL
WELLISI
< -^
/ 1.63 m3/min
(430 gpm)
FLOTATION
MILL
I
k
^f TAILINR A
5.4 m3/min (1,430 gpm) \^^ POND J^
x-^ ^
__to/CLARIFICATION
^\ POND
3.8 m3/min (1,000 gpm)
•DUE TO BEGIN OPERATION IN 1975.
(b) MINE/MILL 9202
(NO DISCHARGE. WATER NOT USED.
BENEFICIATION LIMITED TO
CRUSHING AND/OR GRINDING TO
PROVIDE FURNACE FEED.)
(c) OTHER MERCURY OPERATIONS
310
-------
toxicity, 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.
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
311
-------
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 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 magnesium 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.
312
-------
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
0.20
313
-------
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.
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) subject to NRC 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 wateruse 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-situ
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
economical 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
infiltration from aquifers. All known mills in this
industry segment use a hydrometallurgical process.
The quantity of water used in milling is variable and
depends upon the process used and the degree of recycle.
From these considerations, the effluent quantities are also
314
-------
TABLE V-57. WASTE CHARACTERISTICS AND RAW WASTE LOADINGS
AT MILLS 9201 AND 9202
pH
MILL inpH
unto
•201 6.5
9202
(Pilot Operation! ~
MILL
9201
9202
(Pilot Operation)
MILL
9201
9202
(Pilot Operation)
MINE
9201
9202
(Pilot Operation)
H9
WASTE LOAD
CONCEN-
TBAT,I,PN i" kg/1000 metrie font
In* HI (lb/1000 short tons)
v virisuv ii aae. f»>
-
0.0072 11
122)
in kg/tOOO metric tont
(b/1000 short tons)
of on milled
-
0.094
(0.188)
So
CONCEN-
TRATION
(mg/i)
<0.5
0.03
WASTE LOAD
in kg/1000 nwtrie torn
(lb/1000 short tons)
of concentrate produead
< 6.900
l<1 3,800)
SO
(100)
in kg/1000 nwtrie toni
(to/1000 Ihort tont)
of on milM
<0.06
K0.10)
0.4
10.8)
Tt
CONCEN-
TRATION
(mg/i)
<0.08
"
WASTE LOAD
in kg/1 000 metric tons
(lb/1000 short tons)
of concentrate produoid
< 1.100
K2.200)
-
in kg/IOOO nwtrie tons
(b/1000 ihort tom)
of Of. milwd
< 0.008
(< 0.01 61
-
Zn
CONCEN-
TRATION
(mg/£)
0.14
"
WASTE LOAD
in kg/1000 nwtrie tons
(lb/1000 short tons)
of concentrate produced
1.930
(3.860)
-
m kg/ 1000 nwtrie toni
ll>/1 000 abort tom)
of ora millad
0.014
(0.028)
-
Fa
CONCEN-
TRATION
Imo/d)
-------
Figure V-31. TYPICAL WATER-USE PATTERNS
I
IN-SITU LEACH
MINE
TREATMENT
\
MILL
i
i
^/ IMPOUN
(a) WET MINE/MILL COMPLEX
TREATMENT
MILL
RECYCLE
(b) SEPARATED MILL
I
IN-SITU LEACH
MINE
TREATMENT
DISCHARGE
(c) SEPARATED WET MINE
316
-------
variable. Acid leach mills generally produce between 1.5
and 2.5 tons of liquid per ton of ore; alkaline leach mills
from 0.3 to 0.8 tons of liquid per ton of ore.
Waste Constituents
Radioactive waste Constituents. Radium is one of the most
potentially hazardous radionuelites. The chemistry of
radium is similar to that of calcium, barium and strontium.
The Environmental Protection Agency has proposed interim
drinking water standards for radium -226 and radium -228 at
5pci/l (picocuries per liter) total for both radionuclides.
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 may be
established between the rate of decay of uranium into radium
and the rate of decay of radium into its daughters. Once
this equilibrium is established, the ratio of uranium to
radium equals the ratio of the half-lives—i.e., 2.7
million. An equilibrated ore with a typical grade of 0.22
percent uranium would contain 0.82 microgram of radium per
kilogram. 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 are
approximately 17 to 81,000 picograms per liter. These
concentrations are often quoted in curies (Ci)—i.e., 17 to
81,000 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). In an acid leach ciruit, about 50% of the
thorium and .4 to 6.7% of the radium are dissolved.
Thorium. There are other radioactive species that result
from the decay of uranium. Thorium-230, with a half life of
80,000 years, along with lead 210 and polonium-210r with
half lives of 222 years and 139 days, respectively, are
considered along with radium-226. Thorium is observed in
tailings pond solutions in concentrations from about 10 to
477,000 pCi/liter. A maximum concentration for thorium-230
of 2,000 pCi/liter and for radium-226 of 30 pCi/liter has
been recommended by 10 CFR 20 for release to unrestricted
areas. Generally, it has assumed that methods for control
of radium-226 provide adequate control over thorium and the
other radionuclides of interest.
317
-------
Chemical and Physical Waste Constituents. Chemical
contaminants of milling waste waters 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. One ton of ore containing 4 Ib of U3OJ5 has
about 515 mCi of activity from each member of the decay
chain, with a total combined alpha and beta activity of
about 7,200 mCi. About 85% of the total activity ends up in
the mill waste, and about 1536 is in the uranium product.
With no parent remaining, the thorium-234 and protactium-234
decay out of the mill wastes so that, after a year, the
wastes contain about 70% of the activity originally present
in the ore.
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 quantities of surfactants—
presumably, by saponification (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 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.
318
-------
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)
Alkyl 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
319
-------
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 as Ammonia, Phosphate,
Total Solids, Sulfate, Surfactants
Pb
Acidity, COO, 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 Coliform, Hardness,
Phosphate, Suspended Solids, Total Solids.
Turbidity
At, As, B, Ba, Be, Ca, Cd, Cr, Cu, Fe, Hg, Mg, Mo,
Ni, Pb, Sb. Se, Zn
Ammonia, Chloride, Hardness, Nitrate. 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, Sulfate. Suspended Solids
Al, As, Cr, Fe, Mn, Ni, Pb, Ti, V, Zn
Chloride, COD, Dissolved Solids, Kjeldahl Nitrogen,
Nitrate, Volatile Solids
Co, K. Mn, Na
(None among the analyzed items)
•"Background" is defined in text.
320
-------
Constituents Introduced in Acid Leaching. Acid leaching
(discussed in Section III) dissolves numerous ore
constituents, approximately five 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, molybdic 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 operation 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 condition 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 liguor 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 (NaClO.3) and manganese dioxide (MnO^) 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.
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 sulfides, 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 waste water 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 recarbonation loop. The wastes
discarded to tailings often contain organic compounds
321
-------
Figure V-32. ALKALINE-LEACH WATER FLOW
GROUND ORE
i
TO ATMOSPHERE
T
I
ALKALINE
LEACH
FRESH WATER
OR
TAILING-SOLUTION
RECYCLE
1
FILTERING
I
LEACH
RECYCLE
I
PREGNANT
LEACH
i
PRECIPITATION
CRUDE
PRODUCT
USED
STACK
GAS
I
I
RECARBONATION
t
STACK
GAS
BARREN
LEACH
REPULPED
"TAILINGS"
FRESH
WATER
PURIFICATION
(SODIUM REMOVAL)
WASTE _
WATER ^
k TO TAILING
POND
END
PRODUCT
TO
STOCKPILE
322
-------
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 oxidants 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
sodiumfree 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, 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 contributions 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 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) 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 as a contribution to total organic carbon
(TOC) which is also 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
323
-------
Figure V-33. AMMONIUM CARBONATE LEACHING PROCESS
MINING
ORE
TO ATMOSPHERE
GRINDING
LEACH
SOLUTION"
WASTE GAS
1
AMMONIA AND
CARBON DIOXIDE
ABSORPTION TOWERS
PRESSURE LEACHING
COUNTERCURRENT
DECANTATION
PREGNANT
SOLUTION
SLURRY
•FILTRATE-
TO TAILING
POND
t
STEAM
STEAM STRIPPING
AND URANIUM
PRECIPITATION
SLURRY
FILTRATION
PRODUCT
TO
STOCKPILE
324
-------
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. An analytic procedure for
the entire class of these materials 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 wastewater.
Process Descriptionsr 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 1,000 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 illustrated 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 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
325
-------
TABLE V-60. WATER USE AND FLOWS AT MINE/MILLS 9401, 9402 9403
AND 9404
WATER CATEGORY
Water Supply
Discharge
Supplied to Mill
Recycled to Mill
Loss (Evaporation, etc.)
Makeup Water
Water in Circuit
Discharge
Evaporation and
Seepage
MINE/MILL 9401
m3/day
gpd
WATER USED
MINE/MILL 9402
m^/day
gpd
MINE/MILL 9403
m3/day | gpd
MINE/MILL 9404
m3/day
gpd
MINE PORTION
8,339
3,339
0
5,000
estO
2,700
3,200
0
2,700
2,203,000
882,100
0
1,321,000
est 0
11,552
4,325
5,307
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
MILL PORTION
713,300
845,400
0
713,300
5,307
8,900
0
5,307
1,402,000
2,351,000
0
1,402,000
6,060
6,580
5,400
660
1,601,000
1,738,000
1,427,000
174,400
est 1330
0
0
0
est 1,530
est 404,200
0
0
0
est 404,200
5,300
5,300
0
5,300
1,400,000
1,400,000
0
1,400,000
N/A - Not available
TABLE V-61. WATER TREATMENT INVOLVED IN U/RaA/ OPERATIONS
FEATURE
PARAMETER
MINE/MILL
9401
9402
MINE PORTION
Settling Basin
Evaporating Pond
Ion-Exchange Plant
Area in hectares (acres) II 0.3 (0.74)
Retention Time in hours || est 20
Area in hectares (acres) || N/A
U3 OB Concentration in mg/l II 25
U3 OB Removal in % I 96
0.7(1.7)
est 80
N/A
2 to 12
98
MILL PORTION
Tailing Pond(j)
Ion-Exchange Plant
Rtcarbonizer
Deep Well
Utilization of
Area in hectares (acres)
Number series-connected
Daily Water Use in metric tons (short tons)
Daily Water Use in metric tons (short tons)
Capacity in metric tons (short tons) water per day
Sand/Slime Separators
Decant Facilities
Filters
Coprecipitation
21 (51.8)
1
490 (540)
1,635(1302)
0
Yes
Yes
100(247)
5
N/A
N/A
0
Yes
N/A
N/A
N/A
N/A
N/A
24 (59.3)
3
N/A
520(573)
0
Yes
N/A
N/A
2 (4.9)
N/A
N/A
107 (264)
1
N/A
N/A
1,635(1,802)
Yes
Yes
Yes
TOTAL OPERATION
Ore Handling
Capacity in metric tons (short tons) per day
3,200(3,527)
6.400(7,055)
1,400(1,543)
2,700(2,976)
N/A - Not available
326
-------
Figure V-34. WATER FLOW IN MILLS 9401, 9402, 9403, AND 9404
3,339 m3/dtv
I Ml
i
^*WE
Nc I i |kj ION EXCHANGE L?8?jll!0^K" y» DISCHARGE
1220,200 gpd)
1 IN-SITU LEACH
LUST}— — — - 1».
(713,300 gpd)
1,836 m3/d«y
1431.900 gpd)
6,000 m3/diy
(1.320.000 gpd)
Fflfl miMiy (1*7 0"" grit
MILL |»/
3,200 m3/div
1 W4S.400 gpd)
S TAILING X
>» POND ^^/
2,700 m3/d«y
(713.300 gpdl
I—*- LOSS
(a) MILL 9401
2.136 m3/day
MINE1! I '"^ flp"'
9/»17 m3W
u,..ce | (2.500.000 gpd)
|
^^^ -/ 71V_Ju-, I
(584.000 Jpdi
7^25 m3/diy
(1,900,000 gpdl
1,920 m3/diy
(507^00 gpd)
5,307 m3/d.y
1 11.400.000 gpdl
3,800 m3/d»y
(951.000 gpd)
(578,500 gpdl
ESS USES AND LOSS
MILL
t
(1 ,400.000 gpdl "
QE
4,326 m3/diy
"(1,140.000 gpd)
GE
•S i 5.307 m3/d«y 1
l< 11. 400.000 gpdl
(b) MILL 9402
TO i
ATMOSPHERE 1
EVAPORATION
FAM^.
r—^ 6,080 m3/diy
(1,600,000 gpdl
6,400 m3/d«y
(1, 430.000 gpdl
i
i
860 m3/d« ^^
I (174,400 gpd)
' 520 mj/diy
(137,400 gpd) 1
3 POHO\
••***"^
DISCHARGE f* -- COPRECIPITATION
(c) MILL 9403
t
OPEN-PIT
MINE
_JL_
EVAPORATIOI
POND
HA
1,630 m3/dly I
1404,200 gpd)
1,393 m3/dty "~
(388.000 gpdl
rx
^ 137 m3/diy
(36^00 gpdl
1 TO ATMOSPHERE
UNACCOUNTED LOSS
(d) MILL 9404
327
-------
Figure V-35. FLOWCHART OF MILL 9401
TO ATMOSPHERE
L
TO ATMOSPHERE
TO STOCKPILE
328
-------
Figure V-36. FLOW CHART FOR MILL 9402
MINING
5,300 m3/day
(1,400,000 gpd)
ORE
CRUSHING
AND
GRINDING
H2S04
NadO
NaCI
LEACHING
•SANDS
THICKENERS
SOLVENT
EXTRACTION
PRECIPITATION
YELLOW CAKE
TO STOCKPILE
EVAPORATION
AND SEEPAGE
5,300 m3/day
(1,400,000 gpd)
329
-------
Figure V-37. FLOW CHART OF MILL 9403
WASH EQUIVALENT TO MOISTURE
RETAINED IN CAKE (ELIMINATED
IN DRYER)
TO STOCKPILE
^-"44-HECTARE (108-ACREI"
f TAILING POND WITH 23-HECTARE
N> IBS-ACRE) LIQUID POOL_
330
-------
Figure V-38. FLOW CHART OF MILL 9404
MILL WATER SUPPLY = 5,300 n
RAIN
1,530 m3/day
1 (404,200 gpd)
n3 (1,400,000 gal) per day
TO ATMOSPHERE
EVAPORATION
137 m3/day
(36,200 gpd)
C^^-^^x
320-hectare (790-acre) /^-hectare (4.9-acreWN
OPEN-PIT MINE I EVAPORATION ) )
^^KMV*S/
^^^
TOTAL LOSS OF
5,300 m3/day
It Ann nnn ~**A\ .
U,HUU,UUUgpa/ |\3
LH
[ TAILING A
V POND )
IT
FILTER
CAPACITY OF
1,635 m3/day
(431,900 gpd)
\ •
( DEEP WELL J
MINING
ORE
^^
GRINDING «•—
( WELL 2 J
^~. _-^
1
ACID LEACH
RIM
— ^ * "
EPULP U« SANDS 1 HYDROCYCLONES
T
BARREN
SOLUTION
1 \AfPI 1 d \
1
SLIMES
RESIN-IN-PULP '
ION EXCHANGE -^
1
PREGNANT
SOLUTION
1ELLUE
SE
NT
(HIGH CD
PRECIPITATION
AND
FILTERING
3_j:
WAQHiwr;
-LOW c r-J
331
-------
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).
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
332
-------
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.
Radioactive Nuclides. Decay products of uranium include
isotopes of uranium, thorium, proactinium, radium, radon,
actinium, polonium, bismuth, and lead. 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. Approximately ninety percent or more of the radium
226 remains with solid tailings and sediment in mine-water
settling basins. Concentrations of raw waste of radium and
uranium observed here 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 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 pond— indicate 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,
occasionally, are distinguishable as oils or greases,
surfactants, or phenol. The small amounts of organics that
are observed are reviewed in Table V-63.
Inorganic Anions. These may be distinguished into two
classes: (1) Sulfides, cyanides, and fluorides, for which
technically and economically feasible treatments (e.g.,
oxidation and lime precipitation) are readily available; and
(2) Chlorides, sulfates, nitrates, and phosphates, which are
present in fairly large concentrations in mill wastes and
cannot be removed economically. Distillation and reverse
osmosis, while technically feasible, raise tht cost of
recovered water and requires a large energy expenditure.
Impoundment, in effect, results in distillation in regions
like the southwestern states. Other anions are grouped
together in conjunction with the light-metal cations as
333
-------
TABLE V-62. RADIONUCLIDES IN RAW WASTEWATERS FROM
URANIUM/RADIUM/VANADIUM MINES AND MILLS
RADIONUCLIDE
and units of measurement
RADIUM 226
in picocuries/l
THORIUM
in mg/£
URANIUM
in mg/a
CONCENTRATION
MINES
200 to 3,200
<0.1
4 to 25
ACID MILLS
200 to 700
(4,100)*
(1.D*
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 (TOC)
Oil and Grease
MBAS Surfactants
Phenol
CONCENTRATION (mg/°)
MINES
16 to 45
3to4
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
334
-------
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 waste waters are subject to inclusion in the cate-
gory of total dissolved solids. Calcium, titanium,
magnesium, and aluminum respond to some treatments (e.g.,
lime neutralization) and are shown separately. Table V-65
shows concentrations of aluminum, beryllium, calcium,
magnesium, and titanium found in waste water 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 vanadates. These anionic species are, typically, much
more soluble than cations of these metals that precipitate
as hydroxides or sulfides 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, coprecipitate 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 cationic heavy metals that had been expected to occur
from data on ores and processes include lead, manganese,
a.- jn, 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 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.
335
-------
TABLE V-64. INORGANIC ANIONS IN U/Ra/V RAW WASTEWATER
PARAMETER
Sulfide
Cyanide
Fluoride
Total Dissolved Solids (TOS)
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/RaA/
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 (mg/X )
MINES
0.01 to 0.03
<0.02
0.5 to 1.2
2 to 25
0.5 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
336
-------
TABLE V-67. CONCENTRATIONS OF HEAVY METALS FORMING
CATIONIC SPECIES IN U/Ra/V RAW WASTEWATER
PARAMETER
Silver
Copper
Iron
Manganese
Nickel
Lead
Zinc
CONCENTRATION (mg/£)
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
<0.5to1
0.9 to 1.6
<02to40
0.5
<0.5 to 0.7
0.4
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
337
-------
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-71f 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 TOG 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 contrast,
was concentrating its own, much cleaner, ore. The ratio of
200:1 in IOC is, therefore, expected.
Metal Ores - Not Elsewhere Classified
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
characteristics 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.
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
338
-------
TABLE V-69. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE LOAD
FOR URANIUM MINES 9401 AND 9402
PARAMETER
TSS
COD
TOC
Alkalinity
Ca
Mg
Fe
Mo
V
Ra
Th
U
MINE 9401
CONCENTRATION
(mg/£)
IN WASTEWATER
—
242
15.8
224.4
93
45
0.47
0.5
1.0
3,190*
-
12.1
RAW WASTE LOAD
kg/day
—
2,300
150
2,100
860
420
4
5
9
29,700+
-
113
Ib/day
-
5,200
320
4,600
1,900
920
10
11
20
-
-
248
MINE 9402
CONCENTRATION
IN WASTEWATER
299
600
25
-
117
36
0.23
0.53
<0.5
2,7 10»
<0.1
11.6
RAW WASTE LOAD
kg/day
640
7,000
290
-
1,300
410
3
6
<6
31,100*
<1.2
134
Ib/day
1,400
15,000
640
-
3,000
910
6
13
03
—
<2.5
294
•Value in picocuries/£
Value in picocuries/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
(ma/ )
IN WASTEWATER
294,000
55.6
450
12,200
<0.5
0.92
<0.2
<0.05
0.33
<0.3
17
19,000t
43.9
2.1
TOTAL WASTE
kg/day
3,200,000
150
1,215
32,940
<1.4
2.5
<0.54
<0.14
0.89
<0.81
46
51,300"
118
5.7
Ib/day
7,000,000
331
2.680
72,620
<3
5.5
< 1.2
< 0.3
2
< 1.8
101
261
13
RAW WASTE LOAD
per unit ora milled
kg/metric ton
1,000
0.047
0.38
10
< 0.00042
0.00078
< 0.00017
< 0.000042
0.00028
< 0.00025
0.014
16n
0.041
0.0018
Ib/ihort ton
2,000
0.094
0.76
2)
<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
8,870tf
22
0.98
Ib/snort ton
1,100,000
52
422
11,440
<0.47
0.86
<0.19
< 0.047
0.31
< 0.28
16
20,100"*
45
2.0
tt
•On the basis of 1973 production of 94.5% Uj
^Value in picocur)es/£
Value in microcurim/day
Value in microcuries/metric ton
Value in microcuries/ihort ton
and 5.5% V
339
-------
TABLE V-71. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
LOAD FOR MILL 9402 (ACID- OR COMBINED ACID/ALKALINE-
MILL SUBCATEGORY)
PARAMETER
=
TSS
COD
TOC
Acidity
Al
Cu
Mn
Pb
As
Cr
Mo
V
Ra
U
CONCENTRATION
(mg/£ )
IN WASTEWATER
=====5
525,000
63.5
24.0
35.000
1,594
2.7
105
2.1
2.3
9.0
16.0
125
234t
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
364
RAW WASTE LOAD
per unit ore milled
kg/ metric ton
=====:
1,000
0.082
0.031
45
2.1
0.003
0.14
0.003
0.003
0.012
0.021
0.16
0.30n
0.040
Ib/short ton
=====
2,000
0.16
0.062
91
4.1
0.007
0.27
0.005
0.006
0.023
0.041
0.32
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
0.27*** 136tf
0.080
18
Ib/short ton
900,000
74
28
40,800
1,860
3.1
122
2.4
2.7
10
18.7
146
124***
36
*On the basis of 1973 production of 98.2% U_O0 and 1 8% MO
t o B 3
Value in picocunes/?
"Value in microcuries/day
Value in microcuries/metric ton
***Value in microcuries/short ton
340
-------
TABLE V-72. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
LOAD FOR MILL 9403 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COP
TOC
Alkalinity
C.
Mg
Ti
Al
Cu
F.
Mn
Ni
Pb
Zn
As
Mo
V
Ha
Th
U
Fluoride
CONCENTRATION
(mg/ei
IN WASTEWATER
111,000
27.8
<1
1,150.6
3.200
190
0.395
1 18
1.1
1.6
38
0.52
0.69
< 0.5
1.4
< 0.3
< 0.5
111f
<0.1
3.9
1.4
TOTAL WASTE
kg/day
1,400.000
145
< 5.2
5,980
16.640
990
2.1
94
5.7
8.3
198
2.7
3.6
< 2.6
7.3
< 1.6
< 2.6
580"
<0.5
20
7.3
Ib/day
3,100,000
319
< H
13,190
36.680
2,180
4.5
206
13
18
436
6
7.9
< 5.7
16
< 3.4
< 5.7
-
<1
45
16
RAW WASTE LOAD
per unit on milled
kg/ metric ton
1,000
0.1
< 0.0037
4.3
12
0.71
0.0015
0.07
0.0041
0.0059
0.14
0.0019
0.0026
< 0.0019
0.0052
< 0.0011
< 0.0019
0.411t
< 0.0004
0.032
0.0052
Ib/short ton
2,000
0.2
< 0.0074
8.5
24
1.4
0.0029
0.13
0.0081
0.012
0.28
0.0039
0.0051
< 0.0037
0.01
< 0.0022
< 0.0037
0.37*"
<- 0.0008
0.064
0.01
par unit concentrate produced*
kg/ metric ton
1,050.000
109
3.9
4.500
1.3
743
1.5
70
4.3
6.3
149
2.0
2.7
2
5.5
< 1.2
2
431 "
<0.4
34
5.5
Ib/ihort ton
2,100,000
217
7.8
9,000
2.5
1,486
3.1
141
8.6
13
297
4.1
5.4
4
i -
< 2.3
4
392'"
<0.8
68
11
•On the basil o< 197 3 production of 67% Uj
'value in picocuries/..
"Value in microcunes/day
**Value in microcunes/metnc ton
•"Value in microcuries/short ton
. and 33% CuS.
341
-------
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
Ca
MO
Ti
Al
Cu
Ft
Mn
Ni
Pb
Zn
As
Cr
Mo
V
Ra
U
CONCENTRATION
lmg/2,1
IN WASTEWATER
350,000
629
6.2
4,040
224
550
3
740
0.68
325
210
1.38
0.84
<0.5
0.13
2
< 0.3
120
690*
174.5
TOTAL WASTE
kg/day
2,700,000
3.330
33
21 ,400
1,190
2,920
16
3.920
3.6
1,720
1,110
7 3
4.5
< 27
0.69
11
< 1.6
640
3,660"
925
Ib/day
6,000,000
7,360
72
47.200
2.620
5.430
35
8.650
7.9
3.800
2,450
16
9.8
< 5.8
1.5
23
< 3.5
1,400
-
2,035
RAW WASTE LOAD
per unit ore milled
kg/ metric ton
1,000
1.2
0.012
7.9
0.44
1.1
0.0059
1.5
0.0013
0.64
0.41
0.0027
0.0016
< 0.00098
0.00026
0.0039
< 0.00059
0.24
1.35rt
0.38
Ib/ihort ton
2,000
2.5
0.024
15.8
0.88
2.2
0.012
2.9
0.0026
1.28
0.82
0.0054
0.0033
< 0.002
O.OOO51
0.0079
< 0.0012
0.47
1.23***
0.75
per unit concentrate produced*
kg/ metric ton
530.000
651
6.4
418
232
569
3.1
766
0.7
336
217
1.4
0.9
0.52
0.13
2.1
0.31
124
718"
180
Ib/lhort ton
1,060,000
1,300
12.8
836
464
1.139
6.2
1.532
1.4
673
435
2.9
1.7
1.0
0.27
4.1
0.62
248
652***
361
•On the basis of 1973 production of 100% Ll-j
Value in picocunes/£
"Value in microcuries/day
Value in microcuries/metric ton
•••Value in microcuries/short ton
342
-------
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-UO, and V-41.
Sources of Wastes. There are two basic sources of
effluents: 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 infiltration 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 at this time. Effluents from
beach-sand dredging operations orginate as precipitation,
runoff, and groundwater infiltration. 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.
Waste water 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 discharge, and this comprises the final
discharge. The principal waste water constituents from
these operations are high suspended solid loadings and
343
-------
Figure V-39. WATER FLOWS AND USAGE FOR MINE/MILLS 9901 (ANTIMONY) AND
9902 (BERYLLIUM)
(NO DISCHARGE)
' 306 TO 3(2 m3««y
(80.000 TO 100,000 gpdl
FLOTATION
MILL
286 TO 343 m3/diy
TAILING-
IMPOUNDMENT
EVAPORATION
AND
SEEPAGE
(NO
DISCHARGE)
(a) ANTIMONY MINE/MILL 9901
TO ATMOSPHERE
(b) BERYLLIUM MINE/MILL 9902
344
-------
Figure V-40. WATER FLOWS AND USAGE FOR MINE/MILLS 9903 {RARE EARTHS)
AND 9904 (PLATINUM)
TO ATMOSPHERE
ARGE)
0.08 m'/mmuu
121 gpm)
0.36 m3/minute
(95 gpm)
LEACH CON
'
FLOTAT
i
'
1 .6 m3/mmute
1 0.08 m3/minut« 1
T (20 gpm) 7
EVAPORATION
^1 EVAPO
0.08 m3/minut. ~~ \. K
(20 gpm) ^*— —
, - ' TAI
136m3/minute
(518 gpm)
RECLAIM
TANK
RATION*
)ND )
-ING >
^V POND J
EVAPORATION
0.2 m /minute
(412 gpml
NOTE. FOR BYPRODUCT RECOVERY, SEE PART (b) OF FIGURE V-41 (MINE/MILL 9906)
(a) RARE-EARTH MINE/MILL 9903
RIV
I
'Eft J— |fc{ '
J 24,730 m3/d.v V.
24,730 m3/d«v
(6,480,000 gpdl
>REDGE
POND
./ 49,500 m3/d.y
DREDGE WITH
WET GRAVITY
BENEFICIATION
49300 m3/d«y
(12.960,000 gpd)
(b) PLATINUM MINE/MILL 9904
345
-------
Figure V-41. WATER FLOWS AND USAGE FOR TITANIUM MINE/MILLS 9905 AND 9906
2,668 mj/day
(699,000 gpd)
FLOTATION AND
MAGNETIC-SEPARATION
.MILL
36,069 m3/day
(9,450,000 gpd)
INTERMITTENT
DISCHARGE (SEASONAL)
(9,220,000 gpd)
878 m°/day
(230,000 gpd)
(a) TITANIUM MINE/MILL 9905
TO ATMOSPHERE
BULK
CONCENTRATE"
12,099 ay
(3,170,000 gpd)
12,099 m3/day
(3,170,000 gpd)
TAILING
POND
SYSTEM
12,595 mj/day
(3,300,000 gpd)
DRY
MILL
(ELECTROSTATIC
AND
MAGNETIC METHODS)
7,633 irr/day
(2,000,000 gpd)
EVAPORATION
17,175 m-Vday
(4,500,000 gpd)
DISCHARGE
TO
STREAM
(b) TITANIUM/ZIRCONIUM/MONAZITE MINE/MILL 9906
34C
-------
coloring due to high concentrations of humic acids and
tannic acid from the decay of organic matter incorporated
into former beach sands and gravels being mined.
Waste water 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 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. 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 toxicity, but also because of its effect on
the solubility of the waste constituents.
Waste water emanating from a beach-sand dredging pond
consists of water in excess of that needed to maintain the
pond at the 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 tannic acids which are
indigenous to the ore body and which result in coloring of
the water.
Description of 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
347
-------
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: solubilization 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 commercially valuable ore body of mine
9901 are:
Stibnite (Sb2S_3)
Pyrite (FeS^)
Arsenopyrite (FeAsS)
Sphalerite (ZnS)
Argentite (Ag^S)
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
Northwest Basin. Waste loadings for important constituents
of waste waters 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 (H2BeU^Si209) is recovered by open-pit
methods. A small amount of beryl is also mined in the U.S.
348
-------
TABLE V-74. REAGENT USE AT ANTIMONY-ORE FLOTATION MILL 9901
REAGENT
Dowfroth 250 (Polypropylene glycol
methyl ethers)
Aerofloat 242 (Essentially Aryl
dithiophosphoric acids)
Lead nitrate
PURPOSE
Frother
Collector
Activating
Agent
CONSUMPTION
kg/metric ton
ore milled
0.1
0.5
Ib/short ton
ore milled
0.2
1.0
349
-------
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
Ba
B
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
Total Mn
I
CONCENTRATION (mg/£)
8.3"
8.5
11.0
113*
170
149
68
40
1.5
43
73
6.2
0.23
< 0.002
<0.3
<0.01
0.103
057
0.04
0.12
183
0.13
1.93
0.40
PARAMETER
Hg
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Sb
Mo
Oil and Great*
MBAS Surfactants
Cyanide
Phenol
Fluoride
Total Kjeldahl N
Sulfide
Sulfate
Nitrate
Phosphate
CONCENTRATION (mg/£)
0.0038
0.10
<0.05
<0.2
3.5
0.036
<0.02
2.0
0.11
<0.2
-------
TABLE V-76. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD
AT ANTIMONY MILL 9901
PARAMETER
sass^^sss^^^^ssss^^^ss
PH
TSS
COD
TOC
Fe
Pb
Sb
Zn
Cu
Mn
Mo
CONCENTRATION
(mg/£) IN
WASTEWATER
5==
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/short 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
351
-------
by crude open-cut and hand-picking methods. As indicated in
Figure V-39f 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
approximately 3,061 cubic meters (802,000 gallons) per day
of wastewater are discharged from mill 9902. Waste
constituents 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 SiO,2
Feldspar Al silicates with Ca, K, and Na
Fluorite CaF2
Carbonates
Iron Oxide Minerals
Tourmaline (XY3A16 (B03) 3 (Si6018) (OH) 4)
where X = Na, Ca; 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 relatively high concentration, as are aluminum (552
mg/1) , beryllium (36 mg/1) , and zinc (19 mg/1) .
Rare Earths
Process Description - Rare-Earth Metals Mining. The rare-
earth mineral monazite (Ce, La, Th, Y) PO^t) 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).
352
-------
TABLE V-77. CHEMICAL COMPOSITION OF RAW WASTEWATER FROM
BERYLLIUM MILL 9902 (NO DISCHARGE FROM TREATMENT)
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
Acidity
Alkalinity
Hardness
COO
TOC
Oil and Grease
MBAS Surfactant*
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
CONCENTRATION (mg/£)
17,000"
88*
1.3
18,380
3,035
0
4,000
22
55
<1
0.76
552
0.15
36.0
<5.0
0.65
0.047
43.0
0.20
0.07
-------
As indicated in Figure v-40, no discharge currently exists
at mine 9903.
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. Bastnaesite is
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 concentrate that
averages 57 to 65 percent REOr 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 decantation circuit. The primary thickener
overflow containing the chlorides is fed into the europium
circuit, while the leached solids from the countercurrent
decantation 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
temperature 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 fe "
discussed later.
,_ c—^^-^ wi~J_^./ , J.HV- j.aj-j-j.uauc J..LUIII tue extraction circ
makes up the feed for the lanthanum circuit, which is
After loading the organic with europium, the europium is
stripped in the solvent extraction strip circuit with 4N
hydrochloric acid. The pregnant strip solution contains
354
-------
Figure V-42. BENEFICIATION OF BERTRANDITE, MINED FROM A LODE DEPOSIT.
BY FLOTATION (MILL 9903)
0.36 in3/m*i 1M nT/mimm
(W gpml <61« «pm>
-UNDERFLOW
CYCLONE
FBOTH FIRST AND SECOND ROUGHER
- FROTH _| FLOTAT,ON CELLS
UNDERFLOW
- FROTH
FIRST CLEANER
FLOTATION CELLS
FROTH
SECOND. THIRD, AND FOURTH
CLEANER
FLOTATION CELLS
UNDERFLOW
-UNDERFLOW
SCAVENGER
FLOTATION CELLS
ROTH— I
I
1.96 m'/mmutt
(518 gpm)
LEACHED CONCENTRATE
THICKENER
.TO
WASTE
ALTERNATIVE
SEPARATION OF RARE-
EARTH METALS BY
SOLVENT EXTRACTION
(SEE FIGURE V-43)
TO STOCKPILE
355
-------
Figure V-43. BENEFICIATION OF RARE-EARTH FLOTATION CONCENTRATE BY
SOLVENT EXTRACTION (MILL 9903)
Ul
TO WASTE
1 <
\ t
FILTRATE OVERFLOW
1
FJLTER •*— THICKENER
L— t
DRYER mtCf"J£>0>> -*— AMMONIA
COARSE OR T
FINE
HYDRATE FILTER EUROPIUM
' J PRODUCT
1 "
1 Clllf
THICKENER DRUM _____
FILTER
FILTRATE
1
DRUM
FILTER
FLOTATION
CONCENTRATE
(FROM
FIGURE V-421
Y
DRYER
GADOLIMIUM/SAMARIUM
CARBONATE
SODIUM CARBONATE
1 >
GADOLIMIUIV
PRECIPITA
EURO
I/SAMARIUM
MON TANK
>IUM
PURIFICATION
1
REL
»0
IDE
SODIUM
CARBONATE
' ' ' 1 ' SOLVENT EXTRACTION
i STRIPPING UNIT
3 STOCKPILES , i
SODIUM HYOROSULFIDE
AND AMMONIA
f
PRECIPITATION
TANKS "• RAFFINATE
LANTHANUM CIRCUIT 1
t 1
LOADED BARREN
EUROPIUM ORGANIC
ORGANIC
r
SOLVENT EXTRACTION
UNITS
1ESIOL
PRODUCT
ORGANIC
STORAGE
TANK
£
<•
THICKENER
1
1
OVERFLOW
T
CARBON
FILTERS
1
PREPARATION
TANKS
ORGANIC
HYDROCHLORIC
ACID
EUROPIUM CIRCUIT 1
FEED _.
TANKS
IRON FREE
PREGNANT
EUROPIUM STRIP
SOLUTION
H HEARTH
ROASTER
1 '
| COOLER
uuo m /mm 121 gpml HYDROCHLORIC
' WAI EH ACID SOLUTION
t
LEACH
TANKS
fc THICKENER
"" 2
OVERFLOW
fc CARBON
FILTER
AND — J " " t'~
HEPULPEH Jf
'
CERIUM 1 ' 1 '
CIRCUIT 1
1 ' 1 '
j-T* ,[1
*~ SOLVENT
EXTRACTION
I
LOADED
ORGANIC
SOLVENT EXTRACTION
STRIPPING UNITS
PREGNANT EUROPIUM
STRIP SOLUTION
IRON PRECIPITATION
TANKS
Y
PRESSURE LEAF
FILTER
AGHOUSE 4-, 1 Y '
1 I FILTER ) /~ CERIUM\
1 1 . r .J ( STnoARF 1
CYCLONE CERIUM \-"OIMOS/
CIRCUIT •.
" T i
COARSE OR )
1 ^l FINE CERIUM fc 1
i HYDRATE 1 TO
1 * 1 BLENDED r STOCKPILES
BLENDER | CERIUM fc
SODIUM
"*~ CARBONATE
FERRIC HYnROXinF ^ TO
CAKE WASTE
-------
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 hydroxide 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-UO, 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.
Flotation Circuit
Frother Methylisobutylcarblnol
Collector N-80 Oleic Acid
pH Modifier Sodium Carbonate
Depressants Orzan, Sodium Silicofluroide
Conditioning Agent Molybdenum Compound
Leach Circuit
Leaching Agent Hydrochloric Acid
357
-------
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
SiO2
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/£)
FLOTATION
9.02*
.
;
.
14,476
360,000
-
-
3,100
-
-
-
-
-
-
-
-
-
0.35
-
-
LEACH/
SOLVENT
EXTRACTION
8.23*
345
2,125
80f
5.2
76,162
786
7,220
>1,500
47
<1
21.2
1.25
<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
A8
Na
Sr
Te
Ti
Zn
Mo
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
Cyanide
Phenol
CONCENTRATION (mg/£|
FLOTATION
.
-
0.5
-
-
< 0.3
-
-
-
-
-
-
-
-
-
-
365
-
-
-
-
-
LEACH/
SOLVENT
EXTRACTION
<0.05
6.6
3.0
0.85
<0.1
<0.3
94
0.015
0.09
650
4.5
3.36
7.0
< 0.003
<0.1
54,000
<0.1
2.3
1.50
0.09
<0.01
<0.01
•Value in pH units
Value In cobalt units
358
-------
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/ I )
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
0.27
0.51
0.041
< 0.001
0.006
< 0.001
359
-------
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 (CeFCCH, with La, Nd, Pr, Sm, Gd, and Eu) ;
barite (BaSO^) ; calcite (CaC03J ; and strontianite
The dissolved- solid content of the leach/solvent-extraction
waste stream is extremely high (76,162 mg/1) and is due
largely to chlorides (54,000 mg/i) . 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.
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 separated from the platinum-group metal
alloys and minerals by magnetic separation. The final
platinum- group metal concentrate 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
360
-------
TABLE V-80. CHEMICAL COMPOSITION AND RAW WASTE LOAD FROM RARE-EARTH
MILL 9903
PARAMETER
PH
TSS
TOC
Cr
Mn
V
Fluoride
PH
TSS
TOC
SiO2
Cr
Mn
V
Te
Ni
CONCENTRATION
(mg/£) IN
WASTEWATER
RAW WASTE LOAD f
per unit of concentrate
kg/metric ton ] Ib/short ton
per unit ore milled
kg/metric ton | Ib/short ton
(a) Flotation Mill
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
<0.0016
1.89
(b) Leach/Solvent-Exchange Mill
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
<0.0003
0.003
0.001
_
1.67
0.094
0.00250
0.00008
0.006
< 0.0006
0.006
0.002
_
_
—
_
_
_
-
__
— _
_ ,
—
_
^
t
,__„.
-
Value in pH units
Based upon maximum production achievable (part a) or estimated amount of flotation concentrate
produced (part b)
361
-------
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 waste water 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 (FeCr204)
Ferroplatinum (Fe, Pt, Ir, Os, Ru, Rhr Pd, Cu, Ni) alloy
Iridium/ruthenium/osmium alloy
Taurite (Ru, Ir, Os)S2
Unnamed mineral (Ir, Rh, Pd)S
Mertieite (Pt5(Sb, As) 2)
Sperrylite (PtAsj?)
Gold (Au)
Tin
Tin is recovered in the U.S. as a byproduct of a molybdenum
operation. At this mine (6102), the ore is mined by glory-
hole methods, in which the sides of an open hole are 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 for this mining
milling operation.
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 operations 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
362
-------
TABLE V-81. CHEMICAL COMPOSITION AND LOADING FOR PRINCIPAL WASTE
CONSTITUENTS RESULTING FROM PLATINUM MINE/MILL 9904
{INDUSTRY DATA)
PARAMETER
Alkalinity
Conductivity
Hardness
COD
BOO
TS
TDS
TSS
(N) NH3
Kjeldahl Nitrogen
Al
Cd
Cr
Cu
Total Fe
Pb
Zn
Chloride
Fluoride
Nitrate
Sulfate
Sulfide
CONCENTRATION
(mg/ £ )
WASTEWATER
83
109*
35.6
7.6
3.5
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
363
-------
TABLE V-82. CHEMICAL COMPOSITION OF RAW WASTEWATER
FROM TITANIUM MINE 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Total Kjeldahl N
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/£)
1,000'
11.3*
0.37
1,240
! 14
6.4
138.2
546.4
6.4
10.3
3.0
0.32
2.24
0.1
0.1
0.003
<1
0.01
< 0.002
94.5
<0.01
<0.03
0.33
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Sr
Ag
Na
Sa
Ta
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
1 — 1
CONCENTRATION (mg/K)
<0.05
26.0
<0.01
<0.01
<0.1
<05
13.0
0.129
<0.01
140.0
0.75
< 0.06
< 0.2
0.007
<0.1
< 0.1
< 0.01
183.5
3.20
270
15.52
< 0.05
•Value in micromhos/cm
Value in cobalt units
364
-------
Figure V-44. BENEFICIATION AND WASTE WATER FLOW OF ILMENITE
MINE/MILL 9905 (ROCK DEPOSIT)
MINING
I
ORE
CRUSHING
878 mj/day
(232,000 gpd)
36,069 nT/day
(9,450,000 gpd)
GRINDING
CLASSIFICATION
I
MAGNETIC
SEPARATION
MAGNETICS «-NONMAGNETICS-i
ILMENITE
AND GANGUE
FILTRATE
WATER
"RETURN TO MILL
35,191 m3/day
(9,220,000 gpd)
FLOTATION
CIRCUIT
THICKENER
I
FILTER
1
DRIER
I
CONCENTRATE
f
TO SHIPPING
365
-------
TABLE V-83. CHEMICAL COMPOSITION OF RAW WASTEWATER
FROM TITANIUM MILL 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hardness
COO
TOC
Oil and Grease
MBAS Surfactants
Total Kjeldahl N
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/H)
650*
18.0*
2.2
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
350
0.58
0.43
500
PARAMETER
Pb
Mg
Total Mn
Ni
T(
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
CONCENTRATION (mg/£)
< 0.06
187.5
5.9
1.19
<0.1
2.0
23.7
0.132
0.015
41
0.29
< 0.06
2.08
7.6
< 0.1
< 0.1
< 0.01
19.1
32.5
213
0.68
< 0.05
•Value in micromhos/cm
tValue 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
SSSSSSSSSSSSSS^S^SSS^S^Sl
2.66
1.80
0.016
1.52
3.55
366
-------
TABLE V-85. PRINCIPAL MINERALS ASSOCIATED WITH
ORE OF MINE 9905
MINERAL
llmenite
Magnetite
Pyroxene
Feldspar
COMPOSITION
Fe Ti 03
Fes 04
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
(mg/2,) IN
WASTE WATER
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
0.006
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
367
-------
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
indigenous to the ore body. However, 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-UU) .
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
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 ilmenite 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 waste water 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
principal 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
compilation of the concentrations of the principal
constituents of raw waste water from mill 9905.
368
-------
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 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 operations 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 concentrations of principal waste water
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.
369
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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
I
20,100 m3/day
(5,310,000 gpd)
ORE FED
FROM DREDGE
i
EVAPORATION
POND
WATER
RECYCLE
H2O
J_
VIBRATING
SCREENS
WASH
WATER*
11,000 m3/day
(3,170,000 gpd)
I
7,570 m3/day
(2,000,000 gpd)
—OVERSIZE-
SPIRALS OR LAMINAR
FLOWS (ROUGHERS AND CLEANERS)
I
TO DRY MILL
(FIGURE III-30)
WELL
t—TAILINGS-i
12,000 m3/day
(3,170,000 gpd)
WASTE (DREDGE)
POND
15,615 m3/day
(4,500,000 gpd) ,
DISCHARGE
370
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TABLE V-87. CHEMICAL COMPOSITION OF RAW WASTEWATER
AT MILLS 9906 AND 9907
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
COO
TOC
Total Kieldahl N
Oil and Grease
MBAS Surfactants
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
CONCENTRATION (mg/£)
MILL 9906
200"
51,400*
<0.1
1,644
11,000
47.2
47.6
1,338
972
0.6S
400
<0.01
69.0
0.05
< 0.002
<0.5
0.10
< 0.002
0.10
0.03
<0.03
MILL 9907
40»
16,240*
0.54
370
209
31.4
3.4
362
321
0.65
40.0
<0.01
15.0
0.03
< 0.002
<0.5
0.04
< 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 (mg/£)
MILL 9906
4.9
<0.05
1.63
0.036
<0.01
<0.1
<0.5
3.5
<0.05
<0.01
27.0
<0.05
<0.06
<0.2
0.014
<0.1
<0.1
30.0
0.03
0.35
<0.01
MILL ((07
033
<0.05
0.66
0.01
<0.01
<0.1
<05
1.3
<0.05
' < 0.01
5.0
<0.05
0.15
0.40
0.002
<0.1
<0.1
15.0
<0.01
040
<0.01
•Value in micromhos/cm
Value in cobalt units
371
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TABLE V-88. RAW WASTE LOADS FOR PRINCIPAL WASTEWATER CONSTITUENTS
FROM SAND PLACER MILLS 9906 AND 9907
PARAMETER
TSS
TOC
COO
Oil and Grease
Ti
Fe
Mn
Cr
Phosphate
MILL 9906
CONCENTRATION
IN WASTE WATER
11.000
972
1,337
400
<0.2
4.9
0.36
0.03
0.35
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
330
29.2
40.13
12
< 0.006
0.15
0.0011
0.0009
0.011
Ib/short ton
660
58.4
80.26
24
< 0.012
0.30
0.0022
0.0018
0.022
CONCENTRATION
(mfl/2)
IN WASTEWATER
209
321
361.6
40
0.4
0.93
<0.01
<0.01
0.4
MILL 9907
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
5.01
7.71
8.68
0.96
0.01
0.022
< 0.0024
< 0.0024
0.01
Ib/short ton
10.02
15.42
17.36
1.92
0.02
0.044
< 0.0048
< 0.0048
0.02
372
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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 waste water
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 constituents 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 waste water from each subcategory of this study.
The waste water constituents 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-SEL ECTION CRITERIA
Selection of parameters for use in developing effluent
limitation 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
undesirable tastes and odors in water supplies.
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SIGNIFICANCE AND RATIONALE FOR SELECTION OF POLLUTION
PARAMETERS
pH, 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.
Waters 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, zinc, 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 pH close to 7. This is very
significant for providing safe drinking water.
Extremes of pH or rapid pH changes can exert stress condi-
tions or kill aquatic life outright. Dead fish, associated
algal blooms, and foul strenches are aesthetic liabilities
of any waterway. Even moderate changes from "acceptable"
criteria limits of pH are deleterious to some species. The
relative toxicity to aquatic life of many materials is
increased by changes in the water pH. Metalocyanide
complexes can increase a thousand-fold in toxicity with a
drop of 1.5 pH units. The availability of many nutrient
substances varies with the alkalinity and acidity. Ammonia
is more lethal with a higher pH.
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 ma-
result in eye irritation for the swimmer. Appreciab"
irritation will cause severe pain.
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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; laundries; 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
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
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those benthic 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 inexhaustible food source for sludgeworms and
associated organisms.
Turbidity is principally a measure of the light-scattering
and 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
gravityseparation 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.
Levels of oil and grease which are toxic to aquatic
organisms vary greatly, depending on the type and the
species susceptibility. 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
lubricants and spillage of fuels, as well as the usage of
reagents in many milling processes.
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Chemical Oxygen Demand (COD) and Total Organic Carbon
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
water-quality 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 concentrations that makes them less competitive
and able to sustain their species within the aquatic
environment. For example, reduced DO concentrations have
been shown to interfere 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.
Pish 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.
The total organic carbon (TOC) 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 contamination. Forms of carbon analyzed by this
test, among others, are: soluble, nonvolatile organic
carbon; insoluble, partially 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
377
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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
environmental fates and effects are largely unknown.
Cyanide
Cyanides in water derive their toxicity primarily from
undissociated 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.
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.
378
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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 nonionized 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 presence of dissolved oxygen, is converted to nitrate
(NO3) by nitrifying bacteria. Nitrite (NC>2) , 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 irritation 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
concentration limits, it has been widely recommended that
water containing more than 10 mg/1 of nitrate nitrogen (NO_3-
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
undissociated 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
379
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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-Leute1 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, however, 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 concentrations of aluminum. Where soils are quite
acidic (pH below 5.0), aluminum toxicity to plants becomes
very significant. Aluminum presence is primarily observed
in waste waters from the bauxite-ore mining industry.
Antimony
Antimony is rarely found pure in nature, its common forms
being the sulfide, stibnite (Sb2S_3) and the oxides
cervantite (Sb204J and valentinite (Sb20^) . Any antimony
discharged to natural waters has a strong tendency to
380
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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 byproduct recovery.
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 arsenopyrite (FeS/N FeAs2J .
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 (AsH_3) .
Arsenic trioxide, which also is exceedingly toxic, was
studied in concentrations of 1.96 to UO 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 As^0_3 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 intestinal walls from the air, water, and
food. Arsenic is a normal constituent of most soils, with
381
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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 U
to 12 mg/kg of arsenic trioxide in the topsoil 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
solubilized in mining and milling by oxidation of the ore
and appear in the effluent stream.
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 hydroxide 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 waste waters was detected only in raw-waste
effluents from the mining and milling of bertrandite.
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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.
Cadmium acts synergistically with other metals. Copper and
zinc substantially increase its toxicity. Cadmium is
concentrated 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. It is a common constituent in all zinc ores and
can be expected to be present in most lead-zinc operations
especially those where metals are solubilized.
Chromium
Chromium, in its various valence states, is hazardous to
man. It can produce lung tumors when inhaled and induces
skin sensitizations. Large doses of chromates have
corrosive effects on the intestinal tract and can cause
inflammation of the kidneys. Levels of chromate ions that
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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 documented.
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 precipitation 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.
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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 containing 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.
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
fluoridebearing 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 H 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.
Concentrations of 30 to 50 mg/1 of fluoride in the total
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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 aguifers or from water
contact from 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 micrograms 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
guantities 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 blocked 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
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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 industry
also contributes elevated iron levels through mine drainage.
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 mg/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 solution.
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 cumulative 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 mg/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
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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 a 8-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. 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 consequence, 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
Poly_celis 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.
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On the basis of the literature surveyed, it appears that the
concentrations of manganese listed below are deleterious to
the stated beneficial uses.
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
Manganese concentrations are found in the effluents of iron-
ore, lead, and zinc mining and milling operations and would
be expected from any future operations exploiting manganese
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 organic
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
particularly 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 synthesize 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
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factors 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
waste water 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
distributed in nature. The mean level of molybdenum in the
U.S. has been reported to be 68 micrograms per liter. The
most important water quality aspect of MO is its
concentration in plants with irrigation and subsequent
possible molybdenosis of ruminants eating the plants.
The 96-hour minimum toxic level of fathead minnows for
molybdic anhydride (Mo03) was found to be 70 mg/1 in soft
water and 370 mg/1 in hard water. The threshold
concentration for deleterious effects upon the alga
Scenedesmus occurs at 54 mg/1. EA 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.
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Nickel
Elemental nickel seldom occurs in nature, but nickel
compounds are found in many ores and minerals. As a pure
metal, it is 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
experiments 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 U3 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).
Concentrations 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 waste water in the titanium, rare-earth, mercury, and
uranium.
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 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
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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
contrxbuted 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
concentrations 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.
In soft water, concentrations of zinc ranging from 0.01 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
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species, agef and conditions, as well as with the physical
and chemical 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
longterm 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
operations; and at most ferroalloy-ore mining and milling
sites,
Radiation and Radioactivity
Exposure to ionizing radiation at levels substantially above
that of general background levels has been identified as
harmful to living organisms. Such exposure may cause
adverse somatic effects such as cancer and life shortening
as well as genetic damage. At environmental levels that may
result from releases by industries processing naturally
radioactive materials, the existence of such adverse effects
has not been definitely confirmed. Nevertheless, it is
generally agreed that the prudent public health policy is to
assume a non-threshold health effect response to radiation
exposure. Furthermore, a linear response curve is generally
assumed which enables statistical estimates of risk made
from observed effects occurring at higher exposures to be
applied at low levels of exposure.
The half-life of the particular radionuclides released to
the environment by an industry is extremely important in
determining the significance of such releases. Once
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released to the biosphere, radionuclides with long half-
lives can persist for hundreds and thousands of years. This
fact coupled with their possible buildup in the environment
can lead to their being a source of potential population
exposure for many years. Therefore, in order to minimize
the . potential impact of these radionuclides, they must be
excluded from the biosphere as much as possible.
Plants and animals that incorporate radioactivity through
the biological cycle can pose a health hazard to man through
the food chain. Plants and animals, to be of significance
in the cycling of radionuclides in the aquatic environment
must assimilate the radionuclide, and retain it Such
processes may lead to bioconcentration of the radioactivity
so that the activity per gram of food is greater than the
activity per gram of water. Bioconcentration factors as
great as several thousand have been observed. Even if an
organism is not eaten before it dies, the radionuclides may
remain in the biosphere continuing as a potential source of
exposure.
Aquatic life may assimilate radionuclides from materials
present in the water, sediment, and biota. Humans can
assimilate radioactivity through many different pathways
Among them are drinking contaminated water, and eating fish
and shellfish that have radionuclides incorporated in them
Where fish or other marine products that may accumulate
radioactive materials are used as food by humans, the
concentrations of the radionuclides in the water must be
restricted to provide assurance that the total intake of
radionuclides from all sources will not exceed recommended
-I." v c J_ S •
Naturally occurring radionuclides, particularly of the
uranium-2-38 and thorium-232 series, can be found in
appreciable concentrations in several types of minerals
throughout the country. Radium-226, a member of the uranium
series, is the radionuclide against which the radiotoxicity
ot most other bone seeking radionuclides are compared This
is due to the relatively high dose delivered to bones from
incorporated radium and the wealth of data on the effects of
radium-226 on humans as the result of numerous medical and
industrial exposures. However, other radionuclides in the
uranium and thorium series may be important, particularly if
released into water. These include radium-228, uranium, and
iead-210 and its alpha emitting daughter polonium-210
Radium-228, a member of the thorium series, has been
designated as a radionuclide for which ingestion should be
controlled in proposed drinking water regulations The
isotope lead-210 is of particular interest. Although it is
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a bone seeker, a small fraction of its daughter polonium-210
is released and distributed to soft tissue, where it
concentrates, particularly in the liver and gonads. The
levels of radionuclides other than radium-226 and uranium
present in process streams and treated effluents are
generally not well detailed. consequently, no other
effluent limits are considered at this time. However,
because of their potential public health significance, an
effluent limitation on radium-228, lead-210 and polonium-210
may be warranted in the future.
Radium-226
Radium-226 is a member of the uranium decay series. It has
a half-life of 1620 years. This radionuclide is naturally
present in soils throughout the United States in
concentrations ranging from 0.15 to 2.8 picocuries per gram.
It is also naturally present in ground waters and surface
streams in varying concentrations. Radium-226 is present in
minerals in the earth's crust. Minerals contain varying
concentrations of radium-226 and its decay products
depending upon geological methods of deposition and leaching
action over the years. If ingested the human body
incorporates radium into bone tissue along with calcium.
Some plants and animals also concentrate radium so that it
can significantly impact the food chain.
As a result of its long half-life, radium-226 which is
present in minerals extracted from the earth may persist in
the biosphere for many years after its introduction through
effluents or wastes. Therefore, because of its radiological
consequences, concentrations of this radionuclide need to be
restricted to minimize potential exposure to humans.
Flotation Reagents
The toxicity of organic floation agents—particularly,
collectors and their decomposition products—is an area of
considerable uncertainty, particularly in the complex
chemical environment 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.
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
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TABLE Vl-l. KNOWN TOXICITY OF SOME COMMON FLOTATION REAGENTS
USED IN ORE MINING AND MILLING INDUSTRY
TRADE NAME
Airofloat 25
Aerof loat 31
Aerof loat 238
Aerofloat 242
Aarofroth 65
Aarof roth 71
Aero Promoter
404
Aero Promoter
3477
AROSURF
MG-98A
Oowf roth 250
DowZ-6
DowZ-11
Dow Z 200
Jaguar
M.I.B.C.
-
Superfloc 16
CHEMICAL COMPOSITION
Esuntially aryl dithtophosphoric acid
Essentially aryl dithiophosphoric acid
Sodium di-iecondary butyl
dithiophosphate
Esuntially aryl dithiophosphork acid
Polyglycol typa compound
Mixture of 6-9 carbon alcohols
Mixture of sulfhydryl type compounds
Unknown
Unknown
Chromium salts (ammonium, potassium,
and sodium chromate and ammonium,
potassium, and sodium dichromate)
Copper sulfate
Crasylic acid
Polypropylene glycol methyl ethers
Potassium amyl xanthate
Sodium isopropyl xanthate
Isopropyl ethylthionocarbamate
Based on guar gum
Lime (calcium oxide)
Mathylisobutylcarbinol
Pine oil
Potassium ferr'Ryanide
Sodium ferrocyanide
Sodium hydroxide
Sodium oleate
Sodium silicate
Sodium sulfide
Sulfuric acid
Polyacrylamide
FUNCTION
Collector/Promoter
Collector/Promoter
Collector/Promoter
Collector/Promoter
Frother
Frother
Collector/Promoter
Collector/Promoter
Collector/Promoter
Depressing agent
Activating agent
Frother
Frother
Collector/Promoter
Collector/Promoter
Collector/Promoter
Flocculant
pH modifier and
flocculant
Frothar
Frother
Depressing agent
Depressing agent
pH modifier
Frother
Depressing agent
Activating agent
pH modifier and
flocculant
Flocculant
KNOWN TOXIC
RANGE 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
0.2 to 2.0
10 to 100
10 to 1000
>1000
1 to 100
0.25 to 2.5
1 to 1900
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
Moderate
Moderate
Low
Moderate
Moderate to High
Moderate
Moderate
Moderate
Moderate
Moderate
Moderate
Low
Toxicity Tolerance Level
High
Moderate
Low
O.Omo/Jl
1.0to 1000mg/£
> 1000 mgli
NOTE: Toxic range is a function of organism tested and water quality, including hardness
and pH. Therefore, toxicity data presented in this table are only generally indica-
tive of reagent toxicity. Although the toxicity ranges presented here are based on
many different organisms, much of the data are presented in relation to salmon,
fathead minnows, sticklebacks, and Daphnia.
396
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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,
flexibility, hardness, tensile strength, surface properties,
and other attributes which may affect their respirability,
deposition, 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
asbestos 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 waste water 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
waste water and effluent of tailing ponds at each facility
were examined for the presence or absence of asbestos or
asbestos-like fibers. The method of analysis used for
detection was one based upon published literature and
employed scanning electron microscopy.
Fibers were not detected in any of the samples with the
exception 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 waste waters from mills 1107, 1108, 1109, and 1110
397
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were examined, and no asbestos or asbestos-like minerals
were found.
While the results of the survey indicate the absence of
asbestos 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, however, that a reconaissance evaluation for
asbestos be performed at each iron-ore mining and
beneficiation operation to determine whether possible
asbestos levels of concern are present.
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.
398
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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.
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 waste waters from
mines and mills are not sufficiently high for selenium to be
considered as a harmful pollutant.
Silicates
Silicates may be present in the waste waters from the ore
mining 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
significant 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
399
-------
reduction do not exist, effluent limitations have not been
proposed for this parameter.
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
different 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.
400
-------
TABLE Vl-2. SUMMARY OF PARAMETERS SELECTED FOR EFFLUENT
LIMITATION BY METAL CATEGORY
PARAMETERS
pH (Acidity/Alkalinity)
Total Suspended Solids (TSS)
Chemical Oxygen Demand (COD)
Cyanide
Ammonia
Aluminum
Antimony
Arsenic
Cadmium
Chromium
Copper
Iron
Lead
Mercury
Molybdenum
Nickel
Vanadium
Ztnr
Radium
Uranium
PARAMETERS SELECTED FOR EFFLUENT LIMITATIONS
0»
6
c
0
•
•
+
Copper Ores
•
•
•
•
•
8
6
u
c
N
•o
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(Q
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s
•
•
•
•
•
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6
5
o
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•
•
+
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•
t
Silver Ores
•
4
*
0
•
Bauxite (Al) Ores
•
^
£
0
>
_o
1
5
u.
•
•
•
i
•
•
•
9
•
Mercury Ores
•
•
•
•
Uranium, Radium,
and Vanadium Ores
f
•
•
+
^
4^
•
Metal Ores, Not Elsewhere Classified
1 Antimony
Ores
•
•
•
A
v
1
Is
£6
No separate limitations; zero discharge
1 Platinum
1 Ores
i/i
c £
KO
bdenum
o
E
c.
\
u
3
?
a
»
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W)
re
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0}
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tt
c
(fl
o
CD
•- £
ss
CO —
sy
22;
1 Titanium
Ores
:
t
i
•
^
V
1 Rare-Earth
Ores
No separate limitations; zero discharge
» Zirconium
Ores
E
3
E
re
£
1
' 0
1
a
>.
a
i/i
re
>
c
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£
«
>
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3
C
0
u
1 N
s
0
13
i
V
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&
$
0
401
ft U. S. GOVERNMENT PRINTING OFFICE : 1975 O - 596-127
-------
Pate I*»«e
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