U.S. DEPARTMENT OF COMMERCE
National Technical Information Service
PB-261 052
A STUDY OF WASTE GENERATION, TREATMENT AND
DISPOSAL IN THE METALS MINING INDUSTRY
MIDWEST RESEARCH INSTITUTE
ENVIRONMENTAL PROTECTION AGENCY
OCTOBER 1976
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A STUDY OF WASTE GENERATION, TREATMENT AND
DISPOSAL IN THE METALS MINING INDUSTRY
This final report (SW-L32c) describes work performed
for the Federal solid waste management programs
r under contract No. 68-01-2665
V
and is reproduced as received from the contractor
U.S. ENVIRONMENTAL PROTECTION AGENCY
1976
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BIBLIOGRAPHIC DATA
SHEET
I. Report No.
3. Recipient's Accession No.
4. Title and Subtitle
A Study of Waste Generation, Treatment and Disposal in the
Metals Mining Industry
5. Report Date
October 1976
6.
7. Auchor(s) j)avid Bendersky, Robert E. Gustafson, Charles E. Mumma,
Kenneth R. Walker, and Dennis Costello
8. Performing Organization Kept.
No.
9. Performing Organization Name and Address
Midwest Research Institute
425 Volker Boulevard
Kansas City, Missouri 64110
10. Project/Task/*ork Unit No.
11. Contract/Grant No.
EPA No. 68-01-2665
12. Sponsoring Organization Name and Address
EPA, Hazardous Waste Management Division
Office of Solid Waste Management Programs
401 M Street, S.W.
Washington, D.C. 20460
13. Type of Report & Period
Covered Final
June 1974 to July 197f
14.
15. Supplementary Notes
EPA Project Officer - Allen Pearce
16. Abstracts
The primary objective of the program was to provide SPA with detailed Information concerning the generation,
treatment, and disposal of potentially hazardous wastes In the metals mining and concentrating Industries.
The definition of potentially hazardous wastes used In this study la as follows:
"Any land disposed wastes that contain one or more hazardous substances In concentrations above
that of the land In or on which It Is disposed are considered potentially hazardous."
The metals mining and concentrating Industries covered In this study were categorised by the following Bureau
of the Census Standard Industrial Classification Numbers: 1021 - Copper Ores; 1031 - Lead-Zinc Ores; 1092 -
Mercury Ores; 1094 - Cranium, Radium and Vanadium Ores; and 1099 - Metal Orea—not elsewhere classified - an-
timony, beryllium, platinum, rare earths, tin, titanium, and zirconium.
Potentially hazardous wastes result from waste rock, overburden, and concentrator tailings. All potentially
hazardous land-disposed wastes generated by this Industry are disposed on the property owned by Industry com-
panies. No outside contractors are used and no off-site disposal Is practiced.
Waste disposal and treatment practices are discussed, and estimates are given for the cost of hazardous waste
treatment and disposal at typical facilities.
17. Key Words and Document Analysis.
Metals Mining
Mining Processes
Concentrator Processes
Process Wastes
17o. Descriptors
Hazardous Wastes
Residues
Disposal Treatment Technology
Disposal and Treatment Costs
17b. Identifiers/Open-Ended Terms
17c. COSATI Field/Group
18. Availability Statement
19.. Security Class (This
Report)
_ _ UNCLASSIFIED
1 21. No. of Pages
20. Security Class (This
Page
UNCI.ASSIFIF.D
FORM NTIS-39 (REV. 10-73)
ENDORSED BY ANSI AND UNESCO.
THIS FORM MAY BE REPRODUCED
USCOMM-OC
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This report, has: been, reviewed by. the U.S. Environmental Protec-
tionAgeneyvand; approved for publication. Approval does not signify
thast the contents necessarily re-fleet the views and policies of the
Environmental Protection Agency* nor does mention of trade names or
commercial products constitute end'orsement or recommendation for use
by the U.S. Government.
ii
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PREFACE
This report presents the results of a study of waste generation, treatment,
and disposal in five categories of the metals mining industry. The categories
are SIC (Standard Industrial Classification) Nos. 1021 (Copper Ores), 1031
(Lead and Zinc Ores), 1092 (Mercury Ores), 1094 (Uranium and Vanadium Ores),
and 1099 (Metal Ores, Not Elsewhere Classified).
The study was conducted by Midwest Research Institute (MRI) for the U.S.
Environmental Protection Agency (EPA) under Contract No. 68-01-2665. The
EPA project officer was Allen Pearce of the Hazardous Waste Management
Division, Office of Solid Waste Management Programs.
The principal authors of this report and their areas of responsibility are
David Bendersky, Project Leader; Robert E. Gustafson, Industry Characterization;
Charles E. Mumma, and Kenneth R. Walker, Waste Generation and Characterization;
E. Patrick Shea, Waste Treatment and Disposal Techniques; and Dennis Costello,
Cost Analysis. Other MRI contributors are Patricia Levy, John L. Sealock, Jr.,
and Alan Woodall. Carl Christiansen, University of Missouri School of Mines
and Metallurgy, and Karl C. Dean, formerly with U.S. Bureau of Mines, served
as project consultants.
Many other individuals and organizations contributed to this study. The
following were especially helpful: Sam Morekas and Timothy Fields, Jr., U.S.
EPA, Office of Solid Waste Management Programs, Hazardous Waste Management
Division, for their guidance throughout the project; Brice O'Brien, American
Mining Congress, and Edward C. Bingham, Copper Range Company, for their
assistance in arranging contacts with the mining industry. Various Federal
and State Government agencies provided important data, as referenced in the
report. Finally, the cooperation of the mining companies and their
representatives who provided pertinent information on their operations during
our site visits and via AMC questionnaires is gratefully acknowledged.
Approved for:
MIDWEST RESEARCH INSTITUTE
L. J. Shannon, Director
Environmental and Materials Sciences Division
October 22, 1976
iii
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CONTENTS
Page
EXECUTIVE SUMMARY - 1
Industry Characterization 3
Waste Generation 4
Potentially Hazardous Wastes 9
Waste Treatment and Disposal 12
Cost Analysis 17
References to Executive Summary— 24
I. COPPER ORES (SIC 1021) - - 25
Industry Characterization 25
History of the Industry 25
Domestic Production and Capacity— 26
Number, Location, Age,.and Size of Active Mines and Mills- 26
Employment-- «—- 34
By-Product/Coproduct Relationships---- — -—- 34
Waste Generation and Characterization 37
Waste Generation 37
Mining Processes 44
Concentrator Processes —- 47
Waste Treatment and Disposal 64
Mining Waste Treatment and Disposal 64
Concentrating Process Wastes 64
Waste Disposal Costs - Copper r 71
Waste Disposal Costs, Tailings Pond Using Level I
Technology--— .-.,_—_. 71
Waste Disposal Costs, Tailings Pond Using Level II
Technology .- — -— — .. 71
Waste Disposal Costs, Tailings Pond ,and Wastewater
Treatment Plant Using Level III Technology--—-------- 75
References to Section I — 81
Preceding page Hank
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CONTENTS
II. LEAD-ZINC ORES (SIC 1031) 83
Industry Characterization— 83
History of the Industry
Domestic Production and Capacities 83
Number, Location, Size, and Age of Mines and
Concentrators 88
Employment 90
By-Products and Coproducts 90
Waste Generation and Characterization 94
Quantities and Types of Waste Generated 94
Mining Processes 97
Concentrating Processes 102
Waste Treatment and Disposal 107
Mining Waste Treatment and Disposal • 107
Concentrator Waste Disposal Operation 113
Waste Treatment and Disposal Costs - Lead-Zinc Mining 119
Disposal Cost--Technology Level I"- 119
Disposal Costs--Technology.Levels II and III 119
References to Section II 125
III. ZINC ORES (SIC 1031) 127
Industry Characterization 127
Zinc Mining — 127
Domestic Production and Capacities 128
Number, Location, Size, and Age of Mines and
Concentrators 132
Employment 133
By-Products and Coproducts 133
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CONTENTS
Page
III. (Continued)
Waste Generation and Characterization 135
Mining Processes 138
Concentrator Processes — 138
Waste Treatment and Disposal 141
References to Section III 142
IV. MERCURY ORES (SIC 1092) - 143
Industry Characterization 143
History of the Industry- 143
Domestic Production and Capacities 144
Number, Location, Size, and Age of Mines and Mills 145
Employment 145
By-Products and Coproducts — ------ — 145
Waste Generation and Characterization 145
Waste Generation 145
Mining Processes 148
Concentrator Process 151
Waste Treatment and Disposal 152
References to Section IV 153
vii
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CONTENTS
Page
V. URANIUM-RADIUM-VANADIUM ORES (SIC 1094) - 155
Industry Characterization 155
History of the Industry 155
Domestic Production and Capacities-—— —-— 157
Number, Location, Size, and Age of Mines and Concentrators 165
Employment • 165
By-Product and Coproduct Relationships 168
Waste Characterization 169
Uranium-Vanadium —•• • 169
Waste Characterization—Vanadium Ores 194
Waste Treatment and Disposal—Uranium 202
Open-Pit Mining Wastes < 202
Underground Mining Wastes—; • 209
Concentrator Operation Wastes— —— 213
Cost Ana lysis--Uranium-Radium-Vanadium Ores 219
Disposal Costs—Open-Pit Mine Wastes 219
Disposal Costs—Uranium Tailings Pond, Levels II and III
Technology • 241
References to Section V >• — • 245
VI. MISCELLANEOUS METAL ORES (SIC 1099)- - 249
Antimony ~-.-> ; 249
Industry Characterization-——> —> 249
Waste Generation and Characterization ; 258
Waste Treatment and Disposal-*— •? — 266
Waste Treatment and Disposal Costs.^-Antimony 269
Beryllium - - 271
Industry Characterization > •- • 271
Waste Generation and Characterization-- 272
.Waste Treatment-and Disposal -— 276
.Waste Treatment and Disposal Costs—Beryllium Mines 279
viii
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CONTENTS
Page
VI. (Continued)
Platinum Group Metals 279
Industry Characterization-- — 279
Waste Generation and Characterization 282
Waste Treatment and Disposal 282
Rare Earth Metals 284
Industry Characterization 284
Waste Generation and Characterization 285
Waste Treatment and Disposal 285
Tin 285
Industry Characterization —-- 285
Waste Generation and Characterization • 287
Waste Treatment and Disposal 287
Titanium and Zirconium : 287
Industry Characterization 287
Waste Generation and Characterization 290
Waste Treatment and Disposal 297
References to Section VI- — 298
Appendix A - Active Mine and Concentrator Operations, 1974 299
Appendix B - Mines and Concentrators Visited, Questionnaires
Received, Sample Questionnaire 319
Appendix C - Geology and Mineralogy " 327
Appendix D - Reference Bibliography, Glossary, Units and Conversion
Factors 360
ix
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LIST OF TABLES
No. Title Page
1 Active Mining and Concentrat ing : Opera t ions for 1974 -------------- 3
2 Questionnaires Received or Mines and Concentrators Visited ---- - — 5
3 Composite Statistics for Metals 'Mining 'Industries SICs 1021,
1031, 1092, 1094, and 1099 for 1974 (Metric) .............. ----- 6
4 Quantity of Dry Wastes from Metals 'Mining by State, EPA Regibn, and
SICs .1021, 1031, 1092, 1094, and 1099 (103 TPY) for 1974 (Metric) — - 8
5 Total .and Potentially .Hazardous Waste from Mining and Concentrat-
ing .SICs 102-1, 1031, 1092, 109'4, and 1099 for 1974 (103 TPY)
(Metric) ..... --------------------------- — .... ....... - ................. 10
6 Projected Total and Potentially Hazardous Waste from Mining and
.Concentrating for SICs 1021., 1031, 1092, -1094, and 1099 for
1977 ,(103 TPY) .(Metric) ............ — *- ........... - ........... - 13
7 Projected Total and Potentially Hazardous Waste from Mining and
Concentrating for SICs 1021, 1031, 1092, 1094, and 1099 for
.L9J33 \(103 TPY) (Metric) ..... ----- ^--^ — ^ ...... .. .............. 14
8 Suflsoary pf Costs for Land Disposal of Potentially Hazardous
Wastes from Ore Mining (SICs 1021, 1031, 1094, and
—-^ ------ ....... -------------- ---- "*— "- — — -- — ............ 21
9 Sunpary of Costs for Land Disposal of P6tentially Hazardous
Wastes from Concentrators (SIGs 1021, 1031, 1094, and
1099) ...... - .......... ...... *.***—-. — ...... •- ..... u—;. ........ 22
10 Copper Production, 1963-1983 — ------ -.-*—•— **---«. --- *•_•_****__..• --- — 27
11 U.S. Copper Concentrate Production, 1974*-*- — *- ------- - ------ -- 28
12 Total Copper Mine Production, of Ore by Year------- ----------- - --- 29
13 U.S. Copper Ore Production from Mines by State arid EPA Region,
1974 — ..... .......... — ...... ------ ^---- ........................ 30
14 Projected U.S. Copper Capacity, 1972-1975 — .......... - .......... 31
15 Number, Location, and Size of Active Copper, Mining and
Concentrating' Operations (1974)--- — *- — ---• -------- - ---------- - 33
x
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LIST OF TABLES
No. Title Page
16 Employment at Active Copper Mine Operations (1974) 35
17 Domestic Copper By-Product and Coproduct Relationships (1972)
(Metric) - - 36
18 Total Production Statistics by State and EPA Region for SIC 1021
Copper Ores for 1974 (Metric) 38
19 Concentrator Wastes, Ore Mined and Copper Produced for SIC
1021 (Metric) 39
20 Ratio of Total Waste Rock-Overburden-Concentrator Dry
and Wet Wastes to Ore Mined for SIC 1021 Copper Ores
(Metric) 40
21 Total and Potentially Hazardous Waste from Mining and
Concentrating Copper Ores for SIC 1021 for 1974 (Metric) 41
22 Projected Total and Potentially Hazardous Waste from Mining
and Concentrating of Copper Ores for SIC 1021 for 1977 (Metric)- 42
23 Projected Total and Potentially Hazardous Waste from Mining
and Concentrating Copper Ores for SIC 1021 for 1983 (Metric)- 43
24 Concentrating Processes for Copper Ores 49
25 Analysis of Tailings from a Copper Concentrator---------—----- 50
26 Analytical Data for Tailings Solids for SIC 1021,
Copper 51
27 Major Cost Assumptions, Potentially Hazardous Waste from
Copper Concentrators--Techno logy Level I -—- 72
28 Disposal Costs, Potentially Hazardous Waste from Copper
Concentrators, Tailings Pond, Technology Level I (Expressed
in Equivalent Annual 1973 Dollars) • 74
29 Additional Cost Assumptions—Tailings Pond—Copper
Concentrators--Technology Level II 76
xi
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LIST OF TABLES
No. Title Page
30 Disposal Costs, Potentially Hazardous Waste—Copper Industry—
Tailings Pond, Technology Level II (Expressed in
Equivalent Annual 1973 Dollars)- , „— 78
31 Additional Cost Assumptions—Copper Tailings Pond—Technology
Level III — — 79
32 Disposal Costs, Potentially Hazardous Waste—Copper Concentrator
Tailings Pond and Wastewater Treatment Plant, Technology Level
III (Expressed in Equivalent Annual 1973 Dollars) < 80
33 Lead Production, 1963-1983 84
34 U.S. Lead Concentrate Production for 1974 (Recovered Lead
Concentrate Products.) 85
35 Total Lead Mine Production of Ore by Year •*--. » 86
36 Mine Production of Recoverable Lead in the United States, by
State -.- 87
37 Mine Production of Lead in the United States, by EPA Region
fox 1973 88
38 Number, Location and Size of Active Lead-Zinc Mining and
Gone ent rat ing Op.era t ion,s-—----- -.- - - —- 89
39 U.S. Lead Capacity and Production for 1972—- 92
40 Projected U.S. Lead Production Capacity for 1972-1975—^------- 92
41 Employment at Active Lead-r-Zinc Mine and Concentrator Operations— 93
42 Domestic Lead By-Product and Coproduct Relationships (1972) 94
43 Total Production Statistics by State and EPA Region for Lead-Zinc
and Zinc Ores in 1974 for SIC 1031 (Metric)-., 95
44 Ratio of Total Waste Rock-Concentrator'Dry Wastes and Wastewater
to Ore Mined for Lead-Zinc and Zinc Ores for SIC 1031
(Metric) • • 96
xii
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LIST OF TABLES
No. Title Page
45 Analysis of Tailings from Lead-Zinc Mines and Concentrators 108
46 Analytical Data for Lead-Zinc Tailings Solids for SIC 1031 109
47 Total and Potentially Hazardous Wastes from Mining and
Concentrating of Lead-Zinc and Zinc Ores for 1974 for SIC
1031 (Metric)- 110
48 Projected Total and Potentially Hazardous Waste From Mining
and Concentrating of Lead-Zinc and Zinc Ores for 1977 for
SIC 1031 (Metric) - 111
49 Projected Total and Potentially Hazardous Waste From Mining
and Concentrating Lead-Zinc and Zinc Ores for 1983 for
SIC 1031 (Metric) 112
50 Major Cost Assumption - Tailings Pond, Lead-Zinc Technology
Level I (Coeur d'Alene) 120
51 Disposal Costs, Potentially Hazardous Waste From Lead-Zinc
Tailings Pond, Technology Level I - Coeur d'Alene (in
Equivalent Annual 1973 Dollars) 122
52 Additional Cost Assumptions - Lead-Zinc Mining Technology
Levels II and III 123
53 Disposal Costs, Potentially Hazardous Waste - Lead-Zinc
Mining, Tailings Pond and Wastewater Treatment Plant,
Technology Levels II and III 124
54 U.S. Zinc Production, 1968-1983 (Recoverable Content of Ore) 128
55 U.S. Zinc Concentrate Production - 1974 (Recovered Zinc Con-
centrate Products) 129
56 Total Zinc Mine Production of Ore by Year (Ore Mined) 130
57 Mine Production of Recoverable Zinc in the United States by
State (Metric Tons) 131
58 Mine Production of Recoverable Zinc in the United States by
EPA Region - 1973 132
xiii
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LIST OF TABLES
No. Title Page
59 Number, Location, and Size of Active Zinc Mining and Concen-
trating Operations 133
60 U.S. Zinc Capacity and Production, 1972 (1,000 MT) - 134
61 U.S. Zinc. Production Capacity, 1972-1974 (1,000 MT) 134
62 Employment at. Active Zinc Mine and Concentrator Operations in
1974— — - 135
63 Zinc By-Product and Coproduct Relationships 136
64 Production Statistics by State and EPA:Region for Zinc Ores in
1974 for SIC 1031 (Metric) 137
65 Ratio of Total Waste Rock—Overburden--Concentrator Wastes
to Ore Mined for SIC 1031, Zinc Ore (Metric) 137
66 Mercury Ore Treated in the United States 144
67 Mercury Production - 1968-1973 (Metric Tons) 145
68 Number, Location, and Size of Active Mercury Mining and Mill-
ing Operations (1974) -A 146
69 Production Statistics by State and EPA Region', SIC 1092, for
Mercury (Metric.) • 149
70 Ratio of Total Waste Rock—Overburden^-Concentrator Waste to
Ore Mined,, SIC 1092, for.1 Mercury Ores (Metric) 149
71. Uraaium. Ore Mining. Production in the United States, by State
(Recoverable Content U,0Q Metric Tons) 158
J o
72 Salient Uranium Concentrate (U,0 ) Statistics for the United
States (Metric. Tons U 0' Unless.Otherwise Specified)--^ 160
J o
73 Active Uranium Ore Concentrator Plants- 161
74- Mine Production of Vanadium. Produced in the United States
(MT) — 163
xiv
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LIST OF TABLES
No. Title Page
75 U.S. Vanadium Concentrate Production, 1974 (Recovered Vanadium
Concentrate Products) 164
76 Domestic Uranium Mine Operations and Production Data 166
77 Employment in Active Uranium Mine Operations in 1974 167
78 Total Employment at Domestic Uranium Mines and Concentrators 168
79 Total Production Statistics by State and EPA Region for SIC 1094
(Uranium-Vanadium Ores) in 1974 (Metric) 170
80 Ratio of Total Waste Rock—Overburden—Concentrator Waste to
Ore Mined for SIC 1094 (Uranium-Vanadium Ores) for 1974
(Metric) 171
81 Active U.S. Uranium Ore Concentrator Plants in 1974
82 Distribution of Uranium Ore Concentrator Plants Operated in
1974 by Type of Process and Percent of Total Feed Ore
Capacity — - ----------------------------------- - ------- -._-._
83 Total and Potentially Hazardous Waste From Mining and Concen-
trating Uranium-Vanadium Ores for SIC 1094 in 1974 (Metric)—
84 Projected Total and Potentially Hazardous Waste From Mining
and Concentrating Uranium -Vanadium Ores for SIC 1094 in
1977 (Metric) ................................................ 195
85 Projected Total and Potentially Hazardous Waste From Mining
and Concentrating Uranium-Vanadium Ores for SIC 1094 in
1983 (Metric) ................................................ 196
86 Total Production Statistics, by State and EPA Region for SIC 1094
(Uranium-Vanadium Ores) for 1974 (Metric)— ...... ------ ..... 200
87 Ratio of Total Waste Rock-Overburden-Concentrator Waste to
Ore Mined for SIC 1094 (Uranium-Vanadium Ores) for 1974
(Metric)-- ......... - ......... — ..... - ..... ------------------ 200
88 Major Assumptions of Cost Analysis , Open-Pit Uranium Mine —
Surface Waste Dump — Technology Level I ----------------------- 221
xv
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LIST OF TABLES
No. Title Page
89 Disposal Costs—Potentially Hazardous Waste From Open-Pit
Uranium Mining—Surface Waste Dump, Technology Level I
(Expressed in Annual Costs —1973 (Q4) Dollars) 223
90 Assumptions of Cost Analysis , Open-Pit Uranium Mine—Mine
Backfilling—Technology Level I 225
91 Disposal Costs—Potentially Hazardous Waste From Open-Pit
Uranium Mining—Mine Backfilling, Technology Level I
(Expressed as Annual 1973 (Q4) Dollars) - — 229
92 Major; Cost Assumptions: Technology Levels II and III, Uranium
Open-Pit Mine, Surface Waste -Dump 230
93 Disposal Costs - Potentially Hazardous Wastes From Open-Pit
Uranium Mining - Surface Waste Dump, Technology Levels II
and III (Expressed in Equivalent Annual 1973 Q4 Dollars) 233
94 Major Cost Assumptions—Underground Uranium Mining Surface
Waste Dump—Technology Level I 235
95 Disposal Costs, Potentially Hazardous Waste From Underground
Uranium Mining - Surface Waste Dump,, Level I Technology
(Expressed in Annual Equivalent 1973 Dollars) 236
96 Disposal Costs, Underground Uranium Mining Levels II and III
Technology, Surface Waste Dump (Expressed in Equivalent
Annual 1973 Dollars) - 238
97 Capital Cost Assumptions for Uranium Tailings Ponds (Level I
Technology) in Wyoming and New Mexico (Expressed in 1973
Dollars) 239
98 Operating Cost Assumptions - Tailings Pond (Level I Technology) 240
99 Disposal Costs, Potentially Hazardous Waste From Open-Pit
Uranium Mining—Acid Leach and Alkaline Leach Concentrators—
Levvel I Technology in Wyoming (Expressed in Equivalent Annual
1973 Dollars) 242
100 Disposal Costs, Potentially Hazardous Waste From Open-Pit
Uranium Mining—Acid Leach and Alkaline Leach Concentrators,
Level I Technology in New Mexico (in Annual 1973 Dollars) 243
xvi
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LIST OF TABLES
No. Title Page
101 Land Area of Uranium Tailings Pond Dam and Exposed Beaches
(Expressed in Hectares (Acres)) 244
102 U.S. Production of Miscellaneous Metals - 1974 (Recovered
Metals and Concentrates)- 251
103 Employment in Active Miscellaneous Mining Operations (1974) 252
104 Number, Location, and Size of Miscellaneous Mining, and Milling
Operations (SIC 1099)— — 253
105 World Antimony Production, 1969-1973 (Metric Tons)— - - 254
106 World Production of Platinum-Group Metals, 1968-1973 (Kilograms) 254
107 World Production of Platinum, 1968-1972 (Kilograms) 255
108 Primary Mine Production of Antimony and Antimony Concentrates
in the United States by Year 257
109 Production Statistics by State and EPA 'Region--Miscellaneous
Ores, SIC 1099, for 1974 (Metric) 259
110 Ratio of Total Waste Rock-Overburden-Concentrator Waste to Ore
Mined, SIC 1099, Miscellaneous Ores for 1974 (Metric) 260
111 Total and Potentially Hazardous Waste From Mining and Concen-
trating Miscellaneous Ores, SIC 1099, for 1974 (Metric) 260
112 Projected Total and Potentially Hazardous Waste From Mining
and Concentrating Miscellaneous Ores for SIC 1099, for 1977
(Metric) - - - 261
113 Projected Total and Potentially Hazardous Waste From Mining
and Concentrating Miscellaneous Ores, SIC 1099, for 1983
(Metric) 261
114 Analysis of Tailings From Antimony Mill 268
115 Disposal Costs, Potentially Hazardous Waste From Antimony,
Technology Levels I, IIj and III, Tailings Pond (Expressed
in Equivalent Annual 1973 Dollars) 270
xvii
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LIST OF TABLES
No. Title Page
116 Major Cost Assumptions - Beryllium Mining - Technology Levels
I, II, and III, Tailings Pond .— 280
117 Disposal Cost, Potentially Hazardous Wastes, Beryllium Mining,
Technology Levels I, II, and III,(Expressed in Equivalent
Annual 1973 Dollars) —-— 281
118 U.S. Platinum Group Metal Statistics (Kilograms) .— 283
119 U.S. Production of Tin — .- . — : 286
120 .Production and Mine Shipments of Titanium Concentrates From
Domestic Ores in the United St,ates .- 289
xviii
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LIST OF FIGURES
No. Title Page
1 Location of copper (SIC 1021) mines for 1974 — 32
2 Typical open-pit copper mine 45
3 Typical underground copper mine 46
4 Concentrating of copper 48
5 Concentrating of copper ore, heap leaching and
electrowinning 55
6 Concentrating of copper ore by dump leaching and
precipitation 56
7 Concentrating of copper ore by vat leaching and
electrowinning 58
8 Concentrating of copper ore by in situ leaching and
precipitation 60
9 Level I technology for disposal of copper concentrator
wastes 67
10 Level II technology for disposal of copper concentrator
wastes 68
11 Level III technology for disposal of copper concentrator
wastes 69
12 Location of lead-zinc and zinc (SIC 1031) mines 91
13 Typical underground lead-zinc mine (Missouri Lead Belt) 99
14 Typical underground lead-zinc mine (Coeur d'Alene) 101
15 Mining and concentrating process: lead and zinc ores, Coeur
d'Alene, Idaho district 103
t
16 Mining and concentrating process: lead-zinc-copper ore,
southeast Missouri district 104
xix
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LIST OF FIGURES
No. title Page
17 Level I technology for waste treatment and disposal, lead-
zinc ores (Coeur d'Alene) 114
18 Levels II and III technology for waste treatment and
disposal, lead-zinc ores (Coeur d'Alene)--- • 115
.19 Level I technology for waste treatment and disposal, lead-
zinc ores (southeast Missouri)------ — •—:—---— 116
20 Mining of zinc ore-^ * —_...-—> -^ 139
21 Mining and concentrating of zinc--- ------ 140
22 Location of.mercury (SIC 1092) minds for 1974------ 147
23 Mercury mining and concentrating .process ' 150
24 Mining and concentrating of uranium ore--acid leach process-- 181
25 Mining and concentrating of uranium ore—alkaline leach
process :—•-— • ••=• 187
26 Vanadium mining • • — 198
2.7 Vanadium mining and processing ;-._.w-_-_-_.-— -__> -- 199
28 Level I technology for treatment and disposal of potentially
hazardous wastes in open-pit uranium ore mining and in ore
concentrator operations- :- --• 203
29 Level II/Level III technology for treatment and disposal of
potentially hazardous wastes in open-pit uranium ote
mining and in ore concentrator operatio'ns-------- •-- 207
3,0 Level I technology for treatment and disposal of potentially
hazardous wastes in underground uranium ore mining and i-n
ore concentrator operations---— •- —--- 210
xx
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LIST OF FIGURES
No. Title Page
31 Level II/III technology for treatment and disposal of
potentially hazardous wastes in underground uranium ore
mining and in ore concentrator operations 212
32 Location of miscellaneous ores (SIC 1099) 250
33 Mining and concentrating of antimony 262
34 Waste from antimony operations, technology Levels I, II,
and III - 267
35 Mining and concentrating of beryllium 273
36 Berylliumwastes—technology Levels I, II, and III 277
37 Titanium and zirconium operations--Ti-Zr-13 292
xxi
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EXECUTIVE SUMMARY
This report is the result of a study commissioned by the U.S. Environmental
Protection Agency to assess the waste generation, treatment, and disposal
practices in the metals mining industry. This study is one of a series of
industry studies by the Office of Solid Waste Management Programs, Hazardous
Waste Management Division. The studies were conducted for information
purposes only and not in response to a Congressional regulatory mandate. As
such, the studies serve to provide EPA with: (1) an initial data base
concerning the current and projected types and quantities of industrial
wastes, applicable treatment and disposal technologies and their associated
costs; (2) a data base for technical assistance activities; (3) a background
for guidelines development work pursuant to Section 209 of the Solid Waste
Disposal Act as amended.
The definition of "potentially hazardous waste" in this study was developed
based upon contractor investigations and professional judgment. This
definition does not necessarily reflect EPA thinking since such a definition,
especially in a regulatory context, must be broadly applicable to widely
differing types of waste streams. The presence of a toxic, flammable,
explosive or reactive substance should not be the major determinant of
hazardousness if there are data to represent or illustrate actual effects of
wastes containing these substances in specific environments. Thus, the
reader is cautioned that the data presented in this report constitute only
the contractor's assessment of the hazardous waste management problems in
this industry. EPA reserves its judgments pending a specific legislative
mandate.
The primary objective of this program was to provide EPA with detailed
information concerning the generation, treatment, and disposal of potentially
hazardous wastes in the metals mining and concentrating industries. The metals
mining and concentrating industries covered in this study were categorized by
the following Bureau of the Census Standard Industrial Classifications Numbers;
1021 - Copper Ores; 1031 - Lead-Zinc and Zinc Ores; 1092 - Mercury Ores; 1094 -
Uranium, Radium, and Vanadium Ores; and 1099 - Metal Ores--not elsewhere
classifed - antimony, beryllium, platinum, rare earths, tin, titanium, and
zirconium. The contract specified that particular attention be given to those
wastes which may contain the following materials: asbestos; arsenic; lead;
mercury; halogenated hydrocarbons; pesticides; selenium; and zinc. Other
substances which were believed to be potentially hazardous were also included,
such as carcinogens, chromium, and cadmium.
-------
The definition of potentially hazardous wastes used in this study is as
follows:
"Any land disposed wastes that contain one or more hazardous substances
in concentrations above that of the land in or on which it is disposed
are considered potentially hazardous."
The approach followed in this project was: first, define the industry and
collect and report statistical and historical data on production of ore and
waste from mines and production of metal concentrates and wastes from
concentrators (utilizing secondary sources); second, contact the American
Mining Congress for assistance in obtaining information from the mine and
concentrator operators; third, contact the State Department of Mineral
Resources in each state where the ores of interest were mined and concentrated;
fourth, contact other mining associations, notably the Lead ajnd Zinc Association
and the American Quicksilver Institute; and finally, visit selected mines and
concentrators to obtain all available data on the operation, the quantities and
composition of the various waste materials, and the methods used for land
disposal of the wastes.
The open literature was searched, and the Bureau of Mines, the EPA regional
laboratories and offices, and the U.S. Geological Survey offices and
laboratories were contacted for information on the activities in the mining
and concentrating industries of concern. The U.S. Geological Survey .was contacted
for information on the background concentrations of minerals in and around
the mines. A questionnaire was circulated by the American Minting Congress to
86 mining companies; 45 responses were received covering 64 mines and 52
concentrators. The Lead and Zinc Association was asked to assist in contacting
mining companies for permission to visit their operations. The American
Quicksilver Institute was asked, to assist in contacting mercury mining and
concentrating companies for permission to visit their operations. The relevant
state mineral departments were solicited for information on mining activities
in their states; 10 replied.
All of this information was analyzed, and. representatives of-mining companies
in the five SIC categories were contacted for permission to visit their
facilities. One company refused to grant permission for a visit and also
failed to return the AMC questionnaire.
There were 27 site visits covering 47 mines and 40 concentrators out of a
total of 274 mines and 138 concentrators in the United States. Additionally,
we received questionnaires from 15 companies covering 39 mines and 31
concentrators we had not visited. In all we gathered first hand information
on 86 out of 274 mines, and 71 out of 138 concentrators.
-------
The metals mining industry mines and processes tremendous tonnages of ore
which vary considerably in composition from mine to mine and even within a
single mine. Furthermore, the industry is "easy in - easy out," i.e., quick
to shut down or reopen existing mines in response to changes in profitability
and demand. Statistics in recent years thus bear a definite element of
uncertainty. Projections for future years bear a considerable element of
uncertainty. The statistics in the report generally are reported to more
significant figures than are warranted, mainly to facilitate the tasks of
making percentages add to 100, adding columns and rows to consistent totals,
and maintaining consistency from table to table. The reader is advised to
mentally round off the data as he uses them and to use and interpret them
as representative statistics rather than as precise or exact data.
Industry Characterization
Metals mining is one of the oldest industries in the United States, with
early activities dating back to the 1600"s. Large-scale operations were begun
over 100 years ago. The modern copper mining industry dates from about 1840;
mercury mining started about 1850; and large-scale lead and zinc mining
operations began in the 1860's. A latecomer is uranium mining, which became
prominent with the introduction of the "atomic age" in the 1940's.
Most of the five metal mining industries covered in this study, with the
exception of mercury, are still very active. The copper ore, lead-zinc and
zinc ores, mercury ore, uranium-vanadium ores, and miscellaneous ores
industries had a total of 274 mines and 138 concentrators in operation at
the end of 1974. The breakdown for each industry is given in Table 1. The
uranium-radium-vanadium industry had the largest number of active mines,
followed by copper, lead-zinc, miscellaneous ores, and mercury.
TABLE 1
:
ACTIVE MINING AND CONCENTRATING OPERATIONS FOR 1974
SIC _Ores Mines Concentrators
1021
1031
1092
1094
1099
Total
Copper
(a) Lead-zinc
(b) Zinc
Mercury
Uranium- vanadium
Miscellaneous ores
,57
39
9
2
158
9
274
61
40
7
2
17
11
138
-------
Copper ore production is concentrated in six states (Arizona, Utah,
Montana, New Mexico, Nevada, and Michigan), which produced 99 percent (381
million metric tons (419 million tons)) of the nation's output in 1973.
Domestic lead-zinc and zinc ore production is located in 15 states, with
over 45 percent (8 million metric tons (8.8 million tons)) in Missouri.
Uranium mines are located in six states, with New Mexico and Wyoming
providing more than 75 percent (4 million metric tons (4.4 million tons))
of the output. Vanadium is presently mined only in Arkansas, and mercury
is mined only in California. Miscellaneous metals (antimony, beryllium,
platinum, titanium, zirconium, and rare earths) are mined in small
quantities in eight states, with the largest tonnage in Florida.
The number of returned questionnaires and mines and concentrators visited
(no duplicates) are shown in Table 2.
Composite statistics in 1974 for the five metals mining industries
covered in this study are shown in Table 3. A total of 430 million metric
tons (473 million tons) of raw ore was mined by the five industries, from
which 15 million metric tons (16.5 million tons) of concentrated ore and
miscellaneous concentrates were produced.
The data collected in the site visits and furnished in the questionnaires
were supplemented with data from the U.S. Bureau of Mines Mineral Specialists
and the Engineering & Mining Journal to arrive at the composite statistics
for this industry. Where necessary to complete missing data, engineering
estimates were used. It was assumed that the ore produced, waste generated,
and concentrates produced by the mines and concentrators for which data could
not be found was the average of the values reported to the investigators.
Was te Generation
In the process of mining and concentrating the ores, 769 million metric
tons (847 million tons) of.dry waste, and 1,380 million metric tons (1,518
million tons) of wet waste*were generated .In 1974. The mining wastes
generated consisted of 371 million metric, tons (409 million tons) of waste
rock and 139 million metric tons (153 million tons) of overburden. There
were 259 million metric tons (285 million tons) of dry concentrator wastes
and 869 million metric tons (956 millidn tons) of wet waste.
Region IX was by far the leading generator of metals mining and
concentrating was-tes, with 55 percent (427 million metric tons) of the
national total of dry wastes. Region VIII ranked second with 25 percent
(193 million metric tons), and .Region VI was third with 15 percent (119
million metric tons). These three regions produced 95 percent (738 million
metric tons) of the national metals mining -and concentrating industry wastes
in 1974.
*Wet waste = dry waste + water
-------
TABLE 2
QUESTIONNAIRES RECEIVED OR MINES AND CONCENTRATORS VISITED
Mine type
Ore
type
Copper
Lead-zinc
Zinc
Uranium-rad ium
Vanadium
Miscellaneous
Total
Under-
ground
13
23
7
8
-
1
52
Open- In
pit situ
15 1
-
-
11
3
4 -
33 1
% of
Mines
51
60
78
12
100
55
Concentrator
30
18
5
10
1
7
71
% of
Concentrators
49
40
71
59
100
64
-------
TABLE 3
COMPOSITE STATISTICS FOR METALS MINING INDUSTRIES SICa 1021, 1031, 1092, 1094, AND 1099 FOR 1974 (METRIC)
State Region
Maine I
New Jersey
New York
II
Pennsylvania
Virginia
III
Florida
Kentucky
Tennessee
IV
Illinois
Michigan
Wisconsin
V
Arkansas
New Mexico
Oklahoma
Texao
VI
Missouri VII
Colorado
Montana
Utah
Wyoming
VIII
Arizona
California*
Nevada
IX
Alaska
Idaho
Washington
X
National*
Ore mined
(103 TPY)
48
7,183
1.306
8,489
544
544
1.088
15,422
59
5.180
20,661
294
8,059
468
8,821
363
22,505
213
367
23,448
8,174
1.506
56.471
t. 135,921
1,860
195,758
148.417
22
12.294
160,733
NA
2,067
332
2,399
429,619
Primary
4
172
227
399
54
28
72
388
2
229
619
58
246
15
319
5
628
4
1
638
639
71
366
895
4
1.336
4,877
21
358
5,256
NA
157
2
>59
9,441
Products
Other
metals
3
0
146
146
0
4
4
125
2
0
127
61
0
12
73
0
65
0
0
65
196
44
<0.1
24
0
68
56
14
-
70
NA
126
12
138
890
(103 TPY)
Miscellaneous
0
0
0
0
41
404
445
42
54
3.796
3.892
73
0
0
73
0
0
0
0
9
30
0
<0.l
341
0
341
0
2
0
2
NA
1
0
1
4,784
Hastes ( 103 TPY)
Total
7
172
373
545
95
436
521
555
58
4,025
4,638
192
246
27
465
5
693
4
1
703
865
115
366
1.260
4
1,745
4,933
37
358
5,328
NA
284
14
298
15,115
Waste
rock
1
0
5
5
0
12
12
0
0
101
101
5
7,836
0
7,841
45
54,997
8,165
484
61,691
1,338
37
9.077
15,526
353
24,983
235,868
908
36.287
273,063
NA
394
257
651
371,658
Overburden
0
0
o
0
0
0
0
0
0
0
0
0
0
0
0
862
23,965
181
7.263
32,271
0
0
0
513
72.224
72,737
30,393
487
1,963
32,843
NA
0
1.644
1,644
139,163
Concentrator tailings Total
(dry) (wet) waste
40
125
934
1,059
406
141
547
34
0
781
815
102
7,292
441
7,835
368
21,807
209
367
22,751
7.309
1,394
56 ,099
35.645
1.860
94,998
109,904
166
12.156
122,226
NA
1,719
318
2,037
259,617
160
627
4.670
5,297
2,265
2.091
4,356
374
0
17.395
17,769
510
34,618
2.205
37,333
1,587
70.758
836
2.276
75.457
31.079
9.814
196,433
114,757
5,889
326,893
320,274
986
37,156
358,416
NA
9,620
2.986
12,606
869,366
161
627
4.675
5.302
2.265
2.103
4.368
374
0
17.496
17,870
515
l>2 ,454
2.205
45,174
2,494
147.562
9,182
10.023
169.261
32,417
9,851
205.510
130.796
78.466
424,623
586,535
2,381
25,406
664.322
NA
10,014
4,887
14,901
1,380,187
Ratio of dry 1. of
Waste
to ore
0.85
0.63
0.72
0.71
0.75
0.28
0.51
0.002
-
0.17
0.04
0.4
1.9
0.99
1.8
3.5
4.5
40.2
22
5.06
1.06
0.95
1.15
0.37
40
0.98
2.54
108
4.10
2.66
NA
1.02
6.68
1.8
1.82
Total
Waste to Waste by region
product Dry
5.9 0.01
0.73
2.52
1.95 0.14
4.27
0.35
1.07 0.07
0.06
-
0.22
0.20 0.12
0.56
61.5
16.3
33.7 2.04
255
145
2,140
8.114
169 15.44
10 1.12
12.4
178
40.4
18.609
110 24.97
76.3
3.9
141
80.1 55.52
NA
7.55
159
14.6 0.57
50.9 100
Wet
0.02
0.61
0.50
2.04
4.29
8.68
3.57
37.61
41.23
1.45
100
NA = Nut available.
* Rare earths and platinum not Included.
-------
The quantities of waste generated and land-disposed in 1974 by each of the
five metals mining and concentrating industries are given in Table 4 on a
state, national, and EPA regional basis.
For SIC 1021 (Copper Ores), the total waste amounted to 366 million metric
tons (403 million tons) of waste rock, 44.5 million metric tons (49 million
tons) of overburden, and 241 million metric tons (265 million tons) of
concentrator wastes (tailings).
The copper ore mining and concentrating industry accounted for 85 percent
(651 million metric tons (716 million tons)) of the total national metals
mining and concentrating wastes generated by the five segments examined
during this study.
In SIC 1031 (Lead-Zinc Ores), the total waste quantity in 1974 was 14
million metric tons (15 million tons), consisting of 2 million metric tons
(2.2 million tons) of waste rock, and 12 million metric tons (13 million tons)
of tailings. The lead-zinc and zinc industries produced about 2 percent of the
total national metals mining and concentrating wastes.
California was the only contributor to wastes in SIC 1092 (Mercury Ore) in
1974. There were only eight operating domestic mercury mines in 1974*, and
six were located in California. The total production of waste reported was
1.6 million metric tons (1.7 million tons), consisting of 908,000 MT (1
million tons) of waste rock, 487,000 MT (536,800 tons) of overburden, and
166,000 MT (183,000 tons) of tailings. We were unable to obtain information
from the mercury mines which closed down in 1974. SIC 1092 accounted for
only 0.2 percent of the total national metals mining waste.
The total waste from mineral mining for SIC 1094 (Uranium-Radium-Vanadium
Ores) in 1974 was 103 million metric tons (113 million tons), consisting of 2.3
million metric tons (2.5 million tons) of waste rock, 94 million metric tons
(104 million tons) of overburden, and 6 million metric tons (6.6 million tons)
of tailings. All the waste reported in Table 3 for Arkansas was contributed
solely by the only domestic vanadium mining operation. SIC 1094 accounted for
13 percent of the total national waste from the metals mining industry.
For SIC 1099 (Antimony, Beryllium, Platinum, Rare Earths, Tin, Titanium,
and Zirconium), the total waste generated in 1974 was 1,370,000 MT (1,500,000
tons), of which 1,060,000 MT (1,170,000 tons) was waste rock, and 318,000 MT
(350,000 tons) was tailings. SIC 1099 accounted for only 0.19 percent of the
total national waste from mineral mining and concentrating operations. We were
unable to obtain information from the only platinum mine in the United States;
therefore, the symbol NA is shown for all categories in Alaska.
* Six of these mines closed in 1974, leaving only two operating mercury mines
-------
TABLE 4
QUANTITY OF DP.V WASTES FROM METALS MINING BY STATE, EPA REGION, AND SICs 1021, 1031. 1092, 1094, AND 1099 (103 TPY) FOR 1974 (METRIC) '
co
SIC 1021 wastes
State
Alaska.
Arizona
Arkansas
California
Colorado
Florida
Idaho
Illinois
Kentucky
Maine
Michigan
Missouri
Montana
Nevada
Meu Jers-y
New Mexico
New York
Oklahoma
Pennsylvania
Tennessee
Texas
Utah.
Virginia
Washington
Wisconsin
Wyoming
Ns'tlonal
Region I
Region II
Region. Ill
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
Waste
rock
0
235,868
0
0
0
0
20
0
0
,1
7,836
0
: 9,072
36.287
A
53,87,0
0
8,165
0
23
0
14,515
0
0
0
0
365,657
1
0
0
23
7,836
62,035
0
23,587
272,155
20
Over-
burden
0
30,393
0
0
0
0
0
0
0
0
0
0
. 0
1,963
. n
11,79~3
0
181
0
0
0
181
0
0
0
0
44,511
0
0
0
0
0
11,974
0
181
32,356
0
Tai lines
0
109,904
0
0
0
0
141
0
0
40
7,292
0
56,088
12,156
i
19,233
0
209
0
714
0
35,017
0
0
0
0
240,794
40
0
0
714
7,292
19,442
0
91,105
122,060
141
Total
0
376,165
0
0
0
0
161
0
0
41
15,128
0
65,160
50,406
"
84,896
0
8 , 555
.0
737
0
49,713
0
0
0
0
650,962
41
0
0
737
15,128
93,411
0
114,873
426,571
161
SIC 1031 wastes
Waste Over-
rock burden
0
0
0
0
37
0
289
5
0
0
0
1,338
0
0
•;
b
5
b
0
78
.'.0
66
12.
36
0
0
1,866
0
5
12
78
5
b
1,338
103
0
325
0
0
0
0
0
0
0
0
0
0
0
0
0
0
"
0
0
0
0
0
0
b
0
0
b
0
0
0
0
0
0
0
b
0
0
0
0
Tailings
0
0
0
145
880
. 0
1,400
102
0
0
0
7,309
24
0
12f'
125
934
0
406
67
0
118
141
213
441
0
12,430
0
1,059
547
67
543
"125
7.309
1,022
145
1,613
Total
0
0
0
145
917
0
1,689
107
0*
b
0
8,647
24
0
12s
125
939
0
406
155
. 0
184
153
249
441
0
14,296
0
1,064
595
155
548
125
8,647
1,125
145
1,938
Waste
rock
0
0
0
908
0
0
0
0
0
0
b
0
0
0
rj
0
0
0
0
0
0
0
b
0
0
0
908
0
0
0
0
0
0 •
0
0
908
0
SIC 1092
Over-
burden
0
0
0
487
0
0
0
b
b
0
0
0
0
0
Q
0
0
0
0
0
0
0
b
0
0
0
487
0
0
0
0
0
0
0
0
487
0
wastes
Tailings
0
0
0
16
0
0
0
0
0
0
0
0
0
0
o
0
0
0
0
0
0
0
0
0
0
0
16
0
0
0
0
0
0
0
0
16
0
SIC 1094 wastes
Total
0
0
0
1,411
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
1,411
0
0
0
0
0
0
0
0
1,411
0
Waste
rock
0
0
45
0
0
0
0
0
b
0
0
b
b
0
c
1,127
0
0
0
b
484
38
0
221
0
353
2,268
0
0
0
0
0
1,656
0
391
0
221
Over-
burden
0
0
862
0
0
0
0
0
0
0
0
0
0
0
0
12,172
0
0
0
0
7/263
332
0
1 ,644
0
75,274
94,497
0
0
0
0
0
20,292
0
72,556
0
1,644
Tailings
0
0
368
0
514
0
0
b
0
0
0
0
0
0
o
2/449
0
0
0
0
367
420
0
105
0
1,860
6,083
0
0
0
• 0
0
3,184
0
2,794
0
105
Total
0
0
1,275
0
514
0
0
0
0
0
0
0
0
0
3
15,748
0
0
0
0
8,114
790
0
1,970
0
74,437
102,848
0
0
0
0
0
25,137
0
75,741
0
1,970
Waste
rock
NA
0
0
NA
0
0
85
0
0
0
0
Q
5
0
. 0
0
0
0
0
0
0
961
0
0
0
0
1,051
0
0
0
0
0
0
0
966
0
85
SIC 1099 wastes
Over-
burden
NA
0
0
NA
0
0
0
0
0
0
0
0
0
0
o
0
0
0
b
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Tailings
NA
0
0
NA
0
33
178
0
0
0
0
0
12
0
o
0
0
0
0
0
0
95
0
0
0
0
318
0
0
0
33
0
0
0
107
0
178
Total
NA
0
0
HA
0
33
263
0
0
0
0
0
17
0
G
0
0
0
0
0
0
1,056
0
0
0
0
1,369
0
0
0
33
0
0
0
1,073
0
263
NA = Not available.
* Tailings sold as agricultural lira
-------
Potentially Hazardous Wastes
A determination was made as to how much of the waste being generated by
the five metal mining industries should actually be considered potentially ,
hazardous, taking into account available data on the following factors:
1. The constituents in the wastes.
2. Geological and mineralogical data.
3. The solubility of the metals and minerals in the waste.
4. The concentration of the metals and minerals in the waste.
5. The presence or absence of pyrite, which could cause acid leaching.
6. Background values of hazardous materials.
Table 5 presents summary data on the quantity of potentially hazardous
wastes generated by all mining and concentrating in each state in 1974. In
preparing these data, an identification was made as to the wastes from the
mining and concentrating operations which normally contain hazardous
substances in concentrations above the background levels of the land in or on
which they are disposed. Such wastes were considered to be potentially
hazardous. In most instances, only the concentrator tailings were found to
contain hazardous materials above background levels, and therefore considered
to be potentially hazardous waste. However, for uranium mining and concentrating,
both mine waste rock and concentrator tailings were considered to be potentially
hazardous.
For the SIC 1021 (Copper Ores) category, we concluded that the only
potentially hazardous wastes are the tailings from the concentrator
operations. These tailings can contain small amounts of hazardous materials
at concentrations above disposal area background levels. Furthermore, these
tailing wastes are vulnerable to water leaching (if pyrite is present) and
wind blowing because they consist of finely divided solids. In contrast,
the mining wastes (overburden and waste rock) do not contain concentrations
of hazardous materials above the background levels for the soil and rock on
which they are disposed, and are therefore not considered potentially
hazardous.
The national total of dry potentially hazardous waste from SIC 1021 amounted
to 140 million metric tons (154 million tons) in 1974. This waste consisted
entirely of tailings generated in the ore concentration processes. The wet
weight of potentially hazardous wastes was 461 million metric tons (508
million tons).
-------
TABLE 5
TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING SICs 1021, 1031, 1092, 1094, AND 1099 FOR 1974 (103 TPY) (METRIC)
State
Maine
New York
Virginia
Tennessee
Illinois
Michigan
Wisconsin
New Mexico
Oklahoma
Texas
Missouri
Colorado
Montana
Utah
Wyoming
Arizona
California*
Nevada
Idaho
Washington
National*
Total
process
Region - waste'
I 41
II 939
III 153
IV 737
107
15,128
441
V 15,676
100,769
8,555
8,114
VI 1l7/i38
VII 8,647
1,431
65,200
51,684
74,437
VIII 192,752
376,165
145
__50,406
IX 426,716
2,144
2,219
X 4,363
767,462
Waste rock
Total potentially
hazardous wastet
40
934
141
714
102
7,292
441
7,835
16,201
209
851
17,261
7,309
1,394
36,491
23,426
2,213
63,524
52,167
145
7,900
60,212
1,750
539
2,289
160,259
Total
1
5
12
23
5
7,836
0
7,841
54,997
8,165
484
63,fi46
1,338
37
9,077
15,526
353
24,993
235.868
0
36.287
272,155
394
257
651
370,665
Potentially
hazardous
0
0
0
0
0
0
0
0
1,127
0
484
1,611
0
0
0
38
353
391
0
0
0
0
0
221
221
2,223
Overburden *
Total
0
0
0
0
0
0
0
0
23,965
' 181
7,263
31,409
0
0
0
513
72,224
72,737
30,393
0
1,963
32,356
0
1,644
1,644
138,146
Potentially
hazardous
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Concentrator tailings
Dry
Total
40
934
141
714
102
7,292
441
7,835
21,807
209
367
22,383
7,309
1,394
56,123
35,645
1,860
95,022
109,904
145
12,156
122,205
1,750
318
2,068
298,611
weight
Potentially
hazardous
40
934
141
714
102
7,292
441
7,835
15,074
• 209
367
15,650
7,309
1,394
36,491
23,388
1.860
63,133
52,167
145
7,900
60,212
1,750
318
2,068
158,036
Wet
weight § 7. Potentially harardi-us
Potentially waste In region
Total
160
4,670
2,091
7,483
510
34,618
2,205
37,333
70,758
836
2.276
73,870
31,079
9,814
196,553
114,337
5,889
326,593
320,274
965
37,156
358,395
9,973
2,986
12,959
854,633
hazardous Dry
160 0.02
4,670 0.58
2,091 0.09
7,483 0.45
510
34,618
2,205
37,333 4.89
48,727
836
2.276
51,839 10.77
31,079 4.56
9,814
127,841
75,212
5,889
218,765 39.64
152,021
965
24,147
177,133 37.57
9,973
2.986
12,959 1.43
543,512 100
Wet
0 . 0 .»
0.8t>
0.3S
1.38
6.87
9.S4
5.72
40.25
32.59
2.38
100
* Rare earths and platinum not Included.
t Dry weight.
* Dry weight = wet weight.
} Wet weight = dry weight plus water.
-------
In SIC 1031 (Lead-Zinc Ores), our analysis of the waste characteristics
indicates that the concentration of hazardous materials in the mining
wastes (waste rock and overburden) does not usually exceed the background
concentrations in the native soil and rock on which they are disposed.
Therefore, these mining wastes are not judged to be potentially hazatdous.
On the other hand, the amount of hazardous materials in the concentrator
tailings is commonly in excess of background values; and, therefore, most
tailings from processing lead-zinc ores are classified as potentially
hazardous wastes.
The total amount of dry tailings (potentially hazardous waste) generated
by SIC 1031 in 1974 was 11.8 million metric tons (13 million tons). The wet
weight of potentially hazardous wastes was 61 million metric tons (67 million
tons). SIC 1031 accounted for 7 percent of the total national potentially
hazardous waste from metals mining in 1974.
In SIC 1092 (Mercury Ores), there are no potentially hazardous wastes
generated.
The uranium-radium-vanadium ore industries (SIC 1094) produced potentially
hazardous waste consisting of all uranium waste rock and tailings. The uranium
wastes generally contain significant amounts of radioactive substances, such
as uranium minerals, and in some cases, trace amounts of potentially hazardous
materials such as thorium, radium, arsenic, and other hazardous metals.
The total potentially hazardous wastes generated by SIC 1094 in 1974 were
7,938,000 MT (8,750,000 tons), or 5 percent of the total dry national
potentially hazardous wastes. This waste was comprised of 2,223,000 MT
(2,450,000 tons) of waste rock from the mining uranium ore and 5,715,000 MT
(6,300,000 tons) of dry uranium concentrator tailings. The wet weight of
uranium concentrator tailings was 17 million metric tons (19 million tons)
of potentially hazardous waste, which is 3.2 percent of the total national
wet waste from this industry.
In the miscellaneous ores industry (SIC 1099), the dry weight of potentially
hazardous wastes in 1974 consisted of 189,000 MT (208,000 tons) of antimony
concentrator wastes in Idaho and Montana, and 90,000 MT (99,000 tons) of
beryllium concentrator wastes in Utah. The wet weight of potentially hazardous
wastes was 3.3 million metric tons (3.6 million tons). This total excludes the
rare earth wastes in California, for which no data could be obtained. The
percentage contribution of SIC 1099 to the total national dry potentially
hazardous waste in 1974 was 0.19 percent, and to the total wet weight was
0.68 percent.
11
-------
Tables 6 and 7 contain our projections for the total and potentially
hazardous wastes to be generated and land disposed in 1977 arid 1983. To
arrive at the numbers in the tables, we used the projected mine production
or metal demand data furnished by the U.S. Bureau of Mines.-' The assumption
was made that the ratio of waste to ore would reamin relatively constant
during this period. We further assumed that the ore grade and concentrating
method would also remain as they are. Our evaluation of this industry leads
us to believe that the mining and concentrating methods in use today will
also be used in the immediate future.
We do not, however, believe that ore grades will remain constant, .or that
the ratio of waste to ore will remain constant. Ore grades are constantly
changing, even within a mine. The higher grade ores will be depleted before
the lower grade ores are processed. Another factor that influences the waste
generated is the market price for the products of the mine and concentrator.
As the price for the product falls, metal mines and concentrators are closed
down. There are over 1,600 'abandoned mines and concentrators in these SIC
categories. Some of these were abandoned because the ore was depleted, the
ore grade dropped below the economic break-even point, the market price of
the product dropped below the profitable level, or transportation of raw
materials and products was too expensive.
As shown in Table 6, the total potentially hazardous dry land-disposed
wastes in 1977 are projected to be 187 million metric tons (206 million tons).
The wet wastes are projected to be 632 million metric tons (695 million tons).
These quantities represent a 16 percent increase over 1974.
As shown in Table 7, the total potentially hazardous land-disposed dry
wastes in 1983 are projected to be 260 million metric tons (286 million tons).
The wet wastes are projected to be 851 million metric tons (938 million tons).
This represents a 58 percent increase over 1974, and 36 percent over 1977.
Waste disposal from air and water pollution control devices in the mining
and concentrating of these ores did not add to the waste burden in 1974. All
material collected in the air pollution control devices in the grinding of
the ores was returned to the process for recovery of metal values.
In 1977 and 1983, no change in the operation of air pollution control devices
is envisioned. All material collected will be returned to the process for
recovery of metal values. However, there will be wastes disposed from water-
pollution-control devices in 1977 and 1983. It is MRI's estimate that these
quantities will not exceed 1 percent of the total wastes disposed on land.
t
Waste Treatment'and Disposal
The waste treatment and disposal practices for potentially hazardous wastes
from mining and concentrating of the metal ores of interest are described in
this report.
12
-------
TABLE 6
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING FOB SlCa 1021. 1031, 1092, 1094, AND 1099 FOR 1977 (103 TPY) (METRIC)
Concentrator tailings
State
Maine
New York
Virginia
Tennessee
Illinois
Michigan
Wisconsin
New Mexico
Oklahoma
Texas
Missouri
Colorado
Montana
Utah
Wyoming
Arizona
California *
Nevada
Idaho
Washington
National*
Total
process
Region waste^
I 46
II 1,202
III 195
IV 833
137
17.095
564
V 17,796
121.966
9.667
13.331
VI 144.964
VII 11,068
. 2,018
73.648
58,767
122.300
VIII 256.733
425.066
186
56.959
IX 482.211
2,673
3.556
X 6,229
921,277
Waste rock*
Total potentially
hazardous wastet
45
1.196
180
807
131
8,240
564
8,935
20,160
236
1.398
21,794
9,356
1,971
41,207
26,717
3.636
73,531
58,949
186
8.927
68.062
2,185
809
2.994
186,900
Total
1
6
15
26
6
8,855
0
8,861
62.725
9,226
795
72,746
1,713
. 47
10,256
17,509
' 580
28.392
266,531
0
41.000
307,531
488
409
897
420,188
Potentially
hazardous
0
0
0
0
0
0.
0
0
1.852
0
795
2,647
0
0
0
62
580
642
0
0
0
0
0
363
363
3,652
Overburden *
Total
0
0
0
0
0
0
0
0
33,325
205
11.933
45,463
0
0
0
750
118.664
119.414
34.344
0
2.218
36.562
0
2.701
2,701
204,140
Potentially
hazardous
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Dry
Total
45
1,196
180
807
131
8,240
564
8,935
25,916
236
603
26,755
9,356
1.971
63,391
40,505
3.056
108,923
124.192
186
13.736
138,114
2,185
446
2,631
296.942
weight
Potentially
hazardous
45
1.196
180
807
131
8,240
564
8,935
18,308
236
603
19,147
9,356
1,971
41.207
26.655
3.056
72,889
58,949
186
8.927
68,062
2,185
446
2.631
183,248
Wet
weight) 1 Potentially hazardous
Potentially waste In region
Total
181
5,978
2,676
8,456
653
39,118
2.822
42.593
83.645
945
3.739
88.329
39.781
12,959
221,961
129.698
9.676
374.294
361.910
1.235
41.986
405,131
12,237
3.918
16.155
983,574
hazardous Dry
181 0.02
5,978 0.64
2,676 0.10
8,456 0.43
653
39,118
2.822
42,593 4.78
58 , 750
945
3.739
63.434 11.66
39,781 5.01
12,959
144,316
85,487
9.676
252.438 39.34
171,784
1,235
27.286
200,305 36.42
12.237
3.918
16.155 1.60
631,997 100
Wet
0.03
0.95
0.42
1.34
6.74
10.04
6.29
39.94
31.69
2.56
100
* Hare earths and platinum not Included.
t Dry weight.
t Dry weight « wet weight.
5 Wet weight = dry weight plus water.
-------
TABLE 7
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING FOB SICa 1021, 1031, 1092, 1094, AND 1099 FOB 1983 (103 TPlf) (METRIC)
State
Maine
New York
Virginia
Tennessee
Illinois
Michigan
Wisconsin
New Mexico
Oklahoma
Texas
Missouri
Colorado
Montana
Utah
Wyoming
Arizona
California*
Nevada
Idaho
Washington
National*
Total
process
Region waste *
I 60
II 1.324
III 216
IV 1,094
151
22,390
622
V 23,163-
195.668
12,660
35.966
VI 244,314'
VII 12,192
3,573
96,490-
78,514
330.128
VIII 508,705
556,724
204
74.600
IX 631.528
3,019
9.088
X 12.107
1.434,703
Waste rock*
Total potentially
hazardous waste t Total
59
1.317
199
1,060
144
10,792
622
11,558
34,535
308
3.775^
38.618
10.306
3.521
54,002
35.989
9.815
103,327
77,207
204
11.692
89,103
2,467
1.746
4,213
259,760
1
7
17
34
7
11., 598
' 0
11,605
84,728
12,084
2.147
98,959'
1.887
52
13,433
22,814
1.566
37,865
349,085
0
53.705
402.790
553
Ij031
1,584
554,749
Potentially
hazardous
0
0
0
0
0
0
0
0
4,998
0
2.147
7 , 145
0
0
0
169
1.566
1,735
0
0
0
0
0
980
980
9,860
Overburden *
Total
0
0
0
0
0
0
0
0
71.438
268
32.211
103,917
0
0
0
1 , 740
320.313
322,053
44,980
0
2.905
47,885
0
7.291
7,291
481,146
Potent tally
hazardous
0
0
0
0
0
0
0
0
0
0
0'
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Dry weight
Total
59
1,317
199
1,060
144
10.792
622
11,558
39,502
308
1.628'
41.438
10,306
3,521
83,057
55.086
8.249
149.913
162.658
204
J7.990
180,852
2,467
766
3,233
399,935
Wet weight' t Potential
Potentially
hazardous' Total '
59
1.317
199
1,060
144
10,792
622
11.558
29,537
308
1.628
31-.473
10.306
3,521
54.002
35.820
8.249
101.592
77.207
204
11.692
89,103
2,467
766
3.233
249,900
237
6.585
2.948
11.075
719
51,235
3.109:
55,063
125,136
1.237
10.094'
136.467
43.821
17.141
290', 853
171.624
26.118
505.736
474,000
1.361
54.990
530,351
13,993
5.006
18,999
1,311.282
Potentially waste 1
hazardous Dry
237 0.02
6,585 0.51
2.948 0.08
11,075 0.41
719
51.235
3.109
55.063 4.45
92.531
1.237
10.094
103,862 14.87
43,821 3.97
17,141
189,159
113,720
26 . 1 18
346,138 39.78
224.991
1,361
35.738
262,090 34.30
13,993
5.006
18.999 1.62
850.818 100
ly hazardous
n region
•Wet
0.03
0.77
0.35
1.30
6.47
12.21
5.15
40.68
30.80
2.23
100
* Hare earths and platinum not included.
t Dry weight.
4 Dry weight « wut weight.
J Wet weight = dry weight plus water.
-------
All potentially hazardous land-disposed wastes generated by this industry
are disposed of by this industry on their property. No outside contractors
are used and no off-site disposal is practiced.
The waste disposal practices for the potentially hazardous land-disposed
wastes are described at three levels of technology. These levels are:
Level I - Technology currently employed by typical facilities, i.e.,
broad average present treatment and disposal practice.
Level II - Best technology currently employed. Identified technology at
this level must represent the soundest process from an
environmental and health standpoint, currently in use in at
least one location. Installations must be commercial scale;
pilot and bench scale installations are not suitable.
Level III - Technology necessary to provide adequate health and environmental
protection. Level III technology may be more or less sophisticated
or may be identical with Level I or II technology. At this level,
identified technology may include pilot or bench scale processes
providing the exact stage of development is identified.
The only potentially hazardous land-disposed waste from mining of these ores
is waste rock from uranium mining. Level I technology for the disposal of waste
rock, from open-pit uranium mines is utilized by 90 percent of the operating
mines and is briefly described in this section; a more detailed explanation
is contained in Section VI. Briefly, the waste rock is co-mingled with the
overburden and 93 percent (1.2 million metric tons (1.3 million tons)) is
discarded on a surface waste dump. The dump surface is contoured and stabilized
by vegetation. The rest of the waste rock, 7 percent (93,000 MT (102,000 tons)),
is used as mine backfill.
Level II and III technologies, which are the same for disposal of waste rock
from open-pit uranium mines, are the same as Level I with the following
additions: the final dump is covered with soil and vegetation, and the active
dump is periodically covered with overburden or soil.
Level I technology for the disposal of potentially hazardous waste rock from
underground uranium mines is to pile the rock on a surface waste dump, contour,
and allow natural vegetation to stabilize the dump. All underground mines are
using Level I technology and about 800,000 MT (880,000 tons) are disposed of
this way.
15
-------
Level II and III technologies are practiced by 10 percent (12) of the
uranium mines. To meet Level II and III technology requirements, the waste
dump must be contoured, covered with a layer of topsoil or approved subsoil,
vegetated by seeding, and fertilized to maintain a stabilizing plant cover.
About 100,000 MT (110,000 tons) of waste rock are disposed in this manner.
There is one universal method for the disposal of concentrator tailings
from copper, lead-zinc, beryllium, antimony, uranium, cadmium, and vanadium
processes, and that is the use of a tailings pond for the land disposal of
potentially hazardous wastes. Therefore, Level I technology for land disposal
of all potentially hazardous concentrator wastes is the use of a tailings
pond for all the metal concentrators encompassed in this study.
Level II and III technologies will vary for the different concentrator
wastes. Level II technology for the land disposal of potentially hazardous
wastes from copper concentrators is to compact and vegetate the dike or dam
surrounding the tailings pile. At the present time, three copper companies
are utilizing Level II technology for the disposal of potentially hazardous
concentrator tailings. They are disposing of about 60 million metric tons
(65 million tons), which is about 43 percent of the total potentially
hazardous waste from copper mining and concentrating.
Level III technology for the disposal of copper concentrator wastes is to
add chemicals to the effluent from the tailings pond to precipitate the metal
ions, which are then returned to the tailings pond. This practice is at
present followed by one operator, and a treatment plant is being built at a
second concentrator. Both treatment plants will treat about 23 percent of
the total effluent water from tailings ponds at copper concentrators.
Level II and III technology for the disposal of potentially hazardous
wastes from concentrating lead-zinc ores is the addition of a wastewater
treatment plant to precipitate the minerals from the effluent and return
the solids to the pond. The process has been piloted by at least two companies,
Level II and III technologies for disposal of potentially hazardous wastes
from uranium concentrating operation's consists of adding the zero discharge
(no effluent), and special treatment to avoid erosion losses of tailings.
Level II and III technologies for antimorty and beryllium are the same as
Level I. There are only one beryllium plant and two antimony plants.
There were no potentially hazardous wastes generated in the mining and
concentrating of zinc, mercury, and titanium-zirconium ores, and therefore,
no technology levels were identified. Tin is no longer mined in the United
States, and we were able to learn nothing about the mining or concentrating
of the platinum group ores or the rare earth ores.
16
-------
Cost Analysis
Representative plants for all categories of mining and concentrating, in
which potentially hazardous wastes were generated, were used for the purpose
of the cost analysis. The capacities, handling rates, ore grade, operating
hours and other items listed in the plant specifications were selected to
represent average size plants in most instances. The model plant for each
ore mined and concentrated is representative of the particular ore category;
it is not meant to represent the entire industry. In copper, lead-zinc, and
uranium-ore mining and concentrating, there are significant variations in
size. The process used for copper and lead-zinc ores is flotation. There are
some differences in flotation plants, depending on the other metals present
in the ore. However, the basic process, equipment, and operating
characteristics are similar. There are two main processes for concentrating
uranium ores, and a representative plant was selected for each process.
The scope of the present study prohibited a detailed parametric study of
the economic impact as a function of plant size and configuration, and the
model plants used in the analysis present only a simplified flow diagram for
each type of facility. As a consequence, there are some weaknesses and
shortcomings in the model plant concept. Because of the limitations imposed
by the model plant approach, the results of the economic analysis should be
viewed as indicative of the probable economic impact, and not the absolute
impact.
The cost analysis was conducted from the perspective of a mining
corporation. Only those costs that the corporation pays directly (or that
are "realized" by the firm) are included. External social costs, such as
the value of the pollution emitted, are excluded.
The future effects of inflation are also ignored in the analysis. Keeping
all estimates in terms of constant 1973 dollars enhances comparisons and
avoids the necessity of predicting trends in inflation. However, any changes
in the "real" (or noninflated) value of capital, labor, or resources are
included.
The differential impact of taxes has also been excluded in the methodology.
In other words, the costs calculated represent "before tax" estimates. Taxes
enter the calculation as a separate cost item, and are expressed as a
percentage of invested capital. Handling taxes in this manner helps keep the
analysis easily comprehensible to a wide variety of readers. The method
should also enhance comparisons with other EPA studies and private sector
cost estimates. The discussion below presents details of the methodology.
Capital costs include all buildings, equipment, land, etc., that do not
change with small changes in the amount of hazardous wastes disposed. These
costs are often termed fixed costs.
17
-------
The cost of capital equipment can be calculated in a variety of ways and
with an even wider variety of assumptions. For this study, it was assumed
that all capital is financed through the sale of private debt. The cost of
financing with debt is assumed to be 10 .percent. The 10 percent value was
chosen because it was approximately the cost of private debt in 1973, and
it has been used in numerous other EPA studies.— The use of a discount rate
accounts for the interest cost of debt. Therefore, interest costs do not have
to be calculated separately.
The period of .analysis is equal to the useful life of the longest-lived piece
of capital equipment (usually the tailings pond). Although costs are expressed
in terms of dollars per metric ton of potentially hazardous wastes, the period
of analysis is still important. If the period of analysis is short, relatively
large capital cost items will be undervalued. This undervaluation usually
occurs because the salvage value of the long-lived equipment understates the
equipment's remaining usefulness. Therefore, using a short period of analysis
tends to favor less capital-intensive technologies and equipment.
The nominal cost of the capital equipment, excluding land, is allocated
over its useful life by the use of straight-line depreciation. It is unlikely
that the true value of the asset will decrease by equal amounts each period.
Nonetheless, simplicity and comparability with other EPA reports support the
use of the straight-line method.
The levelized annual capital costs and the annual operation and maintenance
costs, and the yearly .fuel and power costs are summed. This sum represents the
total yearly cost of operating a given hazardous waste disposal facility (using
a prespecified technology and capacity size). The yearly quantity of hazardous
waste disposed by "the representative facility is obtained. The cost of disposal
can then be expressed in dollars per metric ton of potentially hazardous waste.
The technique described above is used in all the cost analyses made in this
study. Important "assumptions are listed with the cost estimates. More than one
cost estimate was .made in most cases to demonstrate the range of results and
the dependence of .those results on input assumptions. In addition, the costs
were aggregated to the .entire industry level. The aggregated estimate was then
compared to the costs of production for that industry.
The analyses and estimation-of costs.associated with the handling and disposal
of wastes in the various metal mining industries relied heavily upon information
provided by respondents to the AMC-mailed questionnaires and field interviews.
Cost data received from the responding companies, .along with information
regarding their waste quantities, characteristics, treatment processes, and
handling techniques, provided the basis for developing standard unit costs
for the various functional operations involved in waste disposal.
18
-------
The raw cost data were related to the specific disposal operations,
categorized generally as follows:
1. Wastes from the mining operation
a. Dump for leaching or further processing
b. Dump for disposal
c. Leave in place
d. Use as surface (pit) backfill
e. Return underground as mine backfill
2. Wastes from the concentration operations
a. Transfer to sand plant for separation
b. Discharge to tailings pond
c. Return underground as mine backfill
d. Use as surface fill
e. Filter liquids
The unit disposal processes associated with each typical mining operation
(classified by type of ore, metal and mining method) were identified and
disposal costs estimated on a process-by-process basis. Flow sheets,
identifying the flow of all materials from the time the ore left the mine
until the metal or concentrate was shipped, were employed in identifying and
quantifying the appropriate disposal operations.
All cost estimates were prepared on the basis of raw material leaving the
mine. Thus, waste rock and overburden left in place, either in an open pit
or underground, were considered to be a cost of mining rather than a waste
disposal cost.
Wide variations were reported in disposal costs by the various respondents;
this is to be expected because of the widely differing conditions under which
wastes are disposed. For example, surface dumping of waste rock, the most
common disposal method, can vary from as little as $0.10/ton to more than
$1.00/ton, depending upon the distance the waste must be hauled to the dump
site. Discharging tailings as a slurry into a pond, the most common method
of waste disposal from the concentration operation, is also subject to wide
variations, again depending on the overall facilities layout and the amount
of pumping force required. Median figures were used as much as possible, with
the same unit costs applied to the same unit operations, regardless of the type
of ore being handled.
19
-------
The other major factor causing variations in waste disposal costs was simply
the quantity of wastes associated with the different mining operations and
mining methods. These variations are especially apparent when comparing disposal
costs among the mining industries.
Finally, to reflect the cost impact of waste disposal in each industry, the
disposal costs were related to the quantity of final output attained in each
typical operation, expressing the disposal costs in dollars per ton of metal
or concentrate produced or shipped.
The total costs of waste disposal associated with typical operations in
each metals mining industry were estimated. The cost estimates were prepared
by applying appropriate unit costs for each disposal function to the quantity
of waste handled in each operation. Disposal costs were broken down between
wastes from the mining operations and wastes from the concentration operations.
All cost estimates were prepared on the basis of quantities of ore leaving the
mine and expressed as costs per million metric tons of potentially hazardous
waste.
Wide variations were found among the different industries, depending on the
ore, mining methods, quantity of waste materials handled, and unit operations
employed. The costs for disposal of potentially hazardous wastes from ore
mining are summarized in Table 8. Level I, II, and III technologies are
displayed in this table. The costs are expressed in dollars per metric ton
of potentially hazardous wastes, and the total cost to the industry. The
cost is also related to the value added for each type of mine. The costs
for disposal of potentially hazardous wastes from concentrators are summarized
in Table 9. Level I, II, and III technologies costs are displayed in this
table.
The following paragraphs summarize the major conclusions of this study.
1. The five metals mining industries covered in this study generate large
quantities of waste which are land-disposed. These wastes usually consist of
overburden, waste rock, and tailings. The copper ore industry produces the
largest amount of these wastes,, followed in descending order by uranium-
vanadium, lead-zinc, mercury, and miscellaneous ores.
2. In most cases, the overburden and waste rock do not contain potentially
hazardous materials in concentrations greater than the land on or in which
they are disposed. Therefore, these wastes generally do not add to the
potential hazards of the area. The exception is uranium ore waste rock,
which usually contains radioactive materials (uranium and radium) above
background levels.
3. Wastes from ore concentration processes (tailings) may contain
potentially hazardous substances in higher concentrations than the land on
which they are disposed, and are therefore considered potentially hazardous.
Furthermore, these wastes are fine-grained*and subject to wind and water
erosion unless properly treated.
20
-------
TABLE 8
SUMMARY OF COSTS FOR LAND DISPOSAL OF POTENTIALLY HAZARDOUS WASTES FROM ORE MIMING
(SICs 1021, 1031, 1094. AND 1099)
ro
Ore mined Type of mine
Uranium Open Pit
Uranium Open Pit
Uraplum Underground
-
Disposal
mode
Surface
Dump
Mine Back-
fill
Surface
Dump
Level I
Range of
cost
($/MT)
$0.004-
$0.25
$0.16-
$0.21
$0.04-
$0.25,
technology
Total cost
to
Industry*
($U)6/yr)
$.008-$. 47
$.023-
$.030
$.008-
$.05
Levels II and III technology
% of value
added In Disposal
raining* mode
0%-.29% Surface
Dump
.01%-. 02% N.A.
0%-.03% Surface
Dump
Range of
cost
($/MT)
$0.20-
$0.24
N.A.
$0.19-
$0.44
Total cost
to
Industry*
(SlO<>/yr)
$.38-$. 45
$.038-
$.088
% of value
added In
mining*
.2 3%-. 28%
.02%-. 05%
N.A. - Not Applicable
* Assuming all mines of that type and disposal mode utilize the stated technology level. Baaed on waste disposal levels In 1974.
t 1972 Value Added In Mining adjusted to 1973 dollars using wholesale price Index. Sources: 1972 Census of Mineral Industries
Bureau of Census, U.S. Covt. Printing Office, 1975 and Economic Report of the President. January 1976, U.S. Covt. Printing
Office 1976.
-------
TABLE 9
SUMMARY OP COSTS FOR LAND DISPOSAL OF POTENTIALLY HAZARDOUS WASTES FROM CONCENTRATORS
(SICa 1021. 1031, 1094, and 1099)
Ore and
Concentrator
Copper
Flotation
Lead-Zinc
Flotation
Uranium
Leach
Uranium
Leach
Antimony
Flotation
Plua
Beryllium
Leach
Location
NA
HA
Wyoming
Nev Mexico
NA
NA
Disposal
mode
Tailings
Pond
Tailings
Pond
Tailings
Pond
Tailings
Pond
Tailings
Pond
Tailings
Pong
Level I
Range of
cost
($/MT)
$.0275
$.0282
$0.55-
$1.63
$0. 18-
$0.30
$0.13-
$0.18
$1.63-
$1.65
$0.54-
$0.58
technology
Total cost
to
Industry!/
($10°/yr)
$6.6-$6.8
$6.8-
$20.3
$0.50-?
$0.84
$0.36-**
$0.51
$0.310-
$0.312
$0.05-
$0.052
Levels II and III techno Ion v
* of value
added in
mining^
.6*
3.27.-
9.5*
0.5*-
0.8*
< HI*
Dlaposal
mode
Tailings
Pond*
Tailings
Pond*
Tailings
Pond*
Range of
cost
($/MT)
$1.55-
$2.22
$0.195-
$0.318
$0.134-
$0.186
Same
Same
Total coat
to 7. of value
industry added In
(|_106/yr) mining^
$19.2- 8.97.-12.9!
$27.6
$ 0.54-?
$ 0.89 0.57*-
$ 0.38-** 0.89*
$ 0.52
as Level 1
aa Level I
Level II technology
Level .III technology
• Ore and
Concentrator
Cooper
notation
Disposal
node
Tailings
Pond??
Range of
cost
($/MT)
$0.32-
$0.33
Total cost
to
Industry**
(S106/yr)
$77.1-
$79.4
7. of value
added In
7.2%
Disposal
node
Tailings.
Pond***
Range of
colt
($/MI)
$0.516-
$0.522
Total cost
to
industry**
(S106/yr)
$124.3-
$125.7
7. of value
added In
mining t
U. 37.-
11.41
NA • Not applicable.
* Derived from.mottle tons of concentrator wastes (dry weight) per year In 1974.
t 1972 value adde
-------
4. The presence of pyrite (FeS-) in the wastes significantly increases the
hazard potential. Pyrite can react in the presence of air and water to form
sulfuric acid, which can leach potentially hazardous metals from the waste.
If pyrite is not present, or if the pyrite is treated with alkaline material
(e.g., lime or soda ash) and the water contacting the waste is kept above pH
7.5, no leaching of potentially hazardous metals will occur. It is, therefore,
important to determine whether the wastes contain free pyrite.
5. There is a general lack of available data as to the potentially
hazardous substances contained in metal mining wastes. Mining companies
usually analyze their land-disposed wastes primarily for the metals being
extracted, not for potentially hazardous substances per se.
6. Based on the limited data presently available, potentially hazardous
substances contained in land-disposed mining wastes which have been identified
are as follows:
* Copper ore wastes (tailings): copper, lead, zinc, pyrite, and cadmium.
* Lead-zinc ore wastes (tailings): lead, zinc, copper, pyrite, and
cadmium.
* Uranium ore wastes (waste rock): uranium and radium.
* Uranium ore wastes (tailings): radium, thorium, and arsenic.
*. Antimony ore wastes (tailings): antimony, lead, and zinc.
* Beryllium ore wastes (tailings): beryllium.
7. Adequate waste treatment and land-disposal methods for potentially
hazardous wastes are available to the companies mining and concentrating
copper, lead-zinc, zinc, mercury, and miscellaneous metals ores.
8. There are wide variations between the various metals mining industries
as to the costs of waste disposal, depending on the type of ore, mining
method, quantity of waste handled, and the processes employed.
All units reported in this report are metric units with English units in
parentheses in the text where possible. A table of conversion units and a
glossary of terms are contained in Appendix D.
Included with this report are four appendices. Appendix A contains a list of
the active mine and concentrator operations in 1974 for SICs 1021, 1031, 1092,
1094, and 1099. Appendix B contains listings of the mines and concentrators
visited, mines and concentrator operators that returned the questionnaire, and
a sample questionnaire. A report on the geology and mineralogy of the various
mining regions is contained in Appendix C. A reference bibliography, the
glossary, and a table of conversion units are contained in Appendix D.
23
-------
REFERENCES
1. U,S, Department of the Interior, Bureau of Mines. Commodity data summaries,
1974; Appendix I to mining and minerals policy.
2, Gruber, G. I. Assessment of industrial hazardous practices, organic chemicals,
pesticides and explosives industries. Environmental Protection Agency,
Report No. 2566-60/0-TU-OO, Jan. 1976.
.24
-------
SECTION I
COPPER ORES (SIC 1021)
Industry Characterization
History of the Industry, Copper has been a contributor to mankind's progress
from the Stone Age to the Space Age. During this span of several thousand
years, uses have ranged from tools, weapons, utensils, statuary, and
architectural applications to the generation and transmission of electrical
energy.
Copper was one of the first metals used by humans, because it was available
naturally in essentially the pure form, and when beaten or forged, it hardened
sufficiently to permit primitive man to fashion utensils and sharp implements
and weapons of lasting quality. Little is known about prehistoric mining of
copper* although it seems certain that both open-pit and underground mining
methods were used.
Evidence of the first mining of copper in North America was discovered by
archeologists in pits in the Upper Peninsula of Michigan and on Isle Royale
in Lake Superior. Copper was first produced in the American Colonies at
Simsbury, Connecticut, in 1709; and a significant ore body was found in
Orange County, Vermont (Ely mine), in 1820. Discovery of the ore deposits
in Michigan in the early 1840's began a new era for copper production in
the United States. Extensive ore bodies were found in Montana and Arizona
during the years 1860 to 1880, and smelter production from domestic ores
increased from 660 MT (762 tons) in 1850 to 27,433 MT (30,175 tons) in
1880. In Bingham Canyon, Utah, the Utah Copper Company (now Utah Copper
Division, Kennecott Copper Corporation) began exploitation of a large
low-grade porphyry deposit in 1906. The success of this operation stimulated
search for and development of other large deposits of similar characteristics
both in the United States and abroad. Development of the froth flotation
process in the first two decades of the 20th century vastly increased the
availability of copper contained in the large low-grade deposits which
constitute most of the U.S. copper reserves.
The United States has been the largest copper-producing nation in the
world since 1883, and in 1973 produced about 22 percent of the total world
copper. Other principal copper-producing nations are the U.S.S.R,, Canada,
Zambia, Chile, Peru, and Zaire.
25
-------
Domestic Production and Capacity. During the period 1963 to 1973, U.S.
copper output (recoverable content of ore*), increased 41 percent from
1,100,000 MT (1,212,542 tons) annually to 1,550,000 MT (1,708,582 tons).
During the same period, world production of copper increased 55 percent
from 4,489,000 MT (4,948,275 tons) to 6,984,000 MT (7,698,542 tons) (Table
10). While the U.S. government statistics are not available for 1974, it
is estimated that the amount of copper concentrate produced in the United
States was nearly 7,324,456 MTt (8,073,830 tons) (Table 11). Table 12 gives
copper ore production by year, and Table 13 gives the estimated copper ore
production by state and region for 1974.
Domestic copper demand is forecast to increase at an annual average growth
rate of 4 percent through 1983. New mine projects and expansion plans are in
approximate balance with the forecast consumption increase and should
maintain the present high degree of self-sufficiency. Based on the projected
rate of demand, copper production in the United States would be approximately
1,822,000 MT (2,008,411 tons) of copper in 1977 and 2,306,000 MT (2,541,929
tons) in 1983.
Copper production is concentrated in comparatively few mines. In 1972, 25
mines accounted for 93 percent of the U.S. output; the five largest provided
41 percent of domestic mine production. In 1974, approximately 47 mining
operations, encompassing 56 mines, were producing copper in the United States.
Mine production of copper in the United States was approximately 88 percent
of the estimated capacity in 1972. Mine production capacity is expected to
have increased by 16 percent between 1972 and 1975, to a total of 1,995,000
MT (2,199,000 tons) (Table 14).
Number, Location, Age, and Size of Active Mines, and Mills. Ninety-eight
percent of the copper production in the United States in 1974 was located
in six states; Arizona, Utah, New Mexico, Montana, Nevada, and Michigan,
and virtually all of the remainder in Tennessee, Maine, Idaho, Missouri,
and Oklahoma (Figure 1). Forty-eight of the 57 copper mines operating in
1974 were located in the first six states (Table 15 and Appendix A).
Copper mines vary in size from small operations employing fewer than 20
workers to mines employing more than 2,500 workers. Employment data from
57 active mines in 1974 indicate that more than one-half of these operations
employed more than 500 workers. More tham 20 percent of the mine operations
employed more than 1,000 workers.
* Recoverable content of ore = the amount of pure copper recovered.
t Recovered copper concentrate = the concentrated copper ore before smelting
and refining. The copper concentrate product will vary from 25 to 90
percent copper depending on the method used in the concentration process,
It is estimated that 95 percent of the product is 25 percent copper
concentrate.
26
-------
TABLE 10
COPPER PRODUCTION, 1963-1983*
(1,000 MT).
Year
United States
Rest of world
Total
1963
1964
1965
1966
1967
1968
1969
1970
1971
1972
1973
1977
1983
1,100
1,131
1,226
1,296
865
1,093
1,401
1,560
1,380
1,510
1,550
l,822t
2,306t
3,389
3,506
3,598
3,679
3,873
4,024
4,244
4,461
4,654
5,124
5,427
—
~ ™
4,489
4,637
4,824
4,975
4,738
5,117
5,645
6,021
6,034
6,634
6,984
--
••••
Source: References 1 and 2.
* Data published in tons and calculated to metric tons.
t Estimated production.
27
-------
TABLE 11
U.S. COPPER CONCENTRATE PRODUCTION, 1974*
(RECOVERED COPPER CONCENTRATE PRODUCT)
State .MT
Arizona 4,876,625
Idaho 17,353
Maine 3,654
Michigan 223,167
Montana 290,299
Nevada 357,498
New Mexico 616,480
Oklahoma 4,445
Tennessee 72,575
Utah 862.359
U.S. total 7,324,456
EPA regions
I
II
III
IV
V
VI
VII
VIII
IX
X
. ... • MT
3,654
--
--
72,575
223,167
620,925
--
1,152,658
5,234,123
™ ™
Source: Reference 3*
* Engineering & Mining
Journal and unpublished
mining company data.
t Data.published in tons
and calculated to metric
tons.
28
-------
TABLE 12
TOTAL COPPER MINE PRODUCTION OF ORE
BY YEAR*
(ORE PRODUCED)
Year 1.000 MT
1967 115,272
1968 154,270
1969 202,984
1970 233,807
1971 220,133
1972 242,065
1973 263,081
Source: References 3 and 4.
* Data published in tons and
calculated to metric tons.
29
-------
TABLE 13
i
U.S. COPPER ORE PRODUCTION FROM MINES BY STATE AND EPA REGION, 1974
(ORE MINED) *
State* : ._ 1.000. MI
Arizona
Idaho
Maine
Michigan
Montana
Nevada
New Mexico
Oklahoma
Tennessee
Utah
tt.S. total
EPA regions, _
I
II
III
-IV
V
VI
vi-i
VIII
IX
X
148,417
364
48
8,059
56,454
12,294
18,829
213
1,996
J35,.552
382,226t
..... , __ .UOOO.MT
48
--
--
1,996
8,059
19,042
'
192,006
160,711
3.64
U*S. total 382,226
•Source: Reference 5. .Engine:e3&ng..{^-Mining ..Jo.urha 1
and utfpublished Mning conipany §a'ta«.
•* States listed are "the only 'ones that have copper mine's,
t Incomplete total-^canhot Estimate quantity of ore in
in situ leach.
30
-------
TABLE 14
PBDJECTED U.S. COPPER CAPACITY, 1972-1975
(1,000 MT)*
1972 1973 1974 1975
Mine and concentrator 1,723 1,814 1,905 1,995
Smelter 1,778 1,814 1,877 1,995
Refinery 2,639 2,639 2,676 2,839
Source: Reference 1.
* Data published in tons and calculated to metric tons.
31
-------
to
Figure 1. Location of copper (SIC 1021) mines for 1974,
-------
u>
TABLE 15
NUMBER, LOCATION, AND SIZE OF ACTIVE COPPER MINING AND CONCENTRATING OPERATIONS* (1974)
State
Arizona
Idaho
Maine
Michigan
Montana
Nevada
New Mexico
Oklahoma
Tennessee
Utah
U.S. total
EPA regions
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
Employees Mining
1-19 20-49 50-99 100-249 250-499 500-999 1,000-2,499 2,500+ operations*
22 4 2 11 6 27
I I
1 1
1 1
1 I
1 11 3
3212 8
1 1
1 1
11 13
362 8 3 15 8 2 47
1 I
1 1
1 I
31212 9
11 24
22 5 2 12 7 30
1 1
No. of
mines
31
1
1
1
2
3
9
1
5
3
57
1
5
1
10
5
34
1
No. of
concentrators
35
1
1
1
2
6
8
I
1
5
61
1
1
I
9
7
41
1
Source: References 4 and 5.
* Location of mining company operations--may or may not include more than one mine at the location designated.
-------
Determination of the age of active mines in the copper industry is difficult.
Secondary source information does not usually give these data, and only
fragmentary statistics are available. Some mines have opened and closed on
many occasions. Some mines started from a multiplicity of small mines and have
grown into one large mine. Where data exist, it is apparent that many active
mines have been operating for a long period of time. For example, the Copper
Queen operation at Bisbee, Arizona, opened in 1885, the Copperhill operation
in Tennessee in 1899, and the Utah Copper operation in 1906, Likewise, thefe
has been some recent copper mine development, such as the Pinto Valley
operation and Oracle operations in Arizona which opened .in 1974, and the
San Xavier, Arizona, operations in 1973, Of the 30 active copper mining
operations for which data are available, six were opened before 1930 and
seven were opened since 1969 (Appendix A),
Employment. Total employment in active copper mines in 1974 i.s estimated at
34,158, About 53 percent of the copper mine workers in-the United States are
employed in Arizona where more than one-half of all copper mines and most of
the large ore-producing mines are located (Table 16).
By-Product/Coproduct Relationships. About 98 percent of the U,S, production
of copper is recovered from ore mines primarily for copper content, with the
remainder being recovered from complex or base metal ores, as in the recovery
of copper from the lead-zinc ores of Missouri, In addition to copper, important
quantities of gold, silver, molybdenum, nickel, selenium, tellurium, arsenic,
rhenium, iron, lead, zinc, sulfur, and platinum-group metals are recovered as
by-products from copper ores.
Lead, molybdenum, iron, and zinc minerals are separated from copper minerals
by selective flotation. Gold, silver, nickel, platinum, palladium, selenium,
and tellurium are recovered from anode sludges at electrolytic copper
refineries. Arsenic and sulfur are extracted .during copper smelting, and
rhenium is obtained in the processing of .molybdenum concentrates recovered
as a coproduct in the treatment of some copper ores. Copper by-products and
coproduct relationships are identified in Table 17.
34
-------
TABLE 16
EMPLOYMENT AT ACTIVE COPPER MINE OPERATIONS
(1974)
State
Employment
Arizona
Idaho
Maine
Michigan
Montana
Nevada
New Mexico
Oklahoma
Tennessee
Utah
U.S. total
EPA regions
I
II
III
IV
V
VI
VII
VIII
IX
X
18,090
200
91
2,444
2,644
3,199
1,820
75
903
4,692
34,158
Employment
91
'
,
903
2,444
1,895
—
7,336
21,289
200
Source: References 5 and 6<
35
-------
TABLE 17
DOMESTIC COPPER BY-PBODUCT AND COPRODUCT RELATIONSHIPS
(1972) (METRIC)*
Source
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
.Copper
Lead
Zinc
Iron
Silver
Gold
Tungsten
Fluorine
By- or
coproduct
Arsenic
Rhenium
Selenium
Palladium
Tellurium
Platinum
Silver
Molybdenum
Gold
Nickel
Sulfur
Zinc
Iron
Lead
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Quantity
(MT)
W
2.772
0.322
W
W
W
0..440
18,375
0.0145
2,272
533,000
6,300
W
900
1,. 485, 000
11,700
7,200
W
1,800
—
'
t-
% of
Total
product
output
100.0
100.0
100.0
100.0
W
39.4
36.0
33.4
15.9
5.8
1.5
.W
0.2
98.4
0.8
0.5
W
0.1
W
W
W
Source: Reference 1.
* Data given in English units and calculated to
metric units.
t Less than 1/2 MT.
W = Withheld to avoid disclosure.
36
-------
Waste Generation and Characterization
Waste Generation. Tables 18 and 19 are compilations of the copper production
and waste generation data for the United States in 1974. A total of 383,133,000
MT (422,331,835 tons) of ore and 7,410,379 MT (8,168,545 tons) of copper
concentrate was produced. There was a total of 626,921,000 MT (691,062,000 tons)
of dry land disposed solid waste generated from the mining and concentrating of
copper ores, 365,787,000 MT (403,211,000 tons) of waste rock, 44,511,000 MT
(49,066,000 tons) of overburden, 216,623,000 MT (238,785,000 tons) of
concentrator tailings, and 531,578,300 MT (585,955,000 tons) of water to
tailings ponds.
The ratios of total waste rock, overburden, and concentrator wastes to ore
mined by state and region are shown in Table 20. In 1974, 0.95 ton of waste
rock was generated and land-disposed for every ton of ore mined. Also, 0.12
ton of overburden was disposed of for every ton of ore. Nationwide, 0.63 ton
of concentrator dry wastes and 2.0 tons of wet waste were generated for each
ton of copper ore mined. A regional variation from 0.01 ton (EPA Region IV)
to 3.26 tons (EPA Region VI) of waste rock for each ton of ore was noted.
Some EPA Regions (I, IV, V, and X) had no overburden because all mines in these
Regions are underground mines. The ratio of overburden to ore varied from
0.0009 ton/ton in EPA Region VIII to 0.63 ton/ton in EPA Region VI. There was
not as great an EPA Regional variation in the ratio of concentrator dry and
wet wastes to ore.
Three factors are involved in the amount of waste per ton of ore: type of
mine, underground or open pit; ore grade; and ore type; oxide, sulfide, or
mixed oxide-sulfide. An underground mine will have less waste rock and no
overburden to remove and dispose on land. In general, the ore grade is higher
in underground mines, hence less concentrator waste. The ore type will determine
the concentrating method used. Concentration by flotation generates more solid
waste than does concentration by leaching and precipitation.
Table 21 shows the potentially hazardous wastes disposed on land for 1974.
Table 22 contains the projected data for 1977 and Table 23 the projected total
potentially hazardous waste for 1983. The three tables show the potentially
hazardous material destined for land disposal by state, EPA Region, and
nationwide for the copper concentrators. Table 21 indicates that in 1974
there were 461 million metric tons (507 million tons) of potentially
hazardous wet waste disposed of on land and 140 million metric tons (154
million tons) of potentially hazardous dry waste. The total process waste
was 1,423 million metric tons (1,566 million tons), of this total, 772
million metric tons (849 million tons) or 54 percent of the total waste
was concentrator waste and disposed of on land. The potentially hazardous
material was 65 percent of the total material disposed in tailings ponds.
In 1977 there will be 521 million metric tons (573 million tons) of
potentially hazardous materials for land disposal, representing an increase
of 13 percent over 1974.
37
-------
TABLE 18
TOTAL PRODUCTION STATISTICS BY STATB AND EPA REGION FOR SIC 1021 COPPER ORES FOR 1974 (METRIC)
Concentrator wastes
State
Maine
Tennessee
Michigan
New Mexico
Oklahoma
Total
Montana
Utah
Total
Nevada
OJ Arizona
00 Total
Idaho
National
Ore. mined
Region 103 TPV
I 48
IV 1,996
V 8,059
19,899
213
VI 20,112
50,454
135.552
VIII 192,006
12,294
148.417*
IX 160,711
X 201
383,133*
Waste rock
103 TPY
1
23
7,836
53,870
8.165-
62,035
9,072
14.515
23,587
36,287
235.868
272,155
20
365,657
Overburden
103 TPY
0
0
0
11,793
181
11,974
0
161
181
1,963
30.393
32,356
0
44,511
Dry
weight
103 TPY
40
714t
7,292t
19,233t
209
19,442
56,088
35.017
91,105
12 , 156
109.904t
122,060
141t
240,794
Wet
weight
103 TPY
160
7.483
34,618
62,933
836
63,769
196,308
111.776
308 , 084
37,156
320.274
357,430
828
772,372
103 TPY
3.654
72,575
246,163
616,480
4.445
620,925
365,596
862.359
1,227,955
357,498
4;876,625
5,234,123
5,024
7,410,379
Metal cone.
103 TPY
3,322
834,610
0
0
0
0
91
10.655
10,746
2
55.645
55,647
0
904,325
Miscellaneous
TPY
0
0
0
0
0
0
98
226.796
226,894
0
0
0
0
226,894
t
Total TPY
6,976
907,185
246,163
616,480
4.445
620.925
365,785
1.099.810
1,465,595
357,500
4.932,270
5,289,770
5,024
8,641,638
Ratio
of
dry
weight
waste
to ore
0.85
0.37
1.88
4.3
40.2
4.64
1.15
0.37
0.6
4.1
2.6
2.7
0.8
1.7
Ratio
of 7.
dry Total
weight dry
waste waste
to In
product region
5.9 0.01
0.8 0.11
61.0 2.32
138
1,924
150 14.36
178
45
78 17.65
140
80
80 65.53
32 0.02
75
1;
Total
wet
waste
in
reeion
0.02
0.97
i.48
_
8.26
_
_
39.89
-
46.28
0.11
Source: References 2, 5, and 15.
* Incomplete total cannot estimate ore for In situ mining.
t Part of tailings are classified and loose sand returned to underground mine for backfilling.
-------
TABU 19
CONCENTRATOR WASTES. ORE MINED AND COPPER PRODUCED FOR SIC 1021
(METRIC)
Number Ore
Process of mined
code* Key** mines 10 J TPY
A A 23 149,313
B B 12 65.807
R C -t- H 7 104,155
S u + 11 1 9,181
CO
vO L D + I 3 5,895
H B + K 3 7,519
N E + H 11 . 16,134
0 C + II 2 25,129
P F + H 5 -
0 F + K 1
Total 68* 383.1338
Percent Coppery
total produced
ore TPY
38.97 1.070.224
17.18 526,458
27.19 824,459
2.40 64,265
1.54 47,159
1.96 37,594
4.21 80,671
6.56 75,387
10.786
4.082
2,741,085
Concen- Percent Ratio of Ratio of Potentially Percent of
Percent trator of waste to waste to hazardous total waste
total waste total copper ore waste potentially
copper 101 TPY waste produced mined 10} TPY hazardous
39.
19.
30.
2.
I.
1.
2.
2.
0.
0.
04 140,179 64.71 131 0.94 140,179 64.71
21 61,494 28.39 117 0.93
08 - - -
34 9,101 4.20 142 0.99
72 5,866 2.71 124 0.99
37 - - - -
94 ... -
75 -
39 . . - - -
15 - - - - -
216,640 79 0.57 140,179 64.71
Source: References 2, 3, and 15.
* Letter denotes process described In Table 24.
f Copper produced Is expressed as tons of pure (1007.) copper.
4 Duplication; some mines use more than one process.
§ Incomplete total; cannot estimate ore quantity In In situ leach.
** See Table 24.
-------
TABLE 20
RATIO OF TOTAL WASTE KOCK-OVERUURDEN-CONCENTRATOR DRY AND WET WASTES TO ORE MINED
FOR SIC 1021 COPPER ORES (METRIC)*
Stace Region
Maine I
Tennessee IV
Michigan V
•P-
O New Mexico
Oklahoma
Total VI
Montana
Utah
Total VIII
Nevada
Arizona
Total IX
Ore mined
103 TPY
48
1,996
8.059
19,899
213
20,112
56,454
135.552
192,006
12,294
148.417
160,711
Waste
rock
103 TPY
1
23
7 .836
53.870
8 . 165
62,035
'9,072
14.515
23.587
36,287
235.868
272,155
Ratio of
total
waste
rock
to ore
0.02
0.01
0.97
2.7
38.3
3.26
0.16
0.11
0.12
2.95
1 .59
1.69
Over-
burden
103 TPY
0
6
0
11,793
181
11,974
0
18J
181
1,963
30.393
32,356
Ratio of
total
over-
burden
'to ore
0
0
0
0
0
0
0
o
0
0
0
0
.6
.85
.63
.001
.0009
.16
.20
.20
Dry
concen-
trator
waste
10 TPY
40
714
7,292
;19,233
209
19,442
56,088
35.017
91.105
12,156
109.904
122,060
Ratio of
dry concen- Wet
trator concentrator
waste
to ore
0.8
0.36
0.91
0.97
0.98
0.97
0.99
0.26
0.47
0.99
"6.74
0.76
waste
103 TPY
7
34
62
63
196
111
308
37
320
357
160
,483
,618
,933
836
,769
,308
.776
,084
,156
.274
,430
Ratio
of wet
waste
to
ore
3.3
3.7
4.3
3.2
3.9
3.2
3.5
6.8
1.6
3.0
2.2
2.2
201
20
0.1
141
0.70
828
4.1
National
383.133
365.657
0.95
44,511
0.12
240,794
6.63
772,372
2.0
* Table compiled from data in Table 18, p. 38.
-------
TABLE 21
TOTAL AMD POTENTIALLY UAZARDOUS WASTE FttCM MINING AND CONCENTRATING COPPER ORES FOR SIC 1021 FOR 1974 (METRIC)*
Ury waste welRht ( 103 TPY)
State Region
Maine I
Tennessee IV
Michigan V
New Mexico
Oklahoma
Tota I VI
Montana
Utah
Total VIII
Nevada
Arizona
Tota I IX
Idaho X
National
Total
process
waste
41
737
IS. 128
84.896
8.555
93.451
65.160
49.713
114.873
50,406
376.165
426,571
161
650,962
Total
potentially
hazardous
waste
40
737
7.292
12,500
209
12.709
36,456
22.760
59,216
7,900
52.167
60,067
141
140,179
Haste
rock
Overburden
Potentially
Total hazardous Total
1
23
7,836
53,870
8.165
62.035
9,072
14.515
23,587
36,287
235.868
272.155
20
365.657
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
11,793
181
11,974
0
181
181
1.963
30.393
32.356
0
44.511
Concentrator tailing*
Potentially
hazardous Total
0
0
0
0
0
0
0
0
0
0
Q
0
0
0
40
714
7.292
19.233
209
19,442
56,0o8
35.017
91,105
12.156
109.904
122.060
141
240.794
Potentially
hazardous
40
714
7.292
12.500
209
12,709
36.456
22 . 760
59,216
7.900
52.167
60,067
141
140.179
Wet weigh
t (I03 TPY)
Concentrator tailings
Potentially
Total hazardous
160
7.483
34.618
62.933
836
63,769
196,308
111.776
308,084
37,156
320.274
357,430
828
772,372
160
7,483
34,618
40,902
836
41.738
127.596
72.651
200,247
24.147
152.021
176.168
828
461,242
Source: Reference 15.
* MK1 engineering judgment.
-------
TABLE 22
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING OF COPPER ORES FOR SIC 1021 FOR 1977 (METRIC)*
Dry waste weights (103 TPY)
State
Mnlne
Michigan
New Mexico
Oklahoma
Total
Montana
Utah
Total
Nevada
Arizona
Total
Idaho
National
Total
process
Region waste
I 46
IV 833
V 17,095
95,932
9,667
VI 105,599
73,631
56,176
VIII 129,807
56,959
425,066
IX 482,025
X 182
735,587
Total
potentially
hazardous
waste
45
8,240
14,125
236
14,361
41,195
25,719
66,914
8.927
58,949
67,876
159
158.402
Waste
Total
1
8,855
60,873
9,226
70,099
10.251
16,402
26.653
41,000
266,531
307,531
23
413,188
rock
Overburden
Potentially
hazardous Total
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Q
0
13,326
205
13,531
0
205
205
2.218
34,344
36,562
0
50,298
Concentrator tailings
Potentially
hazardous Total
0
0
0
0
0
0
0
0
0
0
0
0
0
45
8,240
21.733
236
21,969
63,379
39,569
102,948
13,736
124,192
137,928
159
272,096
Potentially
hazardous
45
8.240
14,125
236
14,361
41,195
25.719
66,914
8.927
58,949
67.876
159
158,402
Wet wetghi
t (103 TPY)
Concentrator tailings
Total
181
39.118
71,114
945
72,059
221,828
126,307
348,135
41,986
361,910
403,896
936
872,781
Potentially
hazardous
181
39,118
46,219
945
47,164
144.183
82.096
226,279
27.286
171,784
199,070
936
521.204
Source: Reference 2.
* MRI engineering judgment.
-------
TABLE 23
PROJECTED TOTAL AMD POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING
COPPEB ORES FOB SIC 1021 FOB 1983 (METRIC)*
State
Maine
Tennessee
Michigan
-P- New Mexico
Oklahoma
Total
Montana
Utali
Total
Nevada
Arizona
Total
Idaho
national
Total
process
Region wastes
I 60
IV 1,094
V 22.390
125,650
12.660
VI 138,310
96.437
73,575
VIII 170,012
74,600
556.72»
IX 631,324
X 238
963.428
Total
potentially
hazardous
waste
59
1.060
10.792
18,500
308
18.808
53,955
33.685
87,640
11,692
77.207
88.899
209
207.467
Concentrator tailings weight (10
Waste rock
Total
1
34
U.59b
79.730
12.084
91.814
13.427
21.482
34.909
53.705
349.085
402.790
30
541.176
Potentially
hazardous
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Overburden
Total
O
0
0
17.455
268
17.723
0
268
268
2,905
44.980
47.885
0
65.876
Potentially
hazardous
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Dry weight
Total
59
1.060
10.792
28,465
308
28,773
83,010
51.825
134.835
17,990
162.658
180.648
209
356,376
Potentially
hazardous
59
1.060
10.792
18,500
308
18,808
53,955
33.685
87.640
11.692
77.207
88.899
209
207.467
Wet
Total
237
11.075
51,235
93 . 140
1.237
94.377
290.536
165.428
455.964
54.990
474.000
528,990
1.225
1.143.103
3 TPY)
weight
Potentially
hazardous
237
11.075
51,235
60.535
1.237
61.772
188.842
107,524
296.366
J5.738
224.991
260.729
1.225
682.639
Source: Reference 2.
* MRI engineering judgment.
-------
In 1983, there will be 683 million metric tons (751 million tons) of
potentially hazardous materials for land disposal. This amount of potentially
hazardous material represents an increase of 30 percent over 1977 and 48
percent over 1974.
Mining Processes. The copper mining industry uses a variety of mining methods.
There are three general types of copper mines: open pit, underground, and
in situ.-'
Open-Pit Mining. There are several different mining methods used in open-pit
mining of copper ores. The most common method is the bench method, in which
drilling and blasting are carried out on one level or bench, and the drilling
equipment moved to a lower bench for drilling and blasting. Af.ter the ore is
loosened, it is loaded by shovels into trucks or railcars for delivery to the
concentrator. At least one mine uses rippers to remove the ore from the ground
and loads it into scrapers for delivery to the concentrator. Some of the smaller
open-pit mines drill and blast the ore loose from the surroundings. They do not
use benches but take everything to the same level.
Typically, open-pit mining involves four basic steps: drilling, blasting,
loading, and hauling. The conventional mining operations include drilling,
blasting, power shovel loading, and truck and rail haulage. The flow diagram
for a typical open-pit mining operation is shown in Figure 2. About 55 percent
of the mines include crushing and grinding at the mine. Ten percent do not
crush or grind. The other 35 percent ship directly to the concentrator where
crushing and grinding are carried out.
Trucks are utilized for transport at the mine site (e.g., to a primary
crushing plant and concentrator), and rail haulage is used to move the ore to
off-site concentrator plants. There are 40 open-pit copper mines, and they
account for 85 percent of the copper ore mined.
Mass balance data for the typical mine are given in Figure 2.
Description of Individual Waste Streams. All waste rock contains less than
0«05 percent copper and is hauled to a waste rock dump. The overburden,
685,000 MT/year (754,870 tons/year), is loaded on trucks and hauled to a waste
dump 0
Identification of Potentially Hazardous Waste, None of the waste rock
disposed on land is considered to be potentially hazardous.
Underground Mining. An underground mining method used in an Arizona copper
mine is described herein as a typical example of such mining practice in the
United States (Figure 3). There are 12 underground copper mines accounting for
14 percent of the copper ore mined.
44
-------
Strip
Overburden
1
Drill Ore
Body
J
Blast- to
Loosen
I
1 f*nt\ <"*»••«»
Load v_;re
[3,356,
Crushing
Primary &
Secondary
1
Concentrator
•» Load •» n
m Overburden 685.000MTPY Dv""K
" Waste Rock 7,586,000 MTPY^ r
585 MTPY
Source: Reference 15.
Figure 2. Typical open-pit copper mine.
45
-------
Drill
Ore Body
I
Blast
Ore Body
Load
Waste Rock
,226,796"MTPY
Dam Construction
Tailings Pond
Load Ore
56,791 MTPY
834,610 MTPY
Concentrator
**• Sand Plant-
249,746 MTPY
Backfill
Stopes in Mine
Source: Reference 15.
Figure 3. Typical underground copper mine,
46
-------
The ore at this mine occurs in a mesothermal deposit, and although the bulk
of the ore has come from a vein, some of the recent production has come from
replacement deposits in limestone^' The ore bodies are "typical" of mesothermal
copper deposits. .
The principal minerals in the vein are chalcopyrite, bornite, enargite,
tennantite, chalcocite, and digenite. The chief gangue minerals are pyrite,
quartz, limestone, and hematite*
Description of an Underground Mining Operation, Mining is conducted using
stope and pillar methods with drilling and blasting. There are several mining
levels and exhaust ventilation shafts. The lowest operating level is about
1,130 m under the surface. The broken ore is raised by a skip hoist and
transported by rail to a nearby concentrator. Waste rock from the mine has
been used for dam construction (tailings pond). Figure 4, a flow diagram of
an underground copper mining operation, indicates that part of the ore from
the mine is shipped directly to a smelter.
Description of Waste Streams. The only waste material in the mining
operation is waste rock.
The waste rock consists principally of granite, schist, and limestone
containing quartz and hematite, and consists mainly of large lumps of material
along with some coarse rock particles.
Identification of Potentially Hazardous Waste Streams. Waste rock, which
is mainly granite and schist from underground mining operations, is not
considered to be a potentially hazardous material, because it is not subject
to wind transport or water leaching of any environmental pollutants.
In Situ Mining. There are three methods used for in situ mining. One method
is to remove overburden and stockpile it; the ore body is then drilled and
blasted in place. A second method is to drill and blast the ore body on the
side of a hill and use the blast to move the ore into an adjacent valley. The
third method is to drill and blast underground to shatter the ore body. Then
the ore is leached by sulfuric acid which is pumped to a concentrating plant
for recovery of copper. One mine has used the first two methods, and the other
two mines use the third method.
Concentrator Processes. Copper concentrating is accomplished using a primary
process or a primary process in combination with secondary processes. Table 24
shows the combinations that are used in the concentrators for copper recovery.
The choice of concentrating method is dictated by the ore type, size of ore
body, and concentration of copper in the ore. It is completely independent of
the mining method used to recover ore from the earth.
47
-------
OPE FROM MlfJ£
SHIPPING- GKADE ORE
TO SMELTER
CRUSHING, GRINDING
& CLASSIFICATION
WASTE
TRAMP
IRON
TO
ADDITIVES'
ROUGHER FLOTATION
CONC.
UNDERFLOW,
CYCLONE
UNDER-
FLOW
CLEANER PLOTATION
THICKENING- A FILTERING-
ADDITIVES'
FLOTATION
NAME
Lino
Aeroflool 404
Iwpropyl Xonfhote
Dow Frolher 250
Polyocrilimida ( Flocculont )
ADDITIVES'
LBAON
OF ORE
2.5
0.003
0.10
0.05
0.01
G/MT
OF ORE
1250.0
1.5
50 0
:s o
50
.
RECLAIM WATER RECYCLED
TO FLOTATION
OVERFLOW _
TAILING THICKENER
UNDER-
FLOW
PYRITE FLOTATION
CONC.
COPPER CONCENTRATE
STORAGE
©
TAIIS
PVRITE STORAGE
SAND PLANT
FINES
(0)
SAND TO MINE AS FILL ©
WASTE ROCK
FOR DAM
CONSTRUCTION
TAILING POND
0
©
Co CONC.
TO SMELTER
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEMS
Quantity. TRY
Metric TPY
_„_, ^y... -
WASTE ROCK
TO DAM
CONST.
250.000
226.794
SMELTER
GRADE
ORE
62.800
54.971
Source: Reference 15.
. - vy •;
ORE
TO
MILL
920.000
834.610
Cu
CONC.
105.000
95.254
PYRITE
CONC.
1,700
1.542
DRY TAILS
TO POND
488.000
442.706
DRY TAILS
TO SAND
FILL
325,300
295,109
WATER
TO POND
4.930.000
4.470.000
Figure 4. Concentrating of copper.
48
-------
TABLE 24
CONCENTRATING PROCESSES FOR COPPER ORES
Ore type
Sulfide-pyrite
Sulfide-pyrite
Sulfide-pyrite
Oxide
Oxide
Oxide
Oxide
Both
Both
Both
Oxide
Oxide
Oxide
Oxide
Oxide
Oxide
Oxide
Sulfide-pyrite
Oxide
Process
code
A
B
C
D
E
F
G
H
I
J
K
L
M
N
0
P
Q
R
S
Process
Flotation
Flotation
Dump leach
Vat leach
Heap leach
In situ leach
Dump leach
Iron precipitation
Electrowinning
Solvent extraction
I + J
D + I
E + K
E + H
G + H
F + H
F + K
C + H
D + H
Waste
material
Yes
Yes
No
Yes
No
No
No
Yes
No
No
No
No
No
Yes
Yes
Yes
No
No
Yes
Potentially
hazardous waste
Yes
Yes
No
No
No
No
No
No
No
No
No
No
No
No
No
No
No
No
No
Source: References 5 and 15.
There are two primary concentrator processes'for recovery of copper from
ore: flotation and leaching. Leaching is accomplished in four different modes:
heap leach, dump leach, vat leach, and in situ leach. In addition to the above
leaching processes, there are three secondary copper recovery processes: iron
precipitation, solvent extraction, and electrowinning. The copper-rich leach
solution must be further processed at the mine. Two methods are used to
precipitate the copper from the leach solution. Precipitation over iron, or
cementation, is the most widely used process, and the product (80 percent
copper) is shipped to a smelter. The use of electrowinning to recover a 99
percent copper product is growing, however, and this material is then sold
to a refinery or copper product manufacturer. The flotation product is shipped
to a smelter. Smelting is the secondary recovery process for flotation
concentrate. (Since smelting is not a mining or concentrating process, it is
not discussed here.)
49
-------
Primary Processes.
Flotation. The wastes from the copper flotation process are considered
tailings, which is a fine sand-like material. Practically all copper flotation
plants dispose of their tailings in a tailings pond.
An analysis of the tailings waste from a copper concentrator operation in
the Coeur d'Alene Region in Idaho is shown in Table 25.-' These data show that
the concentration of cadmium, copper, lead, and zinc (all of which are
potentially hazardous metals) in the tailings stream discharged to a tailings
pond are well above the background concentrations of these metals. The
background data were compiled by analyzing soil samples taken in the area
around the mine, concentrator, and tailings pond.
TABLE 25
ANALYSIS OF TAILINGS FROM A COPPER CONCENTRATOR
Concentration, ppm
Element
Calcium
Cadmium
Copper
Iron
Potassium
Magnesium
Manganese
Sodium
Lead
Antimony
Zinc
Concentrator
1,172
1.4
2,179
264,667 .
115
6,051
19,129
75
1,349
462
868
Background
liSOO
-
21
11,800
1,800
3,700
490
151
51
-
150
Source: Reference 9.
The potentially hazardous waste from the flotation of copper sulfide ores
is the tailings material,, The potentially hazardous wastes in the tailings
shown in Table 26 include cadmium, copper, lead, and zinc0
i
The hazardous substances contained in tailings from copper ore flotation
operations can cause environmental pollution problems if the waste is not
retained at on-site waste disposal impoundments.
50
-------
TABLE 26
ANALYTICAL DATA FOR TAILINGS SOLIDS FOR SIC 1021, COPPER
Concentration (ppo unless Indicated)
Copper
Molybdenum
Sulfur
Iron
Cold (oz/ton)
Silver (oz/ton)
Rhenium (oz/ton)
Aluminum
Arsenic
Cadmium
Lead
Magnesium
Phosphorus
Potassium
Silicon
Sodium
Titanium
Zinc
Zirconium
Cyanides
Mercury
Selenium
Chromium
750 700 2,600 930 1.900 2,000 500 1,000 2,100 1,500 1.300 1,850 1.300 1,800 2,500
0.47 -
8.000
29.000
0.005
0.02
0.004
7,000
Hll 5
5.OOO 60
200 " 7 100
12.000
10,000
50,000
32,000
20,000
10,000
2,300 30 30 1,500 2,000
5,000
Nil
0.01
< I
290
Source: References 9 and 15.
* Data are from 15 mines and represent 32 percent of the mines.
-------
Description of a Flotation Process. The products of copper ore flotation
concentrators are a copper concentrate and by-product concentrates. Crude ore
is crushed in primary and secondary crushers to prepare a feed material for
ball mill grinding. A magnetic head pulley is used on a crushed ore belt
conveyor to separate pieces of scrap iron from the ore. Fine ore from the
crushing operation is wet ground in a ball mill operation in closed circuit
with a cyclone classifier to produce a grind of 88 percent -100 mesh ore for
use in flotation.
The overflow from the cyclone classifier is sent to a bank of rougher
flotation cells. Froth from these rougher cells, containing most of the copper
present in the ore, is processed through cleaner type flotation cells,
thickening, and filtration steps to yield a copper concentrate (25 percent
copper) which is shipped to a smelter.
Underflow from the rougher cells is sent to a cyclone classifier, and
cyclone overflow containing fine sands is sent to tailings thickeners which
separate some reclaim water from the fines for reuse in the process (thickener
underflow). The cyclone underflow is sent to by-product flotation cells.
Concentrate from these cells is by-product metal concentrate. Depending on the
ore produced, the by-product concentrates could be molybdenum, pyrite, or zinc.
Tailings from these flotation cells are sent to a sand plant. The tailings are
classified into two different size fractions in the sand plant. The coarse
fraction is used to build tailings dams or to backfill underground mines. The
fine fraction is discharged to a tailings pond.
A description of the types and quantities of flotation additives used
in the process is given in the flow diagram in Figure 4.
The annual quantities of ore consumed, products, and wastes for this
example process are shown in Figure 4. The overall recovery of copper from
ore is about 95 percent. The consumption of 834,610 MT/year (920,000 tons/
year) of ore corresponds to the production of 95,254 MT/year (104,999 tons/
year) of copper concentrate, 1,542 MT/year (1,699 tons/year) of pyrite
concentrate, and a total of 738,000 MT/year (813,000 tons/year) of tailings.
Description of Individual Waste Streams. Small quantities of scrap iron
waste are recovered in the flotation operation. The scrap metal could be broken
tools or pieces of equipment. This metal is sold as scrap iron or disposed of
as a waste material on a waste dump.
Tailings are the only significant waste material formed in the
concentrating operations. The tailings in the example described above contain
about 0.2 percent copper and 0.15 percent zinc, plus nearly all the gangue
material (e.g., quartz and hematite) in the feed ore used in the process. If
the crude ore contains arsenic minerals, it is probable that the tailings also
contain traces of arsenic; however, we could find no data to confirm the
presence of arsenic in tailings.
o 52
-------
Leaching. The commercially used leaching methods are heap leaching, vat
leaching, in situ leaching, and dump leaching. There are 33 concentrators
leaching copper ore, and they produce about 42 percent (1.1 million metric
tons (1.21 million tons)) of the copper.
Heap Leaching. An open-pit ore deposit in Arizona is used as an example
of this type of operation. This ore body consists of precipitated oxide copper
minerals in host rocks of schist and granite J^'The principal minerals are
copper oxides—chrysocolla, azurite, and malachite. Only trace amounts of
sulfide minerals have been identified in this ore deposit.
Ore is hauled to the leach heap site and dumped in a long thin layer
along an advancing crest. The heap ore is usually started from a beginning
roadway along one side of a canyon. After the ore is spread, a road grader
windrows the ore over the edge to create a gradually advancing crest as more
ore is hauled to the heap. By continually dumping the ore along this crest and
blading it over and down the slope, a 6.1-m (20-ft) layer or lift of ore is
carried across the canyon or across the top of a previously leached lift. Each
heap leach is built up in 6.1-ra (20-ft) lifts, and a complete heap has 10 lifts
for a total height of 61 m (200 ft). However, leaching is started before a heap
is completed.
When a lift is completed, the top 0.3 m (1 ft) is densely compacted
from the travel of the equipment across the surface. To remove this compacted
layer, a ripper breaks it up. Then a scraper is brought in, and the top 0.3
to 0,45 m (1-1.5 ft) is picked up and dumped over the nearest crest. This
operation also helps to level the heap to a common elevation* Then the heap
is ripped about 0,9 m (3 ft) deep to provide a porous top layer.
i '
Under the original heaps, the canyon bottoms and side walls were sealed
with a 10-cm (4-in.) layer of soil cement to prevent loss of copper-rich
solution to the groundwater. This method of providing water-tight basins was
generally satisfactory, but as the weight of millions of tons of ore accumulated
in the heaps, it was noticed that some cracking and pulling of the soil barriers
were occurring, and their integrity became doubtful.
Laboratory tests of the "topsoil" of decomposed granite from this site
showed that if the soil was compacted at 10 to 11 percent moisture to 98 to 100
percent Proctor density, it became as impermeable as an adobe brick when
contacted with dilute acid liquors. For all the new heap areas built from 1968
to date, the canyon bottoms and walls have been compacted very thoroughly, with
water added to achieve the proper moisture content, to prepare them for solution
collection, and to prevent loss of copper-bearing acid solution to the
groundwater.
53
-------
In the canyon bottoms and part way up the walls, collecting berms and
coffer dams are built to collect the solutions flowing from the leach heaps.
Perforated pipes covered with washed gravel are used to form a drain to collect
the pregnant solutions in each collection area. The pregnant solution is drained
to a pond for storage and to control feed to the precipitation plant.
At present there are 20 to 22 leach dumps at this site, but not all dumps
are complete, A complete heap, 61 m (200 ft), is leached for 4 months, rested
for 4 months, and leached again. Mass balance data are shown in Figure 5.
Until a heap is abandoned, there is no waste from the leaching operation.
After the leaching of the ore is terminated, the spent rock residue represents
a potentially hazardous waste material which could cause environmental pollution
through continued natural leaching by rainfall and runoff unless the residual
acid is washed out by water flooding of the heap. Some states require the filing
of an action plan that contains the provision that heaps must be water flooded
before abandonment.
At the present time, four concentrators are using heap leaching.
Dump Leaching,— A typical dump leaching and copper precipitation
operation is described herein as an example of this method for recovering
cement copper from the ore. The flow diagram is shown in Figure 6,
The copper ore recovered from open-pit mining and used for dump leaching
in the example case is a highly altered quartz-monzonite porphyry. The
principal copper mineral is chalcocite, which is deposited in many areas of
the mine as a thin coating on pyrite crystals. The copper content of the ore
ranges up to about 1 percent.
Ore from the open-pit mine is transported by dump trucks to a dump. The
ore may or may not be crushed. The criteria for crushing are the ore size and
contact area available for leaching with the sulfuric acid solution,, About 34
percent (129,284,000 MT/year (142,220,000 tons/year)) of the copper ore is
processed by this method, and 33 percent (900,000 MT/year (990,000 tons/year))
of the copper metal is recovered.
Leaching at this example dump was started in 1956, and has proceeded
simultaneously with dump construction over the years up to 1974. Former leach
areas and haulage surfaces have been ripped with D-9 tractors before deposition
of succeeding layers of ore. The dump surface is prepared for leaching by
ripping, followed by construction of a grid of 3- by 3-m (10- by 10-ft) ponds
on the surface of the leach dump and suitable access roadways.
Make-up water from underground wells combined with barren solution from
copper precipitation is pumped to the dumps through an epoxy-lined pipeline
and distributed through 10- to 20-cm diameter polypropylene lines. Distribution
of the water to the surface of the dump is controlled by use of pinch valves on
the polypropylene pipe.
54
-------
OKE
0
«r
WASTE 80CX
DUMP
TEACHING-
SOLUTION
HEAP LEKHIN&
rSffEO SYSTEM Of
Off LlfTS- EACH 20'H/G-H.
A COMPLETE
/S200'
LEACH
COLLECTING- PONO
EXTRACTION
ORGANIC
SPENT
SOLUTION
Cu
LOOPED
SOLUTION
STRIPPED
OG&4/VIC
STRIPPING-
PREGNANT
SPENT
ELECTROLYTE
STAKTER
CATHODES
£LECTROLVSIS Of COPPER
SULFATE SOLUTION
COPPER CATHOOES-
pRoeucr
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEMS
Quantity, TRY
Metric TRY
Auoy, %Cg
CRUDE
ORE
4,200,000
3,810,177
0.47
^
WASTE
ROCK
5,400,000
4,898,794
0.07
l^
COPPER
CATHODE
PRODUCT
36.925
33,498
99.9+
TYPICAL METALLURGICAL DATA
FLOW ASSAYS G/L
Leach Solution
S-X Plant Feed
Raff i note
Loaded Organic
Stripped Organic
Pregnant Electrolyte
Spent Electrolyte
LITERS/
MIN.
4,939
3,671
—
10.503
2.612
"-••-'
GPM
1305
970
—
2775
—
690
"™ '
Cu. ACID
0.75
2.90
0.30
1.01
.09
36.20
'32.50
6.9
3.8
7.8
—
—
145.0
150.7
Fe
1.0
1.0
—
—
—
Fe"1
0.9
0.9
—
—
—
1.5 i 1.3
•™ —
"^ ~
Current Efficiency 84.25
Pounds of Acid Used , 150,650
Pounds of Acid per Pound Net Elect. Copper 5.10
Source: Reference 15.
Figure 5. Concentrating of copper ore, heap leaching and electrowinning.
55
-------
OPEN PIT MINE
0
MASTS
CKUSHIN& PLANT
B4RPEN
SOLUTION
WATER
LEACH PUMPS
LEACH LIQUOR
COLLECTING POND
SCXAP/PON
Pf?£CIP(T4TORS
CEMENT
COPPER
LIQUOR
(2)
BARREN SOL w
TO LEACH PUMPS
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEMS
Quantity, TPY.
Metric TPY
Assay, % Cu
gpl Cu
PH
\\j
CRUDE
ORE
15,000,000
13,607,800
0.1 - 0.6
ve^ ,
PREGNANT
. LIQUOR
3- 5
2.5- 3.5
i V^i/
CEMENT
COPPER
13,911
12,620
80*
r — \£J
BARREN
SOL'N
0.1 - 0.5
4- 4.2
^
SCRAP
IRON
27,820
25,240
Source: Referenre 15.
*80% Cu on a drug basis. Product normally contains 15- 20% moisture.
Figure 6. Concentrating of copper ore by dump leaching and precipitation.
56
-------
Water is added to the surface of the leach dumps for 2 to 3 months and
then discontinued for a rest period of approximately 1 year. This operation is
then repeated. The period of time used in applying leach water is determined
by productivity, with longer periods being used in areas that tend to produce
higher grade leach liquor. During the water addition period, the water is
pumped continuously to the leach dump at a rate of about 7,570 liters/min
(2,000 gpm). The pyrite contained in the leach dump reacts with percolating
leach water and oxygen from air to generate sulfuric acid and ferric sulfate.
The resulting solution, thus formed, then dissolves copper and other
constituents from the dump material. The dump is self-sustaining in regard
to sulfuric acid production, and no acid is added. If oxide ore is processed,
sulfuric acid is added to the solution before it is distributed on the leach
dump. The copper oxide ore in leach dumps amounts to 25,129,000 MT/year
(27,700,000 tons/year) or about 20 percent of the total copper ore processed
by leaching.
The effluent liquor (pregnant liquor) which drains from the base of the
leach dump is collected in ponds constructed at points of natural drainage egress.
This pregnant liquor from the ponds is pumped to a copper precipitation plant.
Some problems have been encountered in this example leach dump, all of
which are probably typical of most dump leaching operations. These problems
are: (1) a gradually declining copper content in the effluent liquor, despite
continuous addition of new ore to the leach dump; (2) difficulty in obtaining
water penetration of some areas of the dump as indicated by a tendency for the
water to migrate laterally in the upper layers of the dump; (3) the unknown
location and effect of deposits of iron salts in the dump.
The available production data given in Figure 6 show the mass balance of
materials.
The dump leach pile is not considered to be waste material until the
leaching operation is terminated. When leaching is stopped, proper treatment
and disposal practices must be carried out to prevent environmental pollution
from the inactive dump. The State of Arizona requires the filing of a plan of
action for the company to follow before closing the leach dump.
Vat Leaching. A flow sheet for a typical vat leaching and electrowinning
operation is given in Figure 7.
Crushed oxide ore is leached in a series of open-top concrete vats with
an aqueous solution of sulfuric acid (20-25 g/liter H2SO^). The leach solution
is used at ambient temperature, and the pH is maintained at 1 to 1.5. The ore
is leached for 14 days in each vat. Then the slimes are separated and dewatered
by a thickening operation and pumped to a canyon disposal dump which is also
used for disposal of waste rock containing limestone, less than 0.04 percent
copper, and no pyrite. The wasrvj reck is not considered to be potentially
hazardous. The copper-rich leach solution is pumped to an electrolytic copper
recovery plant,
57
-------
OPEN PIT MINE
ORE
WASTE
CRUSHING PLANT
(PRIMARY & SECONDARY)
BARREN
SOLUTION
VAT LEACHING- SYSTEM
©1
Cu CATHODE
f/copt/cr
PREG-NANT
LIQUOR.
SUMES
TO VAT
LEACHIHG-
BARREN
SOLUTION
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEMS
Quantity , TRY
Metric TRY
Assay, % Cu
gpl Cu
gpl H2S04
\\j
CRUDE
ORE
3,700.000
3,356,585
0.9
^j —
PREGNANT
LIQUOR
30
W
LEACH
LIQUOR
5
20
V3/
COPPER
CATHODE
PRODUCT
25,550
23.179
99.5+
SLIMES
DRY
WEIGHT
3,775,400
3,333,400
\SJ
WATER
WEIGHT
12,880,000
14,200,000
Source: Reference 15.
Figure 7. Concentrating of copper ore by vat leaching and electrowinriing.
58
-------
Electrolytic Recovery of Copper. Pregnant liquor, containing 30 to 35
g/liter of copper, is sent to electrolysis cells. The electrolysis cell
operation reduces the copper concentration in the liquor from about 30 to 35
g/liter down to 5 g/liter as the copper is plated onto copper starter cathodes.
The operating liquor temperature in the electrowinning steps is held at about
27eC by steam heat. The copper cathode product contains 99.5+ percent copper»-l='
Barren solution from the electrowinning operation is combined with
sulfuric acid and used as leach solution in the vat leaching operation.
The available production data for this example process presented in
Figure 7 show the mass balance of materials. The only wastes are the partially
dewatered slimes which are produced in the vat leaching step and sent to an
on-site tailings pond. These slimes are not considered to be a potentially
hazardous waste since they contain only oxidized copper and may only be
leached by acid. There are no other waste streams in the concentrating
operations.
13/
In Situ Leaching.— In 1974, there were three domestic copper mines
which were using in situ leaching operations* There were 14,868 MT (16,390
tons) of copper concentrate produced, about 0.5 percent of the total copper
produced. In addition, three other mining companies are known to be doing
experimental work on in situ leaching.of copper ore.
One in situ leaching and precipitation method currently being used at
an Arizona copper mine is described herein as an example of this type of
operation. .The ore is a silicate copper, chrysocolla (CuSi03 • 2H20) and
malachite (CuC03 • Cu(OH)2). No sulfide minerals are present in this ore
body. The host rock is granite and schist, which if left undistrubed, is
impervious to water and acid. The copper content of this ore ranges from
0.2 to 0.6 percent. The flow diagram for this process is shown in Figure 8.
Elimination of conventional mining processing by in situ leaching
results in lower capital and operating costs. Steps eliminated include
loading, transporting, crushing, grinding, waste and tailings disposal,
and land reclamation.
In order to provide permeability of the copper ore deposit, the first
step in the process involves drilling and in-place blasting of the deposit.
This operation fractures and shatters the ore and prepares the ore body for
the percolating solution which must contact the ore.
In the example mine, the host rock (granite and schist) forms barriers
at the perimeter of the ore zone impervious to water and acid, i.e., a
solution control barrier.
59
-------
DRILLING- A BLASTING- OF
Of?E &ODY /W PLACE- TO
PREPARE POR LEACHIN&
SOLUTION
SITU LEACH I N&
(T)
LEACH LIQUOR
COLLBCT/NG POND
SCRAP
PRECIPITATION
Cu CEMENT
TV IN S'TU
LEACHING-
BARREN
SOLUTION
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEMS
Assay, gpl Cu
% Cu
PH
w •- •
CRUDE
ORE
0.2- 0.6
™\&-'
LEACH
LIQUOR
0.2 - 0.4
1.2- 1.7
VJ/ '
PREGNANT
LIQUOR
3-6
2.5-2.8
\gj
Cu
CEMENT
80
— — \^y
BARREN
SOL'N
0.2- 0.5
4.5
Source: Reference 15.
Figure 8. Concentrating of copper ore by in situ leaching and precipitation.
60
-------
Following the in-place blasting of the ore, a leaching solution
containing 2.5 to 5 percent H2S04 (pH of 1.2 to 1.7) is sprayed onto the
surface of the broken ore and allowed to percolate downward through the
ore. Process water is obtained from wells in the area. The pregnant copper
solution from the in-place leaching operation is collected in a pond at
the base of the leaching area. This pregnant liquor is pumped as required
to a precipitation plant.
At the precipitation plant, the pregnant solution containing copper
sulfate is contacted in cone precipitators with scrap iron and chemically
.detinned iron to precipitate copper cement from the solution. The chemical
reaction for this precipitation is:
Fe - > FeS0 + Cu
The cement copper is filtered on a vacuum filter and then further
dewatered by drainage on covered concrete slabs. The dewatered product
(80 percent copper) is shipped to a smelter.
Barren solution from the precipitation step is recycled to the
leaching site. Prior to being sprayed onto the surface of the ore bed,
the barren liquor is mixed with sulfuric acid and makeup water to provide
the desired composition in the leach solution. The available data for this
example process are insufficient to permit a mass balance. However, the
available operating process data are shown in Figure 8.
There are no wastes at an operating in situ site and precipitation
plant. When all the leaching operations are finally terminated, the in situ
leached ore becomes a waste. In situ leaching of a given site lasts for many
years. A thorough water rinsing of the leach area to remove the acid and
recover the copper in the acid solution is planned by the operating companies.
Secondary Processes. There are four secondary processes for recovery of
copper from leach solutions. All of these methods can be used with all of the
primary processes. The selection of the secondary process is guided by the
results of research and development by the individual mining companies and
the Bureau of Mines.
14/
Precipitation of Cement Copper. In the copper concentrating industry,
the three types of precipitation equipment used are cone precipitators
(conical- shaped vertical vessels), launders consisting of a series of open-
top concrete tanks, and vat precipitators.
In the example case, the pregnant liquor is precipitated on a continuous
basis in concrete launders. The pregnant liquor is contacted with scrap iron
in the launders to accomplish the following precipitation reaction:
61
-------
CuS04 + Fe > Cu + FeS04
In common with other cement copper operations in the industry, the
precipitation is carried out at ambient temperature without pH control.
The retention time for the solution in the launders is a few hours.
Continuous or batch operation may be used; most plants have continuous
operations. Uniform distribution of the pregnant copper solution through
the scrap iron is very important in producing a high-grade cement copper.
Precipitation is shown in two flow diagrams, Figure 6 and Figure 8»
The plant uses filtration followed by drainage and natural air-drying
on a concrete pad to dewater the copper cement which is flushed from the
launders. Other facilities conduct the drying operations using a drainage
pad or a gas-fired hot plate. During the drying period, the copper
precipitate is normally moved as little as possible to minimize the amount
of copper oxidation. The moisture content of the dried copper cement is
usually 15 to 20 percent.
In the precipitation plant, approximately 0.1 percent (20,000 MT/year)
of the tailwater from the precipitation plant is discarded to an evaporation
pond as a means of bleeding iron salts out of the leaching circuit. The
remaining tailwater is recirculated to the leach dump. The solids in the
bleed streams are contained in the evaporation ponds and are the only land-
disposed wastes produced from the precipitation of cement copper. No
analytical data are available for the solids in the bleed stream.
Solvent Extraction and Stripping Operations.—4— The liquid ion exchange
process consists of contacting the clarified metal-bearing aqueous solutions
with a water-immiscible, kerosene-soluble organic extractant. The flow
diagram describing the solvent extraction process is shown in Figure 5. The
organic extractant is in a kerosene solution. Copper along with some
impurities concentrates in the organic phase, which is subsequently separated
from the aqueous phase as the result of immtscibility. The metal-bearing
organic phase is then contacted with a new sulfuric acid solution to remove
the copper from the organic extractant, resulting in a copper sulfate
concentrate.
The chemical reaction taking place in the extraction process can be
expressed by the following equation, using the letter R to represent the
organic extractant: CuSO^ + 2HR = CuR2 + H2SCV This reaction shows that
in aqueous solutions of copper sulfate at low acid concentrations, the
organic will combine with the copper ions to form a loaded organic and will
regenerate the sulfuric acid which was associated with the copper. There are
three contact stages, each consisting of a mixer and settler. Spent acid
solution from the extraction step is sent to heap leaching. The purpose of
solvent extraction is to obtain a purified copper sulfate solution suitable
for use in electrolysis.
62
-------
The loaded organic is stripped of copper by treatment with strong sulfuric
acid solution, and stripped organic is recycled to the extraction step. The
chemical reaction involved in stripping is the reverse of the extraction
reaction and is expressed by the following equation: CuR2 + ^SO, = CuSO/ +
2HR. This reaction demonstrates that when loaded organic is contacted with a
high acid strength aqueous solution, the copper is stripped from its organic
form into an aqueous copper sulfate solution, and the organic is returned to
its hydrogen form, ready to be recontacted with, and extract copper from,
slightly acidic aqueous leach liquors.
No significant land-disposed wastes are produced in solvent extraction
and stripping operations. All of the liquid streams are recycled to the
process.
Electrowinning• ' Electrowinning is carried out in electrolysis cells
using purchased starter cathode sheets. Each cell receives pregnant
electrolyte at the same rate. The cell solutions (spent electrolyte)
overflowing at the opposite end from the feed stream are combined and
returned to the stripping operation. Flow diagrams for electrowinning are
shown in Figures 5 and 7.
Each cell contains 40 cathodes and 41 anodes. The anodes are 6 percent
antimonial lead, and are spaced on 11.4-cm (4.5-in.) centers, with all anodes
in any one cell resting on the same bus bar in order to be in parallel
electrically. Similarly, all the cathode support bars in that same cell rest
on the opposite bus bar. The power consumption is about 0.55 kw-hr/kg of
copper.
Copper is deposited on the starter cathodes. After being pulled from
the cells, the finished cathodes are soaked in hot water dip tanks and then
banded in bundles of 20 cathodes for shipment.
Typical metallurgical data for electrowinning are shown in Figure 5.
The product cathodes contain 99.9+ percent copper.
No waste streams of any significant quantity or potential pollution
hazard are involved in the electrowinning operations. All liquid streams
are recycled in the process.
Identification of Potentially Hazardous Materials. Table 21, p.41, shows
the total and potentially hazardous wastes from the mining and concentrating
of copper ores. In 1974, there were 140,179,000 MT (154,520,896 tons) dry
weight and 461,242,000 MT (508,432,269 tons) wet weight of potentially
hazardous wastes resulting from concentrating copper ores.
Tables 22 and 23, p.42 and 43 respectively, show the 1977 and 1983
projections for total and potentially hazardous wastes resulting from the
mining and concentrating of copper ore.
63
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Waste Treatment and Disposal
Mining Waste Treatment and Disposal* There are no potentially hazardous
wastes destined for land disposal from either underground or open-pit mining
of copper ores. The waste rock and overburden that are disposed of on land
are either at or below the background level of the minerals present in the
mine and the adjacent area. The background concentration data for areas
adjacent to the mines were obtained by the U.S. Geological Survey and the
University of Idaho.—'
Concentrating Process Wastes. There are several different concentrating
processes used to recover copper from the mined copper ore. Table 24, p.49, is a
presentation of the different concentrating processes. The table also shows
whether or not waste is generated (i.e., flotation tailings or iron salts
left behind in an evaporation pond), whether or not pyrite (FeS2) is present
in the ore, and also indicates if the waste is a potential hazard. Table 19, p.39,
shows the number of mines using each process, the total ore mined and copper
and waste produced from each process, and the amount of potentially hazardous
waste generated.
Flotation. In Table 24, Process Code A shows that the ore is a sulfide, with
pyrite present; the process is flotation; the waste material is tailings; and
that pyrite is a potential hazard. There are 23 mines mining 148,406,000 MT
(163,247,000 tons), producing 1,070,224 MT (1,177,000 tons) of copper, and
140,179,000 MT (154,200,000 tons) of tailings. Because pyrite is present in
all the mines in this group, the tailings are considered potentially
hazardous, and they constitute 64 percent of the total concentrator wastes
from copper recovery processes.
A tailings dam contains concentrator waste which may be potentially hazardous
material generated from flotation; therefore, it is important that the dam be
constructed to prevent rupture.
Table 26, p.51, shows analytical data on tailings solids reported by all
companies contacted by personal visit and through the American Mining Congress
questionnaire. As can be seen, very little information was available, and
very little interpretation can be made from these data.
The flotation wastes from the 23 mines are disposed of in tailings ponds.
There are differences from company to company in the preparation of a tailings
pond, in the method of addition of the tailings slurry to the pond, and in the
methods used to stabilize the dams and pond surface to assure retention of the
contaminants in the tailings pond.
64
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Some companies remove all vegetation from the surface of the area to be used
as a tailings pond, and a few companies compact the surface or add material to
make the surface impermeable to water. In some copper mining areas of the
country, clearing of the vegetation and sealing the surface of the ground is
unnecessary, as the tailings ponds are located on top of unvegetated impermeable
soil or rock.
The most common method of addition of slurry to the tailings pond is to pump
the slurry to the pond, discharging it to the surface of the pond some distance
away from the dikes. A few companies add the slurry to the inside face of the
dam and utilize the differences in settleability of the fines and coarse to
build up the face of the dam with coarser sand. A few companies separate the
coarser sand from the fines by using a cyclone, discharging the fines on the
inside face of the dike, which aids in sealing the faces of the dikes and dams
against leakage. The coarse material is discharged on top of the dam and
utilized in raising its level.
Stabilization methods used for the dam and pond surface vary from just
keeping water in the pond to extensive measures for stabilization. A few
companies have built backup dams downstream from the tailings pond to collect
seepage and pump it back into the pond. Several companies are vegetating the
surface of the dam and putting in erosion-protection furrows.
One company is compacting the dam continuously and extensively vegetating
the face of the dam and the surface of the dam and pond when it is no longer
in use.
Some companies that operate underground copper mines also operate sand
plants. In these plants, the fines are separated from the coarse material
and are pumped to the tailings pond. The coarse material is pumped to the
mine where it is mixed with concrete and used to backfill the mine.
The filling of depleted mine stopes with tailings sands from the concentrator
waste stream is a we11-developed system at these mines. Lumber cut to size in
the mine is used to prepare skeleton floor and wall-pouring forms. Wire
fencing, burlap, and plastic sheets are used to cover the wooden forms and
serve as a seal during the actual sand-filling operation. The tailings sand
used as fill material is mixed with cement at the mine surface and pumped down
to the fill areas. The bottom 1.22 m (4 ft) of fill consists of 12:1 ratio of
sand-to-cement and the remainder has a ratio of 20:1 sand-to-cement. As much
water as possible is drained from the sand fill before it sets, in order to
obtain a high density and a strong fill. The poured fill sets up in about 24
hr. After the fill has set, the adjacent pillars are mined out to improve the
ore recovery from the mine operations.
65
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Leaching. In leaching of ore with sulfuric acid, there is no potentially
hazardous waste generated,
Electrowinning. There is no potentially hazardous waste generated in
electrowinning of copper.
Levels I, II, and III Technology. Level I technology for disposal of
potentially hazardous wastes from copper flotation plants is the use of
tailings ponds; all flotation concentrators use tailings ponds. Figure 9
shows the waste disposal method for Level I technology used by all concentrators.
The potentially hazardous wastes, dry solids, amount to 82 percent of the ore
processed in the concentrator. This material is disposed on land by pumping the
solids and water to a tailings pond. The slurry pumped to the pond is roughly
31 percent solids. Level II technology for disposal of potentially hazardous
wastes is the vegetative stabilization of the dams and dikes and attention to
surface preparation of the area before establishment of a tailings pond, plus
the compaction of the dam and dike to prevent erosion, and treatment with lime
or limestone to neutralize the effect of the pyrite. Figure 10 shows the waste
disposal method for Level II technology. This method is presently used by six
mines. The material balance for Figure 10 indicates that 41 percent of the ore
processed is pumped to the tailings pond with the water used in the flotation
process. Another 41 percent of the potentially hazardous waste material is
pumped back to the underground mine, mixed with cement and used to backfill
the mine stopes. The slurry pumped to the tailings pond is about 33 percent
solids and contains the fine, smaller than 65 mesh, material.
Level III technology for safe disposal is to incorporate effluent wastewater
treatment. Two of the largest mines are treating the effluent from the pond in
a wastewater treatment plant (and returning the solids precipitated from the
water to the tailings pond) before discharging the effluent. This practice
reduces or eliminates the effect of acid formation due to pyrite and subsequent
leaching of metals in the tailings. Figure 11 shows the waste disposal
techniques for Level III technology. The solids being returned from the
wastewater plant cannot be quantified at this time. The wastewater treatment
plants did not operate until the last quarter of 1974, and solids were not
returned to the tailings pond during 1974.
The enforcement of air and water pollution regulations for 1977 and 1983 will
have only a very small effect, less than 1 percent, on the tonnage of potentially
hazardous wastes disposed on land from copper concentrating operations. There are
very few operations in a copper flotation plant that generate an air pollutant.
If dry grinding of ore is subject to air regulation, wet grinding of the ore is
an obvious solution. Most concentrators wet grind at the present time.
66
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MINE
Ore
10,000 MT
CONCENTRATOR
(FLOTATION)
Water
26,600 MT I
Product
1.800MT
Tailings
8.200MT
TAILINGS POND
Source: Reference 15.
Figure 9. Level I technology for disposal of copper concentrator wastes.
67
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MINE
Products -*
1.800MT
Ore
10.000MT
CONCENTRATOR
(FLOTATION)
Water
Tailings
8,200 MT
SAND PLANT
Water
12.300MT
Fines
4.100MT
.Cement
Coarse
4.100MT
TAILINGS POND
Revegetate
Compacts
Source: Reference 15.
Figure 10. Level II technology for disposal of copper concentrator wastes,
68
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TAILING'S
. 9.6OOMT
PIT
10. OOO A/7
CONCENTRATOR
(FLOTATION)
14. 200 MT
C.4TCH &AS//V
71OO MT
i TME.NT PLANT
SOLIDS
Figure 11.
wastes.
Source: Reference 15.
Level III technology for disposal of copper concentrator
69
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Effluent limitations arising from enforcement of the guidelines will
require removal of trace metals from the effluent water. These solids will
add to the potentially hazardous waste destined for land disposal, but
should contribute < 0.1 percent of the total wastes in 1977 and 1983. Zero
discharge will require that all concentrators install wastewater treatment
to remove traces of metal ions now present in the discharge. New technology
will have to be developed to remove all cations from the effluent.
All technology identified in the program can be transferred to any copper
mine where it is necessary to properly dispose of potentially hazardous waste
materials forfthe protection of the environment. Delivery of equipment and
construction required to enable a concentrator to move from Level I to Level
III technology could require 3 to 5 years.
The application of Level III technology to disposal of flotation tailings
waste where pyrite is present would mean the consumption of considerable
energy and some investment in equipment. There are about 18 concentrating
plants that will have to install wastewater treatment facilities, pumps to
deliver the effluent to the treatment plant, and solids handling equipment
to transfer the treatment plant sludges back to the tailings pond.
Table 18 shows the total wastes disposed on land by copper concentrators.
There were 240,794,000 MT dry weight (265,430,000 tons) of solids disposed
of on land in 1974. There are no radioactive materials in the copper wastes.
The copper and other trace materials present in the tailings at certain
concentration levels are potentially hazardous and'must be kept in the
tailings pond. The material in the pond is about 60 to 80 percent -60 mesh
and has the physical appearance of sand. There is no reliable information
available on the chemical composition. Most tailings will contain from 100
to 1,500 parts per million of copper. Table 26 contains all of the analytical
information furnished by 15 mines on copper concentrator tailings.
The bulk of the material in the tailings is not soluble in water and most
other solvents, including acids. Copper is insoluble until the pH of the
contacting solution is below 3.5. All metals found in the tailings have a
very limited solubility between 3.5 and 7 pH, and the solublity above 8 pH
is below the recommended drinking water standard. If the solutions coming
in contact with the tailings pond are above 7 pH, very little hazard exists.
There are no intermedia pollution problems where Level II or Level III
technology is implemented. The only problem is in tailings ponds where a
considerable quantity of pyrite is present. In this case, sulfuric acid
could be made by water and air contact with the FeS2 (pyrite), and the
sulfuric acid would leach some metals from the tailings, unless lime or
other caustic material is added in the flotation process.
70
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The only places where sampling is needed is where pyrite is present at
concentrations > 2 percent in the ore and is not neutralized by lime addition.
Samples of the effluent water should be taken monthly and analyzed for pH and
metals. Stabilization techniques and progress toward stabilization should be
monitored by quarterly visual surveillance of the tailings pond.
Waste Disposal Costs - Copper
Waste Disposal Costs, Tailings Pond Using Level I Technology. The
representative copper mine used in the cost analysis produces 834,610 MT
(918,000 tons) of ore per year. The representative plant chosen for the cost
estimate is a model plant, based on average figures and practices of several
copper mines. A material balance was calculated for a number of concentrators,
all using the same process, flotation, and all using a tailings pond for
disposal of their potentially hazardous waste. An average ore grade of 1
percent was used in calculating the material balance. A concentrate with 25
percent copper was also used. After the material balance data were derived, a
model plant was designed and used as the representative plant. A model plant
was used to ensure that a particular mine and concentrator could not be
identified. After the concentrating process, 442,706 MT (487,000 tons) of dry
tailings per year are deposited in the tailings pond along with 4,470,000 MT
(4,920,000 tons) of water per year. The tailings pond covers 12.1 hectares
(30 acres) and is located within 0.8 km (0.5 miles) of the concentrator. The
major assumptions of the analysis are presented in Table 27. Many of the costs
are similar to those associated with the uranium tailings pond.
The results of the analysis are presented in Table 28. Two cost estimates
are presented. The differences arise due to different assumptions about the
value of the land after 20 years. The costs per metric ton of disposal of
these potentially hazardous wastes are estimated to be between $0.275 and
$0.282/MT ($0.25-$0,26/ton). Costs estimated by a copper mining corporation
using a tailings pond of approximately the same size as the representative
plant equal $0.38/MT ($0.35/ton). If the entire copper mining and concentrating
industry utilized technology Level I, the cost would be 6.6 to 6.8 million
dollars per year. This cost represents less than 1 percent of the value added
by the copper mining industry.
Waste Disposal Costs, Tailings Pond Using Level II Technology. The disposal
of potentially hazardous wastes from copper concentrators is more expensive
using Level II than Level I technology. In addition to the tailings pond,
Level II technology also requires:
71
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TABLE 27
MAJOR COST ASSUMPTIONS, POTENTIALLY HAZARDOUS WASTE FROM
COPPER CONCENTRATORS — TECHNOLOGY LEVEL I
I. Capital
A. Land
12.1 ha (30 acres) are required for pond*
Land cost at $12,350/ha ($5,000/acre) = $150,000
Levelized annual land cost (capital recovery factor of 0.117) =
$17,550/year
B. Tailings Pump and Pipeline
Tailings pond assumed to be within 0.8 km (0.5 mile) of concentrator*
Tailings pump and pipeline cost = $393,016t
Levelized annual cost at capital recovery factor of 0.117 = $45,982/year
C. Earth Dam
Earth dam capital cost = $35,000
Levelized annual cost - $4,095/year
D. Cyclone Installation
Capital cost of cyclone = $58,000
Levelized annual cost = $6,786/year
.II. Labor
1 operator required for cyclone separator for 5 months., 8 hr/day at
$8.97/hr§ '• '
Labor cost = $7,176/year
III. Supervision
Assumed to be 25% of labor cost§
Supervision = $l,794/year
IV. Maintenance
Assumed to equal 670 of nominal cost of equipment/year§
Maintenance cost = $27,061/year
72
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TABLE 27 (CONCLUDED)
V. Insurance and Taxes
Assumed to equal 270/year of total nominal capital cost (land and
equipment)§
Insurance and taxes = $12,720/year
VI. Energy and Power
A. Tailings pump is assumed to consume electricity with a value
of approximately $l,400/year**
B. Cyclone separator mounted on truck—only fuel required is for
truck as it travels around the pond. Assumed to be less than
$100/year
* MRI communication from individual copper mining corporation.
t Sears, M. B., et al. Correlation of radioactive waste treatment costs
and the environmental impact of waste effluents in the nuclear fuel cycle
for use in establishing as low as practicable guides - Milling of uranium
ores. Oak Ridge National Laboratory, Oak Ridge, Tennessee, ORNL-TM-4903,
v.l, May 1975. p. 187-188. Obtained by ratio of material handled (dry
millings and water) to total cost for pump and pipeline.
§ See corresponding table for uranium tailings pond.
** Obtained in the same manner as the capital cost of the tailings pump
and pipeline (see footnote t).
73
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TABLE 28
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE FROM COPPER
CONCENTRATORS, TAILINGS POND, TECHNOLOGY LEVEL I
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Capital (levelized cost)
Land
Other capital
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Total
Case "A"*
17,550
56,863
7,176
1,794
27,061
12 , 720
1,500
124,664
Case "B"+
14,672
56,863
7,176
1,794
27,061
12 , 720
1,500
121,786
Metric tons of potentially hazardous waste
deposited in pond annually 442,706 442,706
Cost per metric ton of potentially hazardous
waste $0.282 $0.275
Total cost if entire industry uses the
technology level ($106/year) $6.8 $6.6
Total cost as a percent of value added <
in mining* 0.67. 0.6%
* "A" assumes all land is purchased in Year 1 and has 0 resale value
after 20 years.
t "B" assumes all land is purchased in Year 1 and is resold at the same
value after 20 years.
$ 1972 Census of Mineral Industries, Bureau of Census, Government Printing
Office, 1975 corrected for inflation to 1973.
74
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* Vegetative stabilization of dams and dikes;
* Attention to surface preparation before tailings pond is built;
* Compaction of dams and dikes to prevent erosion;
* Treatment with lime or limestone.
Assumptions concerning each of these additional cost factors are presented
in Table 29. Table 30 presents the results of the analysis. Vegetative costs
per hectare are .substantial. However, when the cost is distributed over 20 years
of analysis, the annual equivalent costs are small. Surface preparation and
compaction costs can also be significant, even on the 12.1-ha (30-acre)tOilings
pond. Lime treatment costs, on the other hand, are part of the concentration
process. No cost has been calculated for this treatment, because it is not done
for disposal purposes. The results indicate that Level II technology disposal
costs are approximately $0.32 to $0.33/MT of potentially hazardous wastes.
Total industry costs, assuming all companies utilize Technology Level II,
would be 77.1. to 79.4 million dollars per year. This cost would amount to
7.2 percent of the value added in mining and concentrating copper ores.
Waste Disposal Costs, Tailings Pond and Wastewater Treatment Plant Using
Level III Technology. Level III technology differs from Level II waste
disposal techniques in that a wastewater treatment plant is utilized.
Effluent from the tailings pond is treated in the plant. Solids precipitated
from the water are returned to the tailings pond by truck, and the treated
water is discharged.
The additional assumptions used in calculating the Level III technology
costs are presented in Table 31. Capital, operation, maintenance, materials,
and other costs are presented in the table. The results of the analysis are
presented in Table 32. All costs incurred in Level II technology are also
present in Level III, in addition to the wastewater treatment costs. Table
32 indicates that the disposal costs of potentially hazardous wastes range
from $0.522 to $0.516/MT ($0.475-$0.469/ton) of waste. Total industry costs
would range from 125.7 to 124.3 million dollars per year or 11.4 percent
to 11.3 percent of the total value added in mining and concentrating copper
ores.
75
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TABLE 29
ADDITIONAL COST ASSUMPTIONS--TAILINGS POND-
COPPER CONCENTRATORS—TECHNOLOGY LEVEL II
I. Vegetative Stabilization of Dams and Dikes
1.05 ha (2.6 acres) of berm are assumed to require vegetative Stabi-
lization*
Cost of vegetative stabilization assumed to be $3,213/ha ($1,298/
acre)*
Vegetative stabilization cost = $3,375
Levelized cost per year = $395/year
II. Surface Preparation Before Pond Construction
12.1 ha (30 acres) of pond area are assumed to be bulldozed and com-
pacted
Time required: approximately 16 hr of bulldozing and compacting at
3 mpht
Total cost of surface preparation:
1. Labor at $8.97/hr* = $142.52
Levelized equivalent annual labor cost = $17/year
2. Energy and power (60 gal/8 hr, at $0.50/gal)t = $60.00
Levelized equivalent annual cost = $7.00/year
III. Compaction of Dams and Dikes to Prevent Erosion
A. Capital
One compactor at $70,000 required with useful life of 10 years§
Present value of compactor capital costs at 10% discount rate =
$97,020
Equivalent annual cost = $11,351/year
B. Labor
Compactor assumed to be used 8 hr/week§
Annual labor cost at $8.97/hr = $3,732/year
76
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TABLE 29 (CONCLUDED)
C. Maintenance
Assumed to equal 6% of nominal capital cost or $4,200/year
D. Energy and Power
Compactor assumed to consume 3 gal/hr at $0.50/gal*
Annual energy plus power costs = $624/year
* Ludeke, K. Vegetative stabilization of tailings disposal berms.
Mining Congress Journal, January 1973. p. 32-39. Berm area obtained
from ratio of berm requiring vegetation to tailings pond size obtained
from Pima Copper Mine.
t MRI estimate for 30 acre rectangular tailings pond.
* See previous discussion of uranium open-pit mine disposal costs.
$ MRI estimate.
77
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TABLE 30
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE—COPPER INDUSTRY-
TAILINGS POND, TECHNOLOGY LEVEL II
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Capital
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Case "A"*
85,764
10,925
2,732
31,261
12,720
2,131
Case "B"t
82,886
10,925
2,732
31,261
12,720
2,131
Vegetation of dams and dikes
Total
Metric tons of potentially hazardous
tailings deposited in pond annually
Cost per metric ton of potentially
hazardous waste to tailings pond
Total cost if entire industry used the
technology level ($106/year)
Total cost as a percent of value added
in mining
395
145,928
442,706
$0.330
$79.4
7.2%
395
143,050
442,706
$0.323
$77.1
7.2%
* Case "A" assumes all land is purchased in year 1 and has 0 value
after 20 years.
t Case "B" assumes all land is purchased in year 1 and is resold at
the same value after 20 years.
78
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TABLE 31
ADDITIONAL COST ASSUMPTIONS—COPPER TAILINGS POND--
TECHNOLOGY LEVEL III
Wastewater Treatment Plant
A. Capital
4,470,000 MT/year of water enter the pond
50% evaporation assumed, 25% of water assumed to be held in solids
and 25% assumed treated as effluent*
Therefore, 809,000 gal/day will enter plant
Plant contains two systems: agitated open tank and sedimentation
system
1. Agitated tank capital cost = $90,000 (16 ,860 gal. capacity)!
Useful life = 20 years*
2. Sedimentation system capital cost = $50,000 (809,000 gal/day)§
Useful life = 40 years§
Using S.L. depreciation, net capital cost = $46,900
Equivalent annual cost (tank and sedimentation) = $16,017/year
B. Operation and Maintenance (includes energy and power)
1. Agitated tank operation labor costs, 4,380 hr/year at $8.97/hr
$40,000*
Agitated tank maintenance costs = $9,000/year
2. Sedimentation system operation costs are minimal§
Sedimentation system maintenance costs = $5,250/year§
C. Materials
60 kg of dolomitic lime required per day**
Lime cost equal $22/MTtt
Annual lime cost = $482/year
D. Other cost
Assumed one load of sludge per day is trucked back to tailings pond
Labor cost = $3,274/year
* The amount of water treated can vary from 10 to 25 percent of water
entering the pond.
. t Blecker, H., and T. Nichols. Capital and operating costs of pollu-
tion control equipment modules. Environmental Protection Agency, v.II,
EPA-R5-73-0236, July 1973. p. 10.
* Blecker, H., and T. Nichols. Capital and operating costs of pollu-
tion control equipment modules. Environmental Protection Agency, v.II,
EPA-R5-73-0236, July 1973. p. 11.
§ Blecker, H., and T. Nichols. Capital and operating costs of pollu-
tion control equipment modules. Environmental Protection Agency, v.II,
EPA-R5-73-0236, July 1973. p. 126-127.
** MRI estimate.
tt Environmental Protection Agency. Development document for effluent
limitation guidelines and new source performance standards for the major
inorganic products. EPA-440/l-74-007-a, March 1974. p. 275.
79
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TABLE 32
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE—COPPER CONCENTRATOR
TAILINGS POND AND WASTEWATER TREATMENT PLANT,
TECHNOLOGY LEVEL III
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Capital
Wastewater treatment
Other capital
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Vegetation of dams and dikes
Materials
Total
Metric tons of potentially hazardous
tailings deposited in pond annually
Cost per metric ton of potentially
hazardous waste
Total cost if entire industry utilized
the technology level ($106/year)
Total cost as a percent of value added
in mining
Case "A"*
16,017
85,764
54,199
13,550
45,511
13,040
2,131
395
482
231,089
442,706
$0.522
$125.7
11.47o
Case "B"t
16,017
82,886
54,199
13,550
45,511
13,040
2,131
395
482
228,211
442,706
$0.516
$124.3
11.3%
* Case "A" assumes all land is purchased in Year 1 and has 0 resale
value after 20 years.
t Case "B" assumes all land is purchased in Year 1 and is resold at
the same value after 20 years.
80
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REFERENCES
1. U.S. Department of the Interior, Bureau of Mines. Unpublished Commodity
.Data Report, Copper (H. J. Schroeder), Dec. 1973.
2. U.S. Department of the Interior, Bureau of Mines. Commodity Data
Summaries. 1974, Appendix I to Mining and Minerals Policy.
3. U.S. Bureau of Mines. Metals, minerals, and fuels. 1971 Minerals
Yearbook, 1:461.
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Yearbook. 1:460.
5. Engineering & Mining Journal, 1973-1974, International Directory of
Mining and Mineral Processing Operations, Published by Engineering &
Mining Journal, McGraw-Hill, New York, New York.
6. Mining Enforcement and Safety Administration. Unpublished data.
7. Ward, M. H. Engineering for in situ leaching. Mining Congress Journal,
pp. 21-27, Jan. 1973.
8. Ridge, J. D. Ore deposits of the United States. The American Institute
of Mining, Metallurgical, and Petroleum Engineers, Inc., First Edition,
New York, 1968.
9. Williams, R. E. The role of mine tailings ponds in reducing the discharge
of heavy metal ions to the environment. PB 224, 731. Department of the
Interior, Idaho University, Moscow, Idaho, Aug. 1973.
10. Power, K. L. Operation of the first commercial copper liquid ion exchange
and electrowinning plant. Ranchers Exploration and Development
Corporation, Miami, Arizona, 1970.
i !
11. Hannan, W. S., Jr. Leach dump operations at Bisbee. Phelps Dodge
Corporation, Copper Queen Branch, Bisbee, Arizona, Oct. 5; 1971.
12. Miller, A. Process for the recovery of copper from oxide copper bearing
ores by leach liquid exchange and electrowinning at Ranchers Bluebird
Mine. Ranchers Exploration and Development Corporation.
13. Ward, M. H. Engineering for in situ leaching Ranchers Exploration and
Development Corporation. Miami, Arizona, Jan. 1973.
81
-------
14. Jacky, H. W. Copper precipitation methods at Weed Heights* Mining
Engineering, Society of Mining Engineers, June 1967,
15, MRI communications. Copper raining companies.
82
-------
SECTION II
LEAD-ZINC ORES (SIC 1031)
Industry Characterization
History of the Industry. Lead is one of the oldest metals used by man, and
many of the ancient applications have persisted through the centuries. The
earliest known specimen of lead, dating from 3000 B0C0, is a figure found in
the area of the Dardanelles on the site of the ancient city called Abydos.
Old lead pipes have been found in Egypt, and the hanging gardens of Babylon
were floored with soldered sheets of lead.
The lead mines of Cyprus, Sardinia, and Spain were worked by the Phoenicians,
and the lead-silver mines of Laurium, Greece, are famed as the source of
ancient treasures. The great quantities of silver that assisted in the rise
of Rome were derived from lead smelting-refining in Britain, Sardinia, and
Spain.
In the United States, lead was mined and smelted in Virginia as early as
1621, and the discovery of lead in the Upper Mississippi Valley was reported
in 1690. These shallow, easily processed ores were important sources of
domestic lead during much of the 19th century. The production averaged about
907 MT (999 tons) annually from 1801 through 1810 and increased to about 2,721
MT (2,999 tons) per year from 1831 to 1840, some 10 percent of the world
output. In 1867, discoveries in southeast Missouri led to development of one
of the most productive lead-zinc areas in the world. Since 1904, southeastern
Missouri has been the leading producing lead-zinc district in the United States.
Lead is easily worked and noncorrosive, and since medieval times has been
used in piping, building materials, solders, paints, type, ammunition, and
castings. The advent of the electric starter on internal combustion engines
was followed by a large increase in consumption, which was expanded by the
need for antiknock gasoline additives for use in high compression engines.
United States consumption of lead first reached 1 million metric tons
(1,102,311 tons) in 1941 and has been well above that figure in most years
since. Lead consumption amounted to nearly 1.5 million metric tons (1,653,466
tons) in 1973.
In most lead mines the ore is composed of a combination of lead and zinc
sulfides and the concentrators recover both lead and zinc concentrates.
83
-------
Today, the United States is the world's leading producer of lead and also
the leading consumer* Lead mining and refining is a major basic industry and
ranks fifth worldwide, following steel, aluminum, copper, and zinc in tonnage
of metal produced* Some 53 countries, well distributed throughout the world,
produce lead. In 1973, the leading producers were the United States, U«S.S0R0,
Australia, Canada, Mexico, Peru, and Yugoslavia.
Domestic Production and Capacities. During the period 1963 to 1973, United
States lead production increased from 229,000 MT (252,429 tons) to 548,000 MT
(604,067 tons) (recoverable content of ore)* (Table 33). This increase was
the largest in the world during the 11-year period, and the United States is
the leading world producer. While Government statistics are not available
for 1974, it is estimated that United States production of lead concentrate
product was 1,231,652 MTt (1,357,664 tons) (Table 34), Table 35 gives the ore
produced by year, while Tables 36 and 37 give the mine production of recoverable
lead by state and region, respectively.
TABLE 33
LEAD PRODUCTION, 1963-1983 (1,000 MT)*
Year
1963
1964
1965
1966
1967
1968
1969
1970
1971
1972
1973
1977t
1983t
United States
229
259
273
296
287
325
461
518
524
561
548
585
643
World total
2,544
2,481
2,691
2,847
2,865
2,992
3,196
3,392
3,403
3,491
3,445
__
--
Source: References 1 and 2.
* Data published in tons and calculated to metric tons.
t Estimated production.
* Recoverable content of ore = the amount of metal produced after the
recovery processes have been completed.
t Recovered lead concentrate product = the amount of concentrate produced
before smelting and refining the ore.
84
-------
TABLE 34
U.S. LEAD CONCENTRATE PRODUCTION FOR 1974 (RECOVERED LEAD
CONCENTRATE PRODUCTS)
State
Metric tons
California
Colorado
Idaho
Illinois
Kentucky
Missouri
Montana
New Mexico
New York
Utah
Virginia
Washington
Wisconsin
U.S. total
EPA regions
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
20,629
69,887
150,868
57,930
2,268
638,897
6,668
6,000
226,796
31,751
3,629
1,814
14,515
1,231,652
.
__
226,796
3,629
2,268
72,445
6,000
638,897
108,306
20,629
152,682
Source: Reference 3.
85
-------
TABLE 35
TOTAL LEAD MINE PRODUCTION OF ORE BY YEAR
(ORE PRODUCED)
Year 1,000 MT
1967 287,514
1968 325,820
1969 461,768
1970 518,698
1971 524,851
1972 561,470
1973 547,054
Source: References 4 and 5.
86
-------
TABLE 36
MINE PRODUCTION OF RECOVERABLE LEAD IN THE UNITED STATES, BY STATE
(MT)
State
Alaska
Arizona
California
Colorado
Idaho
Illinois
Kansas
Maine
Missouri
Montana
Nevada
New Mexico
New York
Oklahoma
Oregon
South Dakota
Utah
Virginia
Washington
Wisconsin
Other states
Total
1969
2
197
2,284
19,747
59,509
718
358
--
322,461
1,590
1,288
2,148
1,530
549
*
1
37,496
3,046
7,846
1,000
--
461,768
1970
__
259
1,608
19,827
55,530
1,390
73
--
382,618
904
330
3,221
1,161
723
*
3
41,165
3,045
6,154
690
--
518,698
1971
__
779
2,072
23,356
60,428
1,123
--
--
389,757
558
101
2,695
796
'
—
—
34,718
3,072
4,696
682
18
524,851 .
1972
— —
1,599
1,046
28,437
55,707
1,211
--
77
443,973
260
*
3,250
988
--
—
—
18,784
3,122
2,329
687
--
561,470
1973
5
692
40
25,503
56,013
491
--
185
441,929
160
—
2,319
2,090
--
'
--
12,458
2,392
2,011
766
--
547,054
Source: Reference 4.
* Less than one-half unit.
87
-------
TABLE 37
MINE PRODUCTION OF LEAD IN THE UNITED STATES, BY EPA REGION FOR 1973
Region MT
I
II
III
IV
V
VI
VII
VIII
IX
X
185
2,090
2,392
--
1,257
2,319
441,929
38,121
732
58,029
Source: Reference 4.
The available supply of lead has exceeded U.S. industrial demand, resulting
in a buildup in stocks. The 3 percent average annual growth in demand since
1965 has been due primarily to the growth of transportation requirements
for batteries. This growth in transportation end product applications has been
in contrast to the relatively stable tonnage or even declining consumption
for many historical end products used in construction and communications. The
Environmental Protection Agency's regulations reducing the lead content of
gasoline by 60 to 65 percent over a 4-year period were scheduled to become
effective on January 1, 1975, and will curtail growth in demand for lead.
Domestic demand for lead is forecast to increase at an annual rate of
approximately 1.6 percent through 1983. To meet the projected demand, lead
production is expected to increase to 585,000 MT (644,852 tons) of metal in
1977 and 643,000 MT (708,786 tons) in 1983. The domestic resource base is more
than adequate to support the domestic demand at competitive prices, and •
reliance on imports probably will decline from the current level of about 30
percent of total primary metal consumption.
Number, Location, Size, and Age of Mines and Concentrators. The domestic
lead-zinc mining industry in 1974 comprised 30 mining operations, with about
39 individual mines and 32 concentrators located in 11 states (Table 38).
Lead-zinc mines vary in size from small operations employing fewer than 50
workers to mines employing more than 1,000 workers. Recent data from the
Engineering & Mining Journal and the Mining Enforcement and Safety
Administration indicate that 74 percent of the operations employed more
than 100 workers in 1974 (Table 38).
88
-------
TABLE 38
NUMBER, LOCATION AND SIZE OF ACTIVE LEAD-ZINC MINING AND CONCENTRATING OPERATIONS
State
California
Colorado
Idaho
Illinois
Kentucky
Missouri
New Mexico
New Jersey
New York
Utah
Virginia
Washington
Wisconsin
U.S. total
EPA regions
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region DC
Region X
Employees
1-19 20-49 50-99 100-249 250-499
1
2 2
111
2
2
4
1 1
1
1
1
5 5 11
2
1 2
1 1
4
2 2
1
121
2
1
3
1
1
1
9
1
1
1
3
3
1
Total
500-999 1,000-2,499 2,500+ operations
1
6
1 5
2
2
7
3
1
1 1
1
1
1
1
11 32
0
1 2
I
2
3
3
7
7
1
1 6
Total number
of mines
0
6
6
2
2
10
3
1
2
2
2
2
1
39
0
3
2
2
3
3
10
8
0
8
Total number of
concentrators
1
6
5
2
2
7
3
1
1
1
1
1
1
32
0
2
1
2
3
3
7
7
1
6
Source: References 3 and 6.
-------
The ages of active lead-zinc mines vary considerably; some have been in
operation more than 90 years, while others have only recently opened. The
Bunker Hill mine in Idaho was opened in 1885, while the Cerro Spar Corporation
mine in Kentucky was opened in 1974. The year of mine openings is not known
fot all active lead-zinc mines, however, data for the 17 mines for which
information is available, indicate that active lead-zinc mines are comparatively
young in age, with 14 mines opening after 1940 and nine mines since 1960.
In 1972, the 25 largest lead-zinc mines accounted for over 95 percent of
domestic production, and the leading seven mines accounted for 74 percent of
the total domestic primary production. Currently, domestic lead-zinc mine
production is located in 11 states (Figure 12); this figure also shows the
zinc mines. Five States—Missouri, Idaho, Colorado, New York, and Utah—
produced 73 percent, with Missouri contributing 50 percent of the total
recovered lead concentrate products in 1974. The world lead-zinc mining
industry operated at an estimated 86 percent of capacity in 1972. In the
United States curtailment of production to reduce inventories, as well as a
shortage of skilled underground personnel, held output about 3 percent below
capacity (Tables 39 and 40). A major increase in lead-zinc mine capacity was
affected during the last 5 years. The increases are mainly expansion of
existing mines, although one new mine in Missouri will offset reduction in
output at presently operating mines.
Employment. Total estimated employment in the 38 active lead-zinc mines
in 1974 was approximately 6,100 according to figures issued by Engineering &
Mining Journal, Mining Enforcement and Safety Administration, and from bn-
site investigations made by Midwest Research Institute. More than 50 percent
of the lead-zinc workers were employed in Missouri, Idaho, and Colorado
where the major lead mines are located (Table 41).
By-Products and Coproducts0 The domestic mine production of lead-zinc
continues to come chiefly from ores mined primarily for their lead and zinc
content. Additional lead was derived from ores in which lead and zinc were
comparably valued as coproducts and as a coproduct of ores mined for copper,
gold, silver, zinc, and fluorspar. The complex ores of the Rocky Mountain
area are particularly dependent for economic extraction on the aggregate
value of the lead, zinc, silver, and gold content, and distinction as to the
leading base metal in value is variable. In addition to lead, the principal
metals recovered in processing lead ores and concentrates are antimony,
bismuth, gold, silver, tellurium, zinc, and copper. Appreciable quantities
of sulfur also are recovered as coproducts of lead. The relationship of the
various associated metals and nonmetals to lead production is shown in
Table 42.
90
-------
Figure 12. Location of lead-zinc and zinc (SIC 1031) mines
-------
TABLE 39
U.S. LEAD CAPACITY AND PRODUCTION FOR 1972
(1,000 MT)*
Mine
Smelter
Refinery
Capacity
580
712
712
Production
(recoverable content
561
638
638
of ore)
Source: Reference 2.
* Data published in tons and calculated to metric tons.
TABLE 40
PROJECTED U.S. LEAD PRODUCTION CAPACITY FOR 1972-1975
(1,000 MI)
Production (recoverable content of ore)*
1972 1973 1974 1975
Mine
Smelter
Refinery
580
712
712
589
712
712
589
712
712
589
712
712
Source: Reference 2.
* Data published in tons and calculated to metric tons,
92
-------
TABLE 41
EMPLOYMENT AT ACTIVE LEAD-ZINC MINE AND CONCENTRATOR OPERATIONS
State Employment
California 48
Colorado 989
Idaho 1,669
Illinois 275
Kentucky 40
Missouri 1,389
New Mexico 569
New Jersey 156
New York 501
Utah 286
Washington 92
Wisconsin 87
U.S. total 6,101
EPA regions
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
.
657
'
40
362
569
1,389
1,275
48
1,761
Source: References 3 and 6,
93
-------
TABLE 42
DOMESTIC LEAD BY-PRODUCT AND COPRODUCT RELATIONSHIPS
(1972)
Ore
source
Lead
Lead
Lead
Lead
Lead
Lead
Lead
Lead
Lead
Zinc
Silver
Copper
Fluorine
Gold
By-product
and
coproduct
Bismuth
Antimony
Silver
Zinc
Tellurium
Gold
Copper
Sulfur
Lead
Lead
Lead
Lead
Lead
Lead
Quantity*
(MT)
W
468
226
74,000
W
1.5
11,800
W
490,000
66,000
2,700
900
W
9.9
Percent of
total product
output
100. 0
51.4
20.3
17.2
W
3.4
0.8
W
87.4
11.8
0.5
0.2
W
--
...
W = Withheld to avoid disclosure.
Source: Reference 2.
* Data calculated to metric tons,
Waste Generation and Characterization
Quantities and Types of Waste Generated. Over 14 million metric tons (15
million tons) of wastes were generated by the lead-zinc ore industry in 1974.
These wastes included 1,866,000 MT (2,570,000 tons) of waste rock, and
12,428,000 MT (13,700,000 tons) of tailings.
The amounts of lead-zinc wastes generated in each producing state and EPA
region are given in Table 43. The states which produce the most wastes are
Missouri followed by Idaho, New York, and Colorado. Missouri, Region VII,
produced 42 percent of the total wet wastes.
The ratios of waste rock and concentrator wastes to ore mined are shown in
Table 44.
94
-------
TABLE 43
TOTAL PRODUCTION STATISTICS BY STATE AND EPA REGION FOR
LEAD-ZINC AND ZINC ORES IN 1974 FOB SIC 1031 (METRIC)
EPA
State Region
New Jersey
New York
II
Pennsylvania
Virginia
III
Kentucky
Tennessee
IV
^ Illinois
Ln Wisconsin
V
New Mexico VI
Missouri _ VII-
Colorado
Utah
Montana
VIII
California IX
Idaho
Washington
X
National total
Ore
mined
103 TPY
183
1.306
1,489
544
544
1,088
59
3,184
3,243
294
468
762
157
8,174
992
278
35
1,305
0
1,677
227
1,904
18,122
Waste
rock
103 TPY
0
5
5
0
11
12
0
78
78
5
0
5
0
1,338
37
66
0
103
0
289
36
325
1,866
Concentrator
Cat lings
Dry
103 TPY
123
934
1,057
406
141
547
0
67
67
102
441
543
125
7,309
880
118
24
1,022
145
1.400
213
1.613
12,428
Wet
103 TPY
615
4.670
5.285
2,265
2.091
4,356
.
9.907
9.907
510
2.205
2,715
895
31.079
8,722
608
120
9,450
965
6,614
2.723
9.337
73.989
Concentrator
Lead
cone.
103 TPY
0
227
227
0
4
4
2
0
2
58
15
73
6
639
70
32
7
109
21
151
2
153
1.234
Zinc
COI1C .
103 TPY
59
146
205
54
28
82
2
156
158
61
12
73
26
176
37
14
4
55
14
125
12
137
926
Copper
cone .
103 TPY
0
0
0
0
0
0
0
0
0
0
0
0
0
20
5
0
0
5
0
1
0
1
26
products
Miscellaneous
103 TPY
0
0
0
41
404
445
54
2,961
3,015
73
0
73
0
30
0
114
0
114
2
0
0
0
3,679
Total
103 TPY
59
373
432
95
436
531
58
3,117
3,175
192
27
219
32
865
112
160
11
283
37
277
14
291
5.865
Ratio of dry
weight waste
to ore
0.67
0.72
0.71
0.75
0.28
0.51
--
0.05
0.04
0.4
0.9
0.7
0.8
1.06
0.92
0.66
0.69
0.86
no mine
1.0
1.1
1.02
0.79
Ratio of dry I Total
weight waste dry waste
to product In region
2.1
2.5
2.3 7.43
4.3
Q.35
1.05 3.91
--
0.05
0.05 1.02
0.56
16.3
2.5 3.83
3.9 0.87
10.0 60.49
. 8.2
1.2
2.2
4.0 7.87
3.9 1.02
6.1
17.79
6.66 13.56
2.44 100
I Total
wet waste
In region
7.14
5.89
13.39
3.67
1.21
42.00
12.77
1.30
12.62
100
Source: References 3 and 11.
-------
TABLE 44
RATIO OF TOTAL WASTE ROCK-CONCENTRATOR DRY WASTES AND WASTEWATER TO ORE MINED
FOR LEAD-ZINC AND ZINC ORES FOR SIC 1031 (METRIC)
EPA
State Region
New Jersey
New York
II
Pennsylvania
Virginia
III
Kentucky
Tennessee
IV
Illinois
Wisconsin.
V
New Mexico VI
Missouri "VII
Colorado
Utah
Montana
VIII
California IX
Idaho
Washington
X
National total
Ore
mined
103 TPY
183
1,306
1,489
544
544
1,088
59
1,184
3,243
294
468
762
157
8,174
992
278
35
1,305
0
1,677
227
1,904
18,122
Waste Ratio of total
rock waste rock
103 TPY to ore
0
5
5
0
12
12
0
78
78
5
0
5
0
1,334
37
66
0
103
0
289
36
325
1,866
0.00
0.004
0.003
0.00
0.02
0.01
0.00
0.02
0.02
0.02
0.00
0.007
0.00
0.16
0.04
0.24
0.00
0.08
0.00
0.17
0.13
0.17
0.10
Concentrator
dry waste
103 TPY
123
934
1,057
406
141
547
0
67
67
102
441
543
125
7,309
880
118
24
1,022
145
1,400
213
1,613
12,450
Ratio of dry
concentrator
waste to ore
0.67
0.72
0.71
0.75
0.26
0.50
0.00
0.02
0.02
0.35
0.94
0.71
0.80
0.89
0.89
0.42
0.69
0.78
0.00
0.83
0.94
0.85
0.69
Concentrator
wet waste
103 TPY
615
4_,670
5,285
2,265
2^091
4,356
.
9_,907
9,907
510
2,205
2,715
895
31,079
8,722
608
120
9,450
965
6,614
2.723
9,337
73,989
Ratio of
wet waste
.to ore
3.36
3.58
3.55
4.16
3.84
4.00
NA
3.11
NA
1.73
4.71
3.56
5.70
3.80
8.79
2.19
3.43
7.24
0.00
3.94
12.00
4.90
4.08
Source: References 3 and 11.
-------
The national ratio of waste rock to ore shows that 0.10 ton of waste rock
was mined for every ton of ore. Dry concentrator waste totaled 0.69 ton for
every ton of ore mined, while wet concentrator waste was 4.08 tons. The ratio
of waste rock to ore varied from 0 to 0.24 ton/ton. One reason for this
variation is the age of the mine. The states with the highest ratio of waste
rock to ore are the states where new mines were recently opened. Very little
waste rock is brought to the surface in the older mines because the rock is
used underground for mine roads.
The ratio of dry concentrator wastes to ore mined for lead-zinc ranged from
0.26 ton/ton to 0.94 ton/ton. This variation in ratio is related to ore grade,
percent of lead, and zinc and other metals present in the ore, rather than to
differences in recovery methods and disposal of by-product materials by the
mine. The variations in wet concentrator wastes reflect the difference in the
mines, wet versus dry mines. Some mines must continually pump water out of the
mine to be able to operate, while other mines are relatively dry. All of the
water is pumped to the tailings pond with the concentrator wastes.
Mining Processes. Underground mining systems used in the New Lead Belt of
southeast Missouri and in the Coeur d'Alene region of Idaho are used here as
typical of lead-zinc mining practices.
In Missouri, lead-zinc ore occurs in the Bonne Terre dolomite as irregular
masses of disseminated sulfides which vary greatly in area, shape, and thickness.
Galena is the principal ore mineral, with lesser amounts of sphalerite and
some copper sulfides. The ore-bearing formations are at depths of about 122 m
(400 ft) in the northern part of the district, increasing in depth to about
396 m (1,300 ft) in the southern part. Ore bodies range in thickness from 12
to 30 m (40-100 ft) with the possibility of mineralizations on several horizons
and sections of barren rock between masses of ore. This irregularity of ore
bodies requires a flexible method of mining and the use of the random pillar
arrangement to allow maximum ore recovery while leaving pillars of low grade
ore and waste rock whenever possible.
Considerable groundwater is present throughout the area. Although the amount
of influx varies, all mines must provide large sumps and pumping equipment
with a capacity to handle several thousand gallons per minute.
Southeast Missouri New Lead Belt. All mines in the New Lead Belt district
use the room and pillar method of mining with some differences in equipment
and in methods of mining the ore face. There are about 10 operating mines in
this area.2/
97
-------
Pillar size and spacing is determined by the distribution of ore. Pillars
are generally 6 m (20 ft) wide,and headings have a maximum width of 10.5 m
(34 ft). Some mines have mining levels about 2.3 m (8 ft) apart because of
the manner of deposition of the ore, while the others are developed on single
levels. Since the thickness of the ore varies greatly, several modifications
of the mining system are employed. The maximum height of a single cut is about
7.0 m (23 ft), and the minimum working height is about 3.0 m (10 ft). For ore
exceeding 6.1 m (20 ft) in thickness, the face may be advanced at the bottom
and the high ore broken down by flat hole drilling from the top of the muck
pile or by drilling from a truck-mounted aerial platform. Generally, the method
preferred is to advance a face at the top of the ore and to mine the lower
ore later in one or more benches. Extremely thick ore may be mined by advancing
on two levels, then breaking out the ore separating these levels to form a
single high room. In all cases, completely trackless mining is carried out,
using diesel-powered equipment.
One mine in the southeast Missouri New Lead Belt differs from the others
in that the haulage level is located below the ore body. Mining is carried out
with diesel load-haul-dump equipment carrying broken ore a maximum distance
of 274 m (900 ft) to raises which drop ore to the haulage level. Diesel-powered
trains transport the ore to the shaft bottom. The sublevel is driven first,
followed by a spiral incline at a 10 to 14 percent grade to the top of the
ore body, where stoping commences. In this mine about 2,130 m (7,000 ft) of
development openings were driven before production started, in contrast to
the usual procedure in Missouri mines of starting production with a minimum
of underground development.
Figure 13 is a flow diagram of a typical lead-zinc mine in the southeast
Missouri New Lead Belt.
All entries in the New Lead Belt are by vertical shafts. While almost all
primary entries to the deposits are by conventional shaft sinking techniques,
several bored shafts have been completed including ventilation shafts. A
typical conventional shaft is 6 m (20 ft) in diameter and concrete lined. Some
mines have smaller diameter entries.
In the New Lead Belt mines, some companies have continued to use conventional
drum-type hoisting systems; others have installed friction-type hoists.
Drilling equipment may consist of rubber-tired, self-propelled diesel units
carrying two or three drifters 10 to 11.5 cm (4-5 in.) in diameter capable
of drilling faces up to 10.6 m (35 ft) wide by,9 m (30 ft) high. The newer
methods favor the rotary percussion-type drill over the standard percussion
type.
Dynamite is the standard explosive used for blasting in the district. Ammonium
nitrate has not worked satisfactorily because the ground is too wet.
98
-------
Drill
Ore Body
I
Blast
Ore Body
Load
Hau
i
and
Ore
1.406.0C
Crush and
Grind Ore
Load and Haul
Waste Rock
230.000MTPY
Waste
Rock Pile
I
Concentrator
Source: Reference 11.
Figure 13. Typical underground lead-zinc mine (Missouri Lead Belt)
99
-------
Mines in the New Lead Belt have committed themselves to the load-haul-dump
concept. Principal units employed are the Joy transloader, the Wagner
scoop-trans, Caterpillar front-end loaders, and Emico LHD units. All units are
diesel powered and have rubber tires.
Bored vent shafts are used extensively for ventilation. Since the mines are
not gaseous, the principal need for air is to clear the diesel fumes and the
powder smoke. Air needs range from 2,800 to 11,320 m3 (100,000-400,000 ft3) per
minute. Axivane fans are commonly used to circulate air through flexible
ventilation tubing to the working faces in the mine.
Coeur d'Alene, Idahot District. A flow diagram of mining for lead-zinc ore
in the Coeur d'Alene region is shown in Figure 14. Lead-zinc producers in this
region utilize mining practices which differ from those employed in Missouri.
The entry to the mines is by train in a tunnel (adit) and then by hoist to
the working levels. These companies have developed methods which incorporate
the basic concept of cut-and-fill mining with a standard pillar pattern and
sand backfilling of the stopes. The method has been termed the Bunker Hill
pillar stoping system.
In this method, each 3-x 3-m (10-x 10-ft) pillar is carried vertically from
footwall to hanging wall as mining progresses upward by cutting out one complete
2.43-m (8-ft) floor, filling with mill tailings, then cutting the next floor
above. The only other support in addition to the pillar is an occasional timber
set placed where needed. It is suspected that the pillars crush beneath the
sandfill as successive floors are mined. The support that these pillars afford
on the mining floor is sufficient to carry the load of the hanging wall.
Mining of a new floor begins with a raise-up between a chute and manway-service
raise, as in horizontal cut-and-fill stoping. This is done to a full 2-m (7-ft)
cut height with jackleg drills. Usually raise-ups of 9 to 18 m (30-60 ft) in
length are sufficient to accomplish the objective of establishing the new floor.
In more competent ground, main aisles of up to 45 m (150 ft) in length may be
back-stoped in order to afford more working faces once the stope has been
sandfilled and breasting begins on the new floor.
Stopes ideally are filled to within 0.3 m (1 ft) of the back of the old cut
to allow breasting room. Poor ground conditions, however, often make it necessary
to fill completely to the back of the cut. Care is taken to ensure that the
stope is uniformly filled to the same elevation so that the back of the next
cut will be flat, and the rubber-tired mucker will run on the same elevation.
This requirement necessitates that sandfill outlets be strategically placed
throughout the stope. In large pillar stopes, filling is conducted in two or
three stages. As soon as a large area at one end of a stope cut is mined out,
fill preparation is started.
100
-------
Drill
Ore Body
I
Blast
Ore Body
Load and
Haul Ore
Load and Haul
Waste Rock
36.300MTPY
Tailings Dam
Construction
and Dump
255.400MTPY
Crush and
Grind Ore
I
Concentrator
Sand to Mine Backfill
Source: Reference 11.
Figure 14. Typical underground lead-zinc mine (Coeur d'Alene).
101
-------
Laterals and aisles of the stope cut are driven in line. Rounds, either burn
or breast down, depending on how tightly the stope has been filled, are drilled
3 m (10 ft) wide with jackleg machines. A standard drill pattern, electrical
delay sequence, and explosive loading are used. This incorporates the use of
smooth wall blasting techniques to minimize damage to the pillar and the arch.
A stope crew performs all drilling, blasting, face preparation, and muck
removal. All fill preparation is done by a crew which is responsible for raising
cribbed chutes and manways, building fill fences and traps, burlapping, and
piping for sandfill.
Most raises and chutes are of the cribbed variety with a 1-x 1-m (3-x 3-ft)
inside dimension. All chutes are generally inclined with the dip of the ore
body or down to a 40° minimum.
Forced ventilation is required only in stopes which are not connected from
level-to-level by a raise.
/
Mass Balances of Materials* Mass balance data for representative lead-zinc
mining operations in Idaho and Missouri are shown in Figures 15 and 16.
Description of Individual Waste Streams., The only waste material involved
in lead-zinc ore mining is waste rock.
Identification of Potentially Hazardous Waste. Waste rock from the mining
of lead-zinc ores is not considered to be potentially hazardous since the
waste contains only trace amounts of lead and zinc not exceeding the background
values for elements in the mining district.
Open Pit Mining Processes. There is no open pit mining of lead-zinc ores.
Concentrating Processes.
Flotation Processes. Two basic types of feed ores are processed by flotation
to produce concentrate products: (1) lead-zinc sulfide ores; (2) lead-zinc-
copper sulfide ores.. The concentrating practices and wastes involved in
processing these different feed ores are very similar. Flotation processes are
used for all of them. Descriptions are given in the following sections for
two typical examples of flotation processes, lead-zinc sulfide, and lead-zinc-
copper sulfide ores.
Flotation of Lead-Zinc Sulfide Ores. A typical example of this type of
operation is conducted at a flotation mill which produces lead and zinc
concentrates containing silver. A simplified flow sheet for this flotation
process is shown in Figure 15.
102
-------
TO MINE
AS BACKFILL (g
UNDER&ROUND MINES
DRILL, BLAST: LOAD, t
HAUL TO MILL
QWASTE
ROCK
WASTE ROCK
OPEN DUMP
TAILINGS
POND
© WASTE ROCK
(D MINE WATER
ADDITIVES*
CRUSHING-. GRINDING,
A CLASSIFICATION
ADDITIVES'
WASTE-
WOOD PULP i
TRAMP IKON
LfAD FLOTATION
(ROUG-HERS. CLEANERS.
i REORIND MILL)
Z/A/C CONDITIONER
A ROUG-HER.S
TO POND DIKE
TAILINGS
Z//VC CLEANERS
LEAD CONCENTRATE -
THICKENING; FILTERING,
6. STORAG-E-
21NC CONCENTRATE-
TUICKENING, FILTERING,
A STORAG-E
TO LEAD
SMELTER
TO ZINC
SMELTER.
FLOTATION
NAME
Quicklime
Cooper Sulfatt
Zinc Sulfate
Soda Ash
Methyl liobutyl Carfainoi
Sodium Cyanide
Xonthate-Z-11
Separon
SA 1797 (Z-200)
ADDITIVES'
IB/TON
Of ORE
1.32
0.55
0.22
0.22
0.16
0.15
0.098
0.0003
0.003
G/MT
OF ORE
660.0
275.0
110.0
110.0
60.0
75.0
44.5
0.15
1.5
MATERIALS BALANCE AND COMPOSITIONS
O> TS .O, /-7N
DATA ITEMS
Quantify. TPY
Metric TPY
Quantity. Million GPY
Million Liten-Vr
pH
Auov. % Pb
2n
Ag. oi/'T
Cu
Ai
Cd
Sb
Fe
S
Mn
MgO
AI,O,
Si023
Other
W
WASTE
ROCK
TO DUMP
11.663
10.580
"
WASTE SOCK
TO POND
DIKE
109.703
99.521
MINE
WATER
1548
5859
3 - 5.7
CRUDE
ORE
619.933
562.394
6.18
4.90
3.94
0.06
0.07
0.02
8.20
6.iO
0.63
0.45
12.90
52.00
4.55
©
LEAD
CONC.
37.247
33,790
66.59
4.98
40.45
~"~~®
ZINC
CONC.
47.927
43,479
1.18
54.21
2.83
©/c-s
TOTAL
TAI LINGS -
DRY SOLIDS
534,759
483.125
0.25
0.25
0.18
0.0016
\W
TAILING
LIQUID-
TOTAL
612 - 792
2316 - 2998
1 ® 1
TAILINGS
SOLIDS
TO POND
310.160
281.372
TAILING
SOLIDS TO
MINE BACKFILL
224.599
263.753
NOTE: Company operates !ead and zinc imelten and refineriet on lite.
Source: Reference I I.
Figure 15. Mining and concentrating process:
d'Alene, Idaho district.
lead and zinc ores, Coeur
:„ 1
-------
DRILL, BLAST", LOAD 2
H4UL To MILL.
cauee
oee
™
Q(
MSI
AC
C8VSHIN6-,
i CLASS/r/CAT/Of
ADD/ rives*
I \
sre
Of
c
OUST
SOiLffTOK
f
ptsr
TSffvaess
1
numr-
'
LEAD-COP**
A BECLEANERS '•
_
-------
Crushing, Grinding, and Classification, To reduce the ore size from about
30 cm (12 in.) to 1 cm (0.40 in.), a jaw crusher and cone crushers perform
primary and secondary crushing. Tramp iron is removed by a separator magnet
operating on or under a conveyor belt, and oversize from screening is recycled
to the crushers.
Primary wet grinding is carried out in rod and ball mills operated in closed
circuit with a spiral classifier. Classifier overflow is passed over a vibrating
screen (4-mm (0.16-in.) openings) to remove wood pulp, which is discarded.
The screen undersize is pumped to the lead flotation system.
Lead Flotation. The slurry conditioned by chemical additives (sodium
cyanide, zinc sulfate, xanthate Z-ll, methyl isobutyl carbinol, and soda ash)
is processed in a flotation system consisting of rougher flotation cells and
cleaner cells to produce coarse lead concentrate. Underflow from the roughers
and cleaners is passed through cyclone classifiers to a ball-mill regrind
circuit. The purpose of the regrind circuit is twofold: (1) to promote the
recovery of the lead and zinc from the ore; (2) to reduce the amount of energy
required to liberate the finely disseminated metal sulfides from the gangue
material. From the regrind circuit, the slurry is returned as feed for the
lead rougher flotation cells and is refloated.
After being dewatered by thickening and filtration, the final lead concentrate
containing 66 to 70 percent Pb, 4 to 5 percent Zn, and about 1,370 g/MT of silver
(40 oz/ton), is stored or shipped by railroad cars to a smelter. The classified
fines overflow the hydroclone and are sent to the zinc flotation section of the
plant.
Zinc Flotation. The fines slurry obtained by classification of the lead
rougher underflow is sent to a zinc conditioner and mixed with reagent additives
(xanthate Z-ll, methyl isobutyl carbinol, copper sulfate, quicklime, and
separan). From the conditioner, the slurry is sent to the zinc flotation
machines consisting of rougher flotation cells and cleaner cells. The solids
in the slurry average about 72 percent -200 mesh with 1 percent +48 mesh and
about 57 percent -325 mesh. Tailings from the rougher cells are sent to a sand
plant, and the coarse materials are used for mine backfill. The fines are
disposed of on land in a tailings pond.
The rougher cell middlings are processed in a cyclone classifierj the
coarse fraction (underflow) goes to the ball-mill regrind circuit; and the
fine fraction is returned through the zinc conditioner to the rougher cells
as feed material.
105
-------
Froth from the zinc cleaner cells is dewatered by thickening and filtration
to produce the final zinc concentrate. This concentrate, which contains 54.21
percent Zn, 1.18 percent Pb, and 88 g (2.83 oz) of silver per ton, is stored
or shipped in railroad cars to a zinc smelter or electrolytic zinc refinery.
The depressed material from the zinc cleaner cells is returned to the zinc
rougher cells for more efficient product recovery.
i
Flotation of Lead-Zinc-Copper Sulfide Ores. A simplified flow sheet
representative of this type of flotation process is shown in Figure 16.
Crushing, Grinding, and Classification. The crude ore is crushed by
primary and secondary crushers to -2.54 cm (1 in.) size and then wet ground in
rod mills to produce a 50 percent water slurry containing finely divided solids.
A cyclone classifier separates the over-size which is returned to the rod mill.
A dust collector is provided for the crushing operation, and the material
collected in the cyclone is sent to the flotation circuit for metal recovery.
Flotation Operations. The slurry from the rod mill is sent to a flotation
system (roughers, cleaners, and recleaners) where chemicals are added to produce
a froth to activate the lead and copper minerals which are floated off as froth.
The zinc materials which are not activated and the gangue from the ore flow to
a separate zinc flotation circuit. The froth (containing lead-zinc minerals)
from the lead-zinc flotation circuit is sent to a lead-copper conditioner where
chemical reagents are added and the lead is depressed.
The conditioned froth passes to a copper rougher where the copper minerals
are floated; the depressed lead passes as underflow to a separate lead processing
system which includes thickening, filtering, and drying to yield a concentrate
containing about1 5 percent moisture. The dried lead concentrate, containing
about 73 percent Pb (dry basis), is stored or shipped to a'smelter.
The copper-containing froth from the copper roughing and cleaning'operations
is dewatered in a thickener and a filter. The final copper concentrate containing
10 percent moisture and about 28 percent Cu (dry basis) is stored or shipped to
a smelter.
The copper-rougher underflow containing zinc minerals and gangue is
conditioned and floated in roughers, cleaners, and recleaners to yield a crude
zinc concentrate and tailings. The crude zinc concentrate is then dewatered
in a thickener and filtered to yield a final zinc concentrate containing about
10 percent moisture and 52 percent Zn (dry basis). The final concentrate is
stored or shipped to a smelter.
Mass Balance of Materials. Mass balance data for Example 1, lead-zinc sulfide
ores, and Example 2, lead-zinc-copper sulfide ores, are shown on the table in
Figures 15 and 16. For an ore consumption rate of 562,394 MT- (619,933 tons) per
year (MTPY), the quantities of waste material, callings solids, are 485,125
MTPY (534,758 TPY).
106
-------
Description of Individual Waste Streams. The only waste stream in the
flotation operations is the tailings from the zinc rougher. In the example
lead-zinc flotation plant, the concentrator tailings solids contain 0.25
percent lead, 0.25 percent zinc, and 0.0016 percent cadmium. In the case of
the Pb, Zn, and Cu ores, the tailings could contain 0.2 percent copper. All these
components are potentially hazardous substances and are commonly present in
the tailings from lead-zinc flotation plants. Also, the feed ore may contain
arsenic, a highly toxic material and, therefore, a potentially hazardous
material. However, the arsenic normally floats with the copper and is recovered
at the smelter refinery. No analytical data are available to establish whether
any arsenic is present in the concentrator wastes. Table 45 is analytical data
from the literature on concentrator tailings in the Coeur d'Alene. Table 46
is a compilation of all data reported by the companies contacted in the course
of this program.
Identification of Potentially Hazardous Wastes* Tailings from all lead-zinc
concentrator plants are classified as potentially hazardous wastes which require
disposal safeguards to prevent environmental contamination. The concentrates
from ores containing pyrite are generally considered to be more hazardous,
so special care must be taken in their disposal.
Table 47 shows the total and potentially hazardous wastes from the mining
and concentrating of lead zinc ores. In 1974, there were 11,955,000 MT dry
weight (13,178,000 tons) and 61,817,000 MT wet weight (68,142,000 tons) of
potentially hazardous wastes resulting from concentrating lead-zinc ores.
Tables 48 and 49, respectively, show the 1977 and 1983 projections for total
and potentially hazardous wastes resulting from the mining and concentrating
of lead-zinc ores.
Waste Treatment and Disposal
The waste treatment practices for wastes disposed on land by the lead-zinc
industry are similar to the practices of the underground copper mining industry
(Section I). The mine waste rock is either disposed of in waste rock piles,
used for construction of tailings dams and mine roads, disposed of in the
tailings ponds, or crushed for use as mine backfill. The wastes from concentrating
operations are disposed of in tailings ponds, with some mines using the coarse
fraction for mine backfill.
Mining Waste Treatment and Disposal. There is no generation of potentially
hazardous waste resulting from the mining of lead-zinc ores in the United
States.
107
-------
o
00
TABLE 45
ANALYSIS OF TAILINGS FROM LEAD-ZINC MINES AND CONCENTRATORS (CONCENTRATION IN PPM)
Element
Calcium
Cadmium
Copper
Iron
Potassium
Magnesium
Manganese
Sodium
Lead
Ant imony
Zinc
Concentrator
1
1,549
8.5
71.6;
76,708
335
2 ,460.
5,527
92
2,302
365
2,057
Concentrator
2
3,279
6.3
72.5
59,813
406
2,482
5,932
69
2,624
360
1,316
Concentrator
' 3 ' •-
2,524
19.2
136
88,917
422
3,781
9,123
75
4,380
390
3,547
Concentrator
4
1,598
20.9
117.8
113,667
287
2,680
10,154
69
4,462
1,463
3,366
Background
1,500
—
21
11,800
1,800
3 , 700
490
151
51
—
150
Source: Reference 13.
-------
TABLE 46
ANALYTICAL DATA FOR LEAD-ZINC TAILINGS SOLIDS FOR SIC 1031
(CONCENTRATION IN PPM)
1,700
2,200
7,800
3,140
900
400
3,900
2,000
~ 7
1,500
600
3,590
2,000
6,000
2 ,000
2,100
1,300
380
150
Source
. References 7, 8, and 12.
16
109
-------
TABLE 47
TOTAL AND TOTEKTIALLY HAZARDOUS WASTES FROM MINTNG AND CONCENTRATING
LEAD-ZINC AND 7.TNC ORES FOR 1974 FOR SIC 1031 (METRIC)
OF
Dry waste weight (103 TPY)
Total
process
State Region waste
New Jersey
New York
II
Pennsylvania
Virginia
III
Kentucky
Tennessee
IV
Illinois
Wisconsin
V
New Mexico VI
Missouri VII
Colorado
Utah
Montana
VIII
California IX
Idaho
Washington
X
National total
123
939
1,062
406
153
559
0
145
145
107
441
548
125
8,647
917
184
24
1,125
145
1,689
249
1,938 •
14,294
Total
potentially Waste rock
Concentrator
tailings
hazardous Potentially
waste Total hazardous Total
123
934
1,057
0
141
141
0
0
0
102
441
543
125
7,309
880
118
. 24
1,022
145
1,400
213
1,613
11,955
0
5
5
0
12
12
0
78
78
5
0
5
0
1,338
37
66
0
103
0
289
36
325
1,866
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
123
934
1,057
406
141
547
0
67
67
102
441
543
125
7,309
880
118
24
1,022
145
1,400
213
1,613
12,428
Wet weight (103 TPY)
Concentrator
tailings
Potentially
hazardous Total
123
934
1,057
0
141
141
0
0
0
102
441
543
125
7,309
880
118
24
1,022
145
1,400
213
1,613
11,955
615
4.670
5,285
2,265
2,091
4,356
MA
9,912
9,912
510
2,205
2,715
895
31,079
8,722
608
120
9,450
965
6,614
2.723
9,337
73,994
Potentially
hazardous
615
4.670
5,285 -
0
2,091
2,091
NA
0
0
510
2.205
2,715
895
31,079
8,722
608
120
9,450
965
6,614
2.723
9.337
61.817
Source: References 3 and 12.
-------
TABLE AS
PROJECTED TOTAL AND POTENTIAI.i.Y HAZARDOUS WASTE FROM MINING AND CONCENTRATING
OF LEAD-ZINC AND ZINC ORES FOR 1977 FOR SIC 1031 (METRIC)
Dry waste weight (103 TPY)
EPA
State Region
New Jersey
New York
II
Pennsylvania
Virginia
III
Kentucky
Tennessee
IV
Illinois
Wisconsin
V
New Mexico VI
Missouri VII
Colorado
Utah
Montana
VIII
California IX
Idaho
Washington
X
National total
Total
process
waste
157
1,202
1,359
520
195
715
0
186
186
137
564
701
160
11,068
1,173
235
31
1,439
186
2.162
319
2,481
18,295
Total
potentially
hazardous
waste
157
1,196
1,353
0
180
180
0
0
0
131
564
695
160
9,356
1,126
151
31
1,308
186
1.792
273
2,065
15,303
Waste
rock
Concentrator
tailings
Potentially
Total hazardous Total
0
6
6
0
15
15
0
100
100
6
0
6
0
1,713
47
84
0
131
0
370
46
416
2,387
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
157
1.196
1,353
520
180
700
0
86
86
131
564
695
160
9.356
1,126
151
31
1.308
186
1.792
273
2,065
15,909
Wet weight (103 TPY)
Concentrator
tailings
Potentially
hazardous Total
157
1.196
1,353
0
180
180
0
0
0
131
564
695
160
9,356
1,126
151
31
1,308
186
1,792
273
2,065
15,303
787
5,978
6,765
2,899
2.676
5,575
NA
12,687
12,687
653
2,822
3,475
1,146
39,781
11,164
778
154
12,096
1,235
8,466
3.485
11,951
94,711
Potentially
hazardous
787
5.978
6,765
0
2.676
2,676
NA
0
0
653
2,822
3,475
1,146
39,781
11,164
778
154
12,096
1.235
8,466
3.485
11,951
79,125
Source: References 3 and 12.
-------
TABLE 49
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING
I.EAD-Z1NC AND ZINC ORES FOR 1983 FOR SIC 1031 (METRIC)
NJ
Dry waste weight (10 TPY)
EPA
State Region
New Jersey
New York
II
Pennsylvania
Virginia
III
Kentucky
Tennessee
IV
Illinois
Wisconsin
V
New Mexico VI
Missouri VII
Colorado
Utah
Montana
VIII
California IX
Idaho
Washington
X
National total
Total
process
waste
173
1,324
1,497
572
216
788
0
204
204
151
622
773
176
12,192
1,293
259
34
1,586
204
2,381
351
2,732
20,155
Total
potentially
hazardous
waste
173
1,317
1,490
0
199
199
0
0
0
144
622
766
176
10,306
1,241
166
34
1,441
204
1,974
300
2,274
16,857
Waste
rock
Concentrator
tailings
Potentially
Total hazardous Total
0
7
7
0
17
17
0
110
110
7
0
7
0
1,887
52
93
0
145
0
407
51
458
2,631
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
173
1,317
1,490
572
199
771
0
94
94
144
622
766
176
10,306
1,241
166
34
1,441
204
1,974
300
2,274
17,523
Wet weight (103 TPY)
Concentrator
tailings
Potentially
hazardous Total
173
1,317
1,490
0
199
199
0
0
0
144
622
766
176
10,306
1,241
166
34
1,441
204
1,974
300
2,274
16,857
867
6,585
7,452
3,194
2t948
6,142
NA
13,976
13,976
719
3,109
3,828
1,262
43,821
12,298
857
169
13,524
1,361
9,326
3,839
13.165
104,332
Potentially
hazardous
867
6,585
7,452
0
2.948
2,948
NA
0
0
719
3,109
3,828
1,262
43,821
12,298
857
169
13,524
1,361
9,326
3,839
13.165
87,162
Source: References 3 and 12.
-------
Concentrator Waste Disposal Operation.
Flotation Tailings Disposal. Figures 17, 18, 19 are flow diagrams
describing the tailings disposal methods at three typical lead-zinc
concentrators. The first flotation plant (Figure 17) disposes of tailings '•;
in a tailings pond. Before the plant started up, an earthen dam was built to 'if';',
contain the tailings slurry. Then the land was cleared of vegetation and the |
roots grubbed out. The land was leveled by using a bulldozer and the soil
compacted to present an impervious surface to the tailings slurry which is
pumped to the pond.
A portion (4 percent) of the tailings is sent to a sand plant where the
coarse sand is collected and stockpiled. The fines are discharged to a settling
pond, and the water discharged from the settling pond to a dispersion pond
for final treatment before discharge. When a tailings pond is full, another
tailings pond is put into the system, and the old pond is vegetated with native
trees, shrubs, and grasses.
The second flow diagram (Figure 18) describes a system used in the Coeur
d'Alene region. The flotation tailings are pumped to a sand plant where they
are divided into two fractions. A coarse fraction (42 percent) is pumped back
to the mine, mixed with cement, and used to backfill the rained-out stopes.
The fine fraction (58 percent) is pumped to the tailings pond for disposal.
Waste rock is crushed and used for constructing the dam for the tailings
pond. The tailings slurry is pumped to the pond and discharged about 3 m
(10 ft) from the face of the dam. A water decant system was installed to
recover the water from the pond so it could be pumped to a wastewater treatment
plant. The effluent from the wastewater treatment plant is discharged, by state
permit, into a river, and the sludge from the treatment plant is pumped to the
tailings pond.
The acid mine water is also pumped to the tailings pond; the analysis of
the solids in the mine drainage as well as the analysis of the tailings
are shown in Tables 45 and 46. Pumping of acid mine water to the tailings
pond is considered to be a poor practice; this is recognized by the company,
and they are contemplating building a treatment plant to treat the mine water
before it is pumped into the pond. Neutralization of the mine water before
being added to the tailings pond will help to prevent hazardous materials from
leaching out of the pond and appearing in the stream. The solubility of the
metals is drastically reduced by raising the pH of the tailings slurry to above
8.
For stabilization, the dam at this plant is being revegetated with native
shrubs. There are no plans for abandonment of this tailings pond during the
life of the mine and concentrating plant. As more pond space is needed, the
dam is raised with a waste rock and tailings mixture,,
113
-------
Ore
lOO.OQOMTPY
CONCENTRATOR
(FLOTATION)
Product
31.000MTPY
Mine and
Process Water
470.000MTPY
Tailings
2.650MTPY
LJ
65.750MTPY
TAILINGS POND
Revegetate
Fines
" 540MTPY
Source: Reference 11.
SAND PLANT
Coarse
2.110MTPY
STOCKPILE
Figure 17. Level I technology for waste treatment and disposal, lead-zinc
ores (Coeur d'Alene).
114
-------
Mine Backfill
I Ore
10.000MTPY
1
CONCENTRATOR
(FLOTATION)
Products
14.000MTPY
Tailings
86.000MTPY
SAND PLANT
Waste Rock
l.OOOMTPY
Cement
Coarse
36.100MTPY
1 1 I
49.900MTPY (Dry)
Mine Drainage
and Process Water
336.000MTPY
WASTE ROCK FOR
DAM CONSTRUCTION
TAILINGS POND
Solids
Revegetate
Water
117.600MTPY
WASTE WATER
TREATMENT PLANT
H2O Effluent
Source: Reference 11.
Figure 18. Levels II and III technology for waste treatment and disposal,
lead-zinc ores (Coeur d'Alene).
115
-------
Ore
50,OOOMT
1
Concentrator
1
Products
4,000 MT
Product
900 MT
Tails 46,OOOMT
, Slurry
Cyclone Separator
Source: Reference 11-.
Revegetation
H20
Tailings
Dry
Coarse
•18,400 MT
Earthcore Dam
with Keyway
Bottom Red Clay
Impervious to Water
Tailings Pond
Seepage
Catch Basin
& Dam
Fines
27,600MT
Fines
Figure 19. Level I technology for.waste treatment and disposal, lead-zinc
ores (southeast Missouri).
116
-------
The third flow diagram, Figure 19, shows the procedure for disposal of
flotation tailings in the southeast Missouri New Lead Belt.
The tailings are pumped to a cyclone separator. The coarse material (40
percent of the tailings) is discharged on the dam and used for dam construction.
The fines (60 percent of the tailings) are discharged on the pond side (inside
face) of the dam in the hope that they will seal the dam and prevent seepage.
The dams for these tailings ponds are earth-core dams, keywayed for strength
and stability. The pond is in a valley, and the dam is built across the end
of the valley and tied into the hills. The red clay valley floor in this area
is impervious to water.
The construction of an earthen dam downstream from the tailings pond has
formed a catch basin. Seepage water from the pond is caught in the basin and
pumped back to the tailings pond.
A portion of the dry fines (2.5 percent) is sold to local farmers and used
as agricultural lime.
The stream survey study conducted in this lead belt by the University of
Missouri at Rolla concluded that there was no metal pollution from the tailings
ponds.—' There was a problem with mine water, however, and this was solved.
There is a problem with transportation of concentrates which has not yet been
solved.
Levels I, II, and III Technology. Level I technology for disposal of
flotation tailings in the lead-zinc industry is the use of tailings ponds for
final disposal and the revegetation of these ponds. All of the lead-zinc mining
and concentrating companies use tailings ponds and revegetate the dams for
stabilization, Figures 17 and 19.
Level II technology for disposal of flotation tailings is the practice of
backfilling the stopes in the mines with the coarse tailing in areas where
this is practical. The tailings are pumped to a sand plant and separated into
two fractions. The coarse fraction is pumped to the mine, mixed with cement,
and used to fill mined-out stopes. The fines are pumped to a tailings pond.
Approximately 25 percent of the companies are using Level II technology. An
example of Level II technology is shown in Figure 18.
Level II technology for lead-zinc mines with acid mine water is neutralization
of the acid mine water before discharge to the tailings pond. At present, one
company out of the six which have acid water present is studying the use of
this practice. Level II technology is equivalent to Level III technology, and
an example is shown in Figure 18. About 17 percent of the companies are using
Levels II and III technology.
117
-------
The use of tailings ponds into which no unneutralized acid mine water is
pumped, with revegetation of the pond and dam, is environmentally adequate.
The only risk encountered is when unneutralized acid mine water (pH 2-2.5)
is pumped to the pond, lowering the pH of the pond below 6. The solubility
of some metal salts increases when the pH of the contacting solutions is below
7. Therefore, acid mine water must be treated to raise the pH above 8 before
it is discharged to a tailings pond. The tailings slurry from the lead-zinc
flotation is generally above 10 pH, and at this pH there is practically no
solubility of the metal salts*2ii2/
Lead, zinc, cadmium, copper, and bismuth are contained in the waste. These
metals are potentially hazardous if the pH of the contacting solutions is
below 8. The pH of the contacting solutions affects the potential hazard
presented by the waste. There is no radioactive material in the wastes.
Based on the returned questionnaires, company data, and our engineering
judgment, there were in 1974 approximately 11,900,000 MT dry weight (13,000,000
tons) of potentially hazardous waste resulting from the concentrating of
lead-zinc ores (Table 47).
Future Adequacy of the Technology Identifiedo The adequacy of the Level II
technology treatment methods will not be changed by air and water pollution
enforcement regulations. There are no predicted changes in volume and composition
of waste due to the future imposition of air and water pollution controls.
There is one change that might occur if air pollution controls are imposed
on the crushing and grinding plants. All plants which do not have dust
collectors will probably change to wet grinding rather than install dust
collectors. However, a waste material is not generated in this step. In fact,
collectors and wet grinding are processing steps used to increase the feed
to the concentrator.
The Level II technology identified can be retrofitted to all lead-zinc mines
and concentrating facilities. However, the capital cost and energy requirement
will be considerable where it is necessary to build a mine water treatment
plant. The lead time would vary from 18 months to 2 years.
Energy and Cost Requirements. The energy amounts and cost associated with
the proposed treatment and control technologies have been estimated as a
portion of the t'otal cost necessary to implement the recommended technologies,
and included in the waste treatment and disposal costs discussion which
follows next.
118
-------
Waste Treatment and Disposal Costs - Lead-Zinc Mining
Disposal Cost--Technology Level I« Potentially hazardous waste disposal
practices in lead-zinc mining (Technology Level I) utilize a tailings pond.
A starter dam is built initially and a sand plant is used to raise the dam
as the amount of tailings increases. The sides of the dam are vegetated to
help reduce erosion. The pond used as the representative plant covers 40 hectares
(100 acres). Land costs (assumed to be f12,350/hectare ($5,000/acre)) are a
significant fraction of the total disposal costs. Therefore, three alternative
cases are presented to show how different resale values and methods of
accounting affect the disposal cost per metric ton*
Table 50 presents the major assumptions for technology Level I, Table 51
presents the result for the Coeur d'Alene area. In general, the disposal
costs associated with lead-zinc are substantially higher than the same costs
for uranium or copper mining. The costs range from $0.55 to $1.63/MT of
potentially hazardous waste. Total industry costs, if all companies used the
technology, would be 6.8 to 20.3 million dollars per year. The total cost
would amount to 3.2 percent to 9.5 percent of the value added in mining and
concentrating of lead-zinc ores.
Technology Level I differs between Coeur d'Alene and southeast Missouri.
Mines at the Missouri location usually contain a catch basin or dam, as well
as a tailings pond. The additional dam adds to the cost of construction. In
addition, approximately four more hectares (10 acres) are needed for the catch
basin. These items increase the cost of potentially hazardous waste disposal
in Missouri.
Disposal Cost--Technology Levels II and III. In lead-zinc concentrating,
waste disposal technology Level II is identical to technology Level III.
Both levels utilize a wastewater treatment plant, as well as a tailings pond.
Major cost assumptions of the treatment facility are presented in Table 52.
Capital, operation, maintenance, and material costs are included. Table 53
presents the results for these more advanced technologies. Disposal of
potentially hazardous waste material ranges in cost from $2.22 to $1.55/MT
of waste. The cost to the industry would be 19.2 to 27.6 million dollars
per year, or 8.9 percent to 12.9 percent of value added in mining.
119
-------
TABLE 50
MAJOR COST ASSUMPTION - TAILINGS POND, LEAD-ZINC
TECHNOLOGY LEVEL I (COEUR d'ALENE)
I. Capital
A. Land
Tailings pond assumed to have a maximum area of 40 ha (100 acres)*
Land cost at $12,350/ha (f5,000/acre) = $500,000
Levelized annual cost (assuming no resale value) = $58,500
B. Tailings Pump and Pipeline
Tailings pond assumed to be within 0.8 km (0.5 mile) of concentrator
Tailings pump and pipeline cost = $43,860t
Levelized annual cost at capital recovery factor of 0.117 = |5,100/year
C. Tailings Pond Starter Dam
Starter dam cost = $38,000t
Levelized annual cost = $4,446/year
D. Sand Plant
Sand plant capital cost = $58,000 t
Useful life = 10 years
Levelized a'nnual cost = |9,406/year
II. Labor
Equivalent of one operator for sand plant and tailings pump and pipeline
for 5 months; 8 hr/day at |8.97/hour*, labor cost = $7,176/year
III. Supervision
Supervision cost (25% of labor cost)* = $l,794/year
IV. Maintenance
Assumed to equal 6%/year of nominal cost of equipment*
Maintenance cost = $6,112/year
120
-------
TABLE 50 (Concluded)
V. Insurance and Taxes
Insurance and tax (2% of nominal capital cost; land equipment)* =
fl2,797/year
VI. Energy and Power
Tailings pump and pipeline assumed to cause electricity with a value
of less than $l,000/year§
VII. Vegetative Stabilization
Approximately 4 ha (10 acres) assumed to require stabilization §
Vegetative stabilization costs at $3,245/ha ($l,298/acre)** = $12,980
Levelized annual stabilization costs = $l,519/year
* MRI estimate.
t Sears, M. B., et. al. Correlation of radioactive waste treatment costs and
the environmental impact of waste effluents in the nuclear fuel cycle for use in
establishing as low as practicable guides - Milling of uranium ores. Oak Ridge
National Laboratory, Oak Ridge, Tennessee, ORNL-TM-4903, v.1, May 1975. p. 187-188.
Obtained by ratio of material handled (dry millings and water) to total cost for
pump and pipeline.
^ See corresponding table for copper mining.
§ MRI estimate.
** Ludeke, K. Vegetative stabilization of tailings disposal berms. in:
Mining Congress Journal« January 1973. p. 39.
121
-------
TABLE 51
DISPOSAL COSTS', POTENTIALLY HAZARDOUS WASTE FROM LEAD-ZINC TAILINGS
POND", TECHNOLOGY LEVEL I - COEUR d'ALENE
(IN EQUIVALENT ANNUAL 1973 DOLLARS)
Case "A"* Case "B"t Case "D"*
Capital
Land
Equipment and dam
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Vegetative stabilization
Total
$58,500
18,867
7,176
1,794
6,122
12,797
1,000
1,519
$107,765
$48,906
18,867
7,176
1,794
6,122
12,797
1,000
1,519
$98,171
0
$18,867
7,176
1,794
6,122
377
1,000
. - 1,519
$36,555
Metric ton of potentially
hazardous waste deposited
in tailings pond/year $66,290 $66,290 $66,290
Cost/metric ton of
potentially hazardous
waste $1.63 $1.48 $6.55
Total cost if entire
industry adopted
($106/year) $20.3 -- $6.8
Total cost as a percent
of value added in
mining 9.57» -- 3.2%
* Case "A" assumes all land required is purchased in year 1 and has no resale
value after 20 years.
t Case "B" assumes all land required is purchased in year 1 and resold at
same price after 20 years.
4s Case "D" assumes land used for tailings pond has no opportunity cost to the
company.
122
-------
TABLE 52
ADDITIONAL COST ASSUMPTIONS - LEAD-ZINC MINING
TECHNOLOGY LEVELS II AND III
A. Capital
The plant will handle 117,000 MT of water per year or 321,237 liters/day
(84,862 gal/day).
The plant contains two systems; agitated open tank and sedimentation system.
1. Agitated tank capital cost 6,700 liter capacity (1,800 gal) = $40,000.*
Useful life = 20 years.*
2. Sedimentation system capital cost= $ll,000t (84,862 gal/day). Useful
life = 40 years.
Useful SL depreciation, net capital cost = $10,100.
3. Conveyor transport of solids = $15,000.*
4. Equivalent annual cost (tank, conveyor and sedimentation) = $7,617/year.
B. Operation and Maintenance (includes energy and power)
1. Agitated tank, labor cost = $20,000/year.§ Agitated tank, maintenance
cost $4,000/year.
2. Sedimentation system, labor cost = minimal.**
Sedimentation system, maintenance cost = $l,155/year. **
C. Materials
The kilograms of dolomitic lime required per day.t
Lime cost (at $22/MT)tt approximately equal $100/year.
* Blecker, H., and T. Nichols. Capital and operating costs of pollution control
equipment modules. Vol. II—Data Manual. Environmental Protection Agency EPA-R5-
73-0236. July 1973. p. 10.
t Blecker, H., and T. Nichols. Capital and operating costs of pollution control
equipment modules. Vol. II—Data Manual. Environmental Protection Agency EPA-R5-
73-0236.'July 1973. p. 126-127.
* MRI estimate.
§ Blecker, H., and T. Nichols. Capital and operating costs of pollution control
equipment,modules. Vol. II—Data Manual. Environmental Protection Agency EPA-R5-
73-0236. July 1973. p. 11. (
** Blecker, H., and T. Nichols. Capital and operating costs of pollution control
equipment modules. Vol. II—Data Manual. Environmental Protection Agency EPA-R5-
73-0236. July 1973. p. 127.
tt U.S. Environmental Protection Agency. Development document for effluent
limitation guidelines and new source performance, standards for the major inorganic
produces. EPA-440/l-74-007-a. March L974. p. 275.
123
-------
TABLE 53
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE - LEAD-ZINC MINING, TAILINGS
POND AND WASTEWATER TREATMENT PLANT, TECHNOLOGY LEVELS II AND III
Capital
Wastewater treatment plant
Land
Other equipment and dam
Labor
Supervision
Maintenance
Materials
Insurance and taxes
Energy and power§
Vegetative stabilization
Total
Case "A"*
$ 7,617
58,500
18,867
27,176
6,794
11,267
100
14,117
1,000
1,519
$U6,957
Case "B"t
$ 7,617
48,906
18,867
27,176
6,794
11,267
100
14,117
1,000
1,519
$137,363
Case "D"*
$ 7,6.17
0
18,867
27,176
6,794
11,267
100
14,117
1,000
1,519
$102,674
Metric tons of potentially
hazardous waste 66,290 66,290 66,290
Cost/metric ton of
potentially hazardous
waste $2.22 $2.07 $1.55
Total cost if entire
industry adopted
($!06/year) $27.6 ' -- $19.2
Total cost as a percent
of value added in
mining 12.97, -- 8.9%
* Case "A" assumes all land purchased in year 1 with no resale value after
20 years.
t Case "B" assumes all land purchased in year 1 and resold at the same value
after 20 years. •
$ Case "D" assumes land used for tailings pond has no opportunity cost to the
company.
§ Not including energy and power requirements of wastewater treatment plant.
124
-------
REFERENCES
1. U.S. Department of the Interior. Bureau of Mines. Commodity data, summaries,
1974. Appendix I to Mining and Minerals Policy, p. 88-89.
2. U.S. Department of Interior. Bureau of Mines. Lead (J. P. Ryan). January
1974. Unpublished commodity data report.
3. International Directory of Mining and Mineral Processing Operations.
Published by Engineering & Mining Journal. McGraw-Hill, New York,
New York.
4. U.S. Bureau of Mines. Metals, minerals, and fuels. 1973 Minerals Yearbook.
v.I. p. 685-713.
5. U.S. Bureau of Mines. Metals, minerals, and fuels. 1971 Minerals Yearbook.
v.I. p. 665-693.
6. Mining Enforcement and Safety Administration. Unpublished data.
7. Rausch, D. 0., and B. C. Mariacher. Mining and concentrating of lead and
zinc. v.I of AIME World Symposium on Mining and Metallurgy of Lead
and Zinc. 1970.
8. Wixcon, B. G., and J. C. Jennett. An interdisciplinary investigation of
environmental pollution by lead and other heavy metals from industrial
development in the new lead belt of southwest Missouri. v.I&II, NSF.
Rann. 1974.
9. Weast. R. C. Handbook of chemistry and physics. 49th ed. The Chemical Rubber
Company, Cleveland, Ohio. 1963.
10. Lange, N. A. Handbook of chemistry. 9th ed. Handbook Publishers, Inc.
Sandusky, Ohio, 1956.
11. MRI communication with lead-zinc mining companies.
12. American Mining Congress questionnaires and MRI site visits.
13. Williams, R. E., and A. T. Wallace. The role of mines tailings ponds in
reducing the discharge of heavy metal ions to the environment. U.S.
Bureau of Mines open file report 61(l)-73 PB 224 730, PB 224 731.
125
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SECTION III
ZINC ORES (SIC 1031)
Industry Characterization
Zinc Mining. The metallurgy of zinc began as an alloy material long before
it became known as a metal. The oldest known piece of zinc is in the form of
an idol found in the prehistoric Dacian settlement at Doroseh, Transylvania.
The Romans were well acquainted with brass, an alloy of zinc, as early as 200
B.C. Zinc appears to have been known in India as early as 1000 to 1300 A.D.
and was probably smelted commercially in the 14th Century.
About 1730, the technology of smelting zinc was brought from China to
England; in 1739, a patent was obtained for a distillation method, and a
smelter was erected at Bristol, England, with a mentioned capacity of 180
MT (198 tons) of zinc annually.
In the United States the first zinc was produced in 1835 at the Arsenal
in Washington, D.C. The U.S. Government imported workers from Belgium and
built a small zinc furnace utilizing zincite ore from New Jersey. The primary
purpose of the furnace was to provide zinc to form brass for manufacture of
standard units of weight and measure.
The U.S. zinc industry began in 1860 upon successful operation of a plant
at LaSalle, Illinois, for the treatment of Wisconsin ores and also a plant
at South Bethlehem, Pennsylvania, the same year. The processing technology
of zinc improved rapidly with adoption of horizontal retorts, mechanically
rabbled roasters, successful hydrometallurgy techniques to produce, by
leaching, a zinc sulfate for lithopone, and the American process for
production of zinc oxide directly from ore. In 1895, natural gas was
discovered in Kansas, which led to the building of a number of smelters
in the southwestern United States utilizing natural gas as a fuel.
The commercial introduction of the froth flotation process early in the
20th Century was the most significant technological advance in the zinc
industry. The problems of smelting lead-zinc sulfide ores and concentrates
had previously prevented efficient recovery of the zinc at lead smelters
because the zinc ended up in the slag. The flotation process also made it
possible to recover the zinc in mixed copper-lead-zinc ores.
Preceding page blank
127
-------
The zinc industry is an international basic industry with world-wide
influence in mining, smelting, and trade. Canada is the world's largest
producer with more than double the output of the Soviet Union, followed
by the United States, Other large zinc-producing nations include Australia,
Peru, Mexico, and Japan.
Domestic Production and Capacities. Domestic production of recoverable zinc
was 433,000 MT (477,000 tons) in 1973 compared to 480,000 Ml (529,000 tons)*
in 1968 (Table 54). While Government data are not available for 1974, it is
estimated that zinc ore concentrate production amounted to 1,138,000 MT
(1,254,000 tons) (Table 55). Zinc production in 1974 was reported in 16
states. New Mexico led the nation with an estimated 235,500 MT (260,000
tons) of recovered zinc concentratet product, followed by Missouri, Tennessee,
and New York. The total zinc ore production by year is given in Table 56. The
mine production of recoverable zinc in the United States, by state and year,
is given in Table 57. Table 58 shows the mine production of zinc by EPA
regions for 1973.
TABLE 54
U.S. ZINC PRODUCTION, 1968-1983
(RECOVERABLE CONTENT OF ORE)*
Year
1968
1969
1970
1971
1972
1973
1977
1983
Metric tons
480,313
501,794
484,568
455,907
433,930
432,734
487,000t
581,000t
Source: Reference 1.
* Data reported in tons and calculated to
metric tons,
t Estimated production.
* Recoverable zinc = the amount of pure zinc recovered.
t Recovered zinc concentrate — the amount of concentrate produced before
smelting and refining of the ore.
128
-------
TABLE 55
U.S. ZINC CONCENTRATE PRODUCTION - 1974*
(RECOVERED ZINC CONCENTRATE PRODUCTS)t
State
Metric tons
California
Colorado
Idaho
Illinois
Kentucky
Missouri
Montana
New Mexico
New York
New Jersey
Pennsylvania
Tennessee
Utah
Virginia
Washington
Wisconsin
U.S. total
EPA regions
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
13,753
37,105
125,367
60,963
2,268
176,056
4,442
235,484
59,439
146,446
54,431
156,354
13,608
28,123
11,793
12,156
1,137,788
Metric tons
•M ~
205,885
82,554
158,622
73,119
235,484
176,056
55,155
13,753
137,160
Source: Reference 2.
* Engineering & Mining Journal and unpublished
mining company data.
t Data reported in tons and calculated to metric
tons.
129
-------
TABLE 56
TOTAL ZINC MINE PRODUCTION OF ORE BY YEAR
(ORE MINED)*
Year 1,000 Metric tons
1969 501,785
1970 484,560
1971 455,899
1972 433,922
1973 434,405
Source: Reference 3.
* Data reported in tons and calculated to
metric tons.
130
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TABLE 57
MINE PRODUCTION OF RECOVERABLE ZINC IN THE
UNITED STATES BY STATE
(METRIC TONS)*
State 1969 1970 1971 1972 1973
Arizona
California
Colorado
Idaho
Illinois
Kansas
Kentucky
Maine
Missouri
Montana
Nevada
New Jersey
New Mexico
New York
Oklahoma
Pennsylvania
South Dakota
Tennessee
Utah
Virginia
Washington
Wisconsin
Other States
Total
8,200
3,018
48,729
50,711
12,487
1,723
4,525
6,929
37,284
5,572
853
22,748
22,051
53,277
2,489
29,968
--
112,973
31,662
16,967
8,834
20,775
--
501,785
8,725
3,187
51,431
37,241
15,237
1,075
3,800
8,268
46,013
1,321
115
26,020
15,060
53,140
2,404
26,810
1
107,283
31,468
16,386
10,846
18,718
--
484,560
7,040
2,724
55,502
40,894
11,526
--
4,779
5,307
43,739
327
64
27,194
12,663
57,533
--
24,891
—
108,222
23,315
15,267
5,245
9,656
2
455,899
9,172
1,090
57,879
35,059
10,321
--
1,614
5,279
56,175
10
--
34,560
11,553
55,110
—
16,641
. __
92,280
19,824
15,230
5,881
6,235
_-
433,922
7,644
18
52,924
41,827
4,762
--
247
17,817
74,706
66
--
29,961
11,182
73,894
--
17,106
—
58,215
15,240
15,134
5,786
7,867
__
434,405
Source: Reference 3.
* Data reported in tons and calculated to metric tons.
131
-------
TABLE 58
MINE PRODUCTION OF RECOVERABLE ZINC IN THE
UNITED STATES BY EPA REGION - 1973
Region Metric tons*
I 17,817
II 103,857
III 32,241
IV 58,463
V 12,629
VI 11,183
VII 74,707
VIII 68,231
IX 7,663
X 47,614
Source: Reference 3.
* Data reported in tons and calculated to
metric tons.
In 1972, the 25 leading zinc mines accounted for 89 percent of the domestic
mine production. The five leading mines produced 36 percent; the first 10
together, 56 percent.
The U.S. mine capacity is not sufficient to meet the total requirements
of the nation, and the U.S. dependence on imports for primary supply of
zinc approximates 50 percent. Ores and concentrates are imported from
throughout the world, with Canada supplying 60 percent, Mexico 24 percent,
Peru 8 percent, and other countries, about 8 percent of the imports. The
demand for zinc is expected to increase at 3 percent annually through 1983.
The projected rate of demand would increase U.S. zinc production to 487,000
MT (537,000 tons) of metal in 1977 and 581,000 MT (640,000 tons) in 1983.
Number, Location', Size, and Age of Mines and Concentrators. The domestic
zinc mining industry in 1974 comprised eight mining operations with a total
of eight mines and seven concentrators (Table 59)«i' Zinc mines vary in size
from medium with less than 100 employees to large ,with over 500 employees.
Recent data from the Engineering & Mining Journal and the Mining Enforcement
and Safety Administration indicate that 88 percent of the operations employed
more than 100 workers in 1974.
132
-------
TABLE 59
NUMBER, LOCATION, AND SIZE OF ACTIVE ZINC MINING
AND CONCENTRATING OPERATIONS
Employees Total Total No. Total Mo.
State 50-99
Pennsylvania
Tennessee 1
New Jersey
100-249
1
4
1
250-499 500-999 operations
1
1 6
1
mines
1
6
1
concentrators
1
5
1
Source: References 2 and 5.
The active zinc mines have been in operation for many years, some more than
90 years. The owners and operators have changed, but the mines have stayed
in production. In Tennessee, two companies have recently purchased operating
mines and concentrators from the former operators.
Lead and zinc statistics are combined because it is difficult to separate
lead and zinc mining operations, as they are often mined together. However,
approximately 50 percent of the domestic zinc production is from ores designated
as zinc ores. These are ore deposits found principally in New Jersey, Tennessee,
and Pennsylvania (Figure 12). The domestic mines operated essentially at
capacity during 1972, but smelter production was down (Tables 60 and 61).
Employment. Total estimated employment in the eight active zinc mines in 1974
was approximately 1,800 according to figures issued by Engineering & Mining
Journal, Mining Enforcement and Safety Administration, and from on-site
investigations made by MRI. More than 65 percent of the zinc workers were
employed in Tennessee (Table 59). Employment by region is shown in Table 62.
By-Products and Cooroducts. Zinc production affects, and in turn is affected
by, the demand for a variety of coproducts and by-products. Ores containing
zinc also contain varying amounts of other valuable and recoverable materials
including cadmium, copper, fluorspar, gallium, germanium, gold, indium, lead,
manganese, silver, sulfur, and thallium. Zinc ranges from the major product,
as in the Tennessee, Pennsylvania, and New Jersey deposits, to a coproduct
as in the complex western ores and in the Missouri Lead Belt.
133
-------
TABLE 60
U.S. ZINC CAPACITY AND PRODUCTION, 1972
(1,000 MT)
Capacity
Source: Reference 7.
Production
(recoverable
content, of ore)
Mine
Smelter
494
603
478
541
TABLE 61
U.S. ZINC PRODUCTION CAPACITY, 1972-1974
(1,000 MT)
Production
(recoverable content of ore)
1972 1973 1974
Mine
Smelter
478
633
479
541
492
540
Source: Reference 7.
134
-------
TABLE 62
EMPLOYMENT AT ACTIVE ZINC MINE AND
CONCENTRATOR OPERATIONS IN 1974
State
Employment
Pennsylvania
Tennessee
New Jersey
EPA regions
207
1,328
207
Employment
III
IV
II
207
1,368
207
Source: References 2 and 5.
The major products associated with zinc and recovered at zinc plants in
stack gases, flue dusts, and residues are sulfur, cadmium, germanium, thallium,
indium, and gallium. Zinc by-product and coproduct relationships are identified
in Table 63.
Waste Generation and Characterization
Data on the production statistics by state and region for zinc ores are
shown in Table 64. Data on the ratio of waste rock, overburden, and
concentrator wastes to ore mined by state and region for SIC No. 1031
(zinc ores) are shown in Table 65.
The total national production of zinc ore in 1974 amounted to 3,912,000 MT
(4,312,241 tons). The total national concentrator products for 1974 amounted
to 3,272,000 MT (3,607,000 tons). There were 270,224 MT (297,871 tons) of
zinc concentrate and 3,002,000 MT (3,309,000 tons) of miscellaneous products.
The corresponding generation of waste rock was 79,000 MT/year (87,000 tons/year),
overburden (0), and concentrator waste (tailings 596,000 MT/year (657,000 tons/
year). The national ratio of waste rock to ore was only 0.02, and the ratio
of concentrator waste co ore was 0.15.
135
-------
TABLE 63
ZINC BY-PRODUCT AND COPRODUCT RELATIONSHIPS
Ore
source
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Zinc
Lead
Fluorine
Copper
Silver
Gold
By-product
and
coproduct
Cadmium
Germanium
Indium
Thallium
Gallium
Manganese
Silver
Sulfur
Gold
Calcium
Copper
Stone
Iron
Zinc
Zinc
Zinc
Zinc
Zinc
Total product
Quantity* output
(MT) (%)
1,718
21,800
6.9
1,181
W
5,000
124.1
267,000
0.6
679,000
4,500
W
4,500
51,000
10,800
8,100
1,800
4.5
100.0
100.0
100.0
100.0
W
16.7
12.6
3.0
1.4
0.8
0.4
W
—
10.7
2.2
1.8
0.4
™ »
Source: Reference 4.
W'— Withheld to avoid disclosing individual
company confidential data.
*: Data calculated to metric units.
136
-------
TABI£ 64.
PRODUCTION STATISTICS BY STATE AND EPA HERION FOR
ZINC ORES IN 1974 FOR SIC 1031 (METRIC)
EPA Ore mined
State Region ( 103 TPY)
New Jersey II 183
Tennessee IV 3.184
Pennsylvania III 545
National total 3,912
Source: References 2, 6, and 7.
r-1
U)
-si
Wastes
Waste Over-
rock burden
0 0
79 0
0 0
79 0
(103 TPY)
Concentrator Total Zr.
wasted vaste cone.
125 123 59,439
67 146 156,354
406 406 54,431
596 , 675 270,224
Products (TPY)
Miscellaneous Total
products products
0 59,439
2,961,126 3,117,480
40,823 95,254
3,001,949 3,272,173
Ratio of Total
Total waste Total waste waste
to ore to product (7-1
0.68 2.1 18.28
0.05 0.05 21.64
0.75 4.3 60. 03
0.17 0.205 100.00
Concentrator
wet waste
(I03 TPY)
736
7,915
2,094
10,642
TABLE 65
RATIO OF TOTAL WASTE ROCK--OVERBURDEN--CONCENTRATOR WASTES TO
ORE MINED FOR SIC 1031, ZINC ORE (METRIC)*
State
New Jersey
Pennsylvania
Tennessee
National
UPA
Region
III
III
IV
Ore mined
(103 TPYJ
183
545
3,184
3,912
Ratio of
total waste
Waste rock rock to
(103 TPY) ore
0
0
79 0.025
79 0.020
Concentrator
wet
(10
2
7
10
waste
3 TPY)
736
,094
,915
.642
RHt
wet
to
4.
3.
2.
2.
lo of
vas te
ore
02
84
49
72
Concentrator
dry
waste
(103 TPY)
125
406
67
596
Ratio of
concentre Lor
dry
waste to ore
0.673
0.745
0.021
0.152
•'•• Cofi.plled from data In Table 64,
-------
On a regional basis, the ratio of waste rock plus overburden to ore ranged
from 0.020 to 0.025. The ratio of concentrator waste to ore ranged from 0.02
in Region IV to 0.75 in Region III. No overburden was produced in any of the
operating regions, and Regions II and III also had no waste rock.
The zinc mines are underground units and are all located in the eastern
half of the nation (Tennessee, Pennsylvania, and New Jersey). A discussion
of typical mining and concentrating operations for zinc ores is presented
in the following subsections. These operations are generally similar to
those used for lead-zinc ores.
Mining Processes.
Underground Mining Process. In one typical zinc ore deposit, the zinc
exists principally as zinc sulfide (sphalerite). Other minerals found in
the ore deposit include zinc carbonate, dolomite, and limestone. Cadmium
is also present in this ore.
As shown in Figure 20, the basic mining method consists of drilling,
blasting with dynamite or AN-FO,* loading the broken ore by mobile load-
haul units, hoisting ore and waste rock to the surface by skip hoist,
and transporting ore by dump truck to a crushing plant.
Waste rock removed from the mine is deposited on an open stockpile. In
this example, the rock is periodically removed from the stockpile, crushed
and ground, and sold as a ground rock product 'for road-building uses.
Following the initial startup period for these mines, there is no overburden
waste.
Mass Balance of Materials. Figure 21 contains mass balance data for a
zinc mine and concentrator. It shows that 454,000 MT (500,000 tons) of ore
were mined for 27,216 MT (30,000 tons) of zinc concentrate recovered.
Description of Individual Waste Streams. The only waste stream in domestic
miiiing of zinc ores is waste rock, and it is not considered to be potentially
hazardous. '•"
Concentrator Processes.
Flotation Process. All domestic zinc ores are concentrated by flotation.
A representative flotation operation including available production data is
shown in Figure 21.
AN-FO - Ammonium nitrate-fuel oil blasting agent; it generally contains
5 percent fuel oil.
138
-------
Drill Ore
Body
Blast Ore
Body
1
Load and
Haul Ore
Waste
Rock
Crush and
Grind Ore
I
Stockpiling
and
Grinding
Concentrator
Source: Reference 6.
Figure 20. Mining of zinc ore,
139
-------
1
OlfE
©
C8USHING,
WET GRINDING-, t
CLASSIFICATION
R-23
SODIUM
tPPEK SULFAT*
1 ,
-39 MCSM
OSE
FLCTATIOH F£ED
CONDITIONING-
4
gouG*Ee FLOTATION
,
tifjnenci OM>
SUMP
\
f'HES
TAiLii\&3
ff£C3v££t
POKDS-
CYCLOME
CLASSIFIERS
fW£S
\ \
AGB/CLILTllOAL. MOffrAK POND SAND
_„_.., LIMESTONE SAA/D PPCDVCT
r""(3TH
IUQWD BECYCLEO
TO PROCESS
CLEAN ?& FLOTATION
ZINC
co/vcefurQATE
VACUUM FILTRATION
T »
&£C*CLED TO WET- Z'HC FILTER CAKE
Gff/HDING C/IXUIT TO DA ii-£ipe SIN
MATERIALS BALANCE AND COMPOSITIONS
DATA
Quantity.
Assay. %
/TS /S\_
ITEMS
TPY
Metric TPY
Zn
Cd
Pb
Cu
S
Unspecified
Limestone
Dolomite
CRUDE
ORE
500.000
453.593
4.07
Present
0
0
Present
. NA
Present
Present
\£/
ZINC
.CONCEN- .
TRATE
30,000
27,216
64.00
0.25
0.00
0.00
32.00
3.75
NA
NA
'There ore no waste materials in this plant other than liquid discharged
to railings recovery ponds. All solids are sold as commercial products.
NAME
Copper Sulfate
R-23 (Dow Chemical)
Frother 65*
Sodium Aerofloat*
FLOTATION ADDITIVES*
LBAON,
OF ORE
0.5
0.08
0.02
0.07 .
G/MT
OF ORE
250
40
10
35
Source: Reference 7.
* Produced by American Cyanamid.
Figure 21. Mining and concentrating of zinc,
140
-------
Crude ore delivered from the mine is crushed and wet-ground to -35 mesh.
Crushing is done by a jaw crusher, a gyratory crusher, and secondary cone
crushers. Grinding is carried out by using rod mills in closed circuit with
ball mills. The ground ore is classified by a cyclone, with the oversize
returned to the ball mills, and the undersize (-35 mesh) material sent to
flotation.
Ground ore is conditioned by incorporation of flotation additives, and
then floated in banks of flotation cells and cleaner cells. The quantities
of flotation additives used are shown in Figure 21. Zinc concentrate
discharged by the cleaner cells is vacuum filtered to produce a zinc filter
cake which is stored or shipped by rail to a smelter.
Underflow from the flotation machines is passed to a sump, and dewatered
and classified as to coarseness by cyclones to yield mortar sand and
agricultural limestone products. Slurry from the sump is discharged to
tailings recovery ponds. All of the tailings sand is reclaimed from these
ponds and sold for commercial uses.
As shown in Figure 21, for the example process, 27,216 MT/year (30,000
tons/year) of 64 percent zinc concentrate are produced from 454,000 MT
(500,000 tons) of zinc ore containing 4.07 percent zinc. The product which
is shipped to a smelter also contains 0.25 percent cadmium, a potentially
hazardous material. This cadmium is recovered at the refinery. The by-
products include ground' rock formed from waste rock, agricultural limestone,
mortar sand, and pond sand produced by classifying the flotation tailings.
The concentrator wastes (tailings) consist of finely divided sands (-35
mesh). These wastes contain a small amount' of zinc, along with large amounts
of limestone and dolomite. Since the feed ore contains a cadmium compound,
a potentially hazardous substance, the tailings may also contain trace
amounts of this material. However, no analytical data are available to
establish the presence of cadmium in this waste. Based on the information
available to the investigators, it was concluded that zinc flotation
tailings are not potentially hazardous.
Waste Treatment and Disposal
There are no potentially hazardous wastes resulting from the treatment and
disposal of wastes generated in the mining and concentrating of zinc ores.
This conclusion was reached as a result of the site visits.
141
-------
REFERENCES
1. U.S. Department of the Interior, Bureau of Mines. Commodity data summaries,
zinc. 1974. pp. 188-189.
2. Engineering &.Mining Journal, 1973-1974. International directory of mining
and mineral processing operations. Published by Engineering & Mining
Journal, Mc-Graw Hill, New York, New York.
/
3. 1973 Minerals Yearbook, Metals, minerals, and fuels. U.S. Bureau of Mines,
v.I. pp. 1239-1269.
4. Heindl, R. A. Zinc. Mineral Facts & Problems. Bulletin 650, 1970 Edition.
5. U.S. Department of the Interior, Bureau of Mines, Mining Enforcement and
Safety Administration.
6. MRI communication with zinc mining companies.
7. Unpublished mining company data.
142
-------
SECTION IV
MERCURY ORES (SIC 1092)
Industry Characterization
History of the Industry. The first recorded mention of mercury, also known
as quicksilver, was by the Greek philosopher Aristotle in the fourth century
B.C.', when it was used in religious ceremonies. Until the 16th century
consumption was small and chiefly for medicinal and cosmetic purposes.
Following the introduction in 1557 of the Patio amalgamation process for
recovery of silver, large quantities of mercury were used for amalgamation
purposes in Mexico, Peru, and other countries. The invention of the barometer
in 1643 introduced mercury into scientific research, and in 1720 Fahrenheit
invented the mercury thermometer.
The early history of mercury in the United States is closely associated with
the discovery of gold and the development of gold mining in California. Output
of mercury began shortly before 1850 in the United States with the advent of
the "gold rush" to California in 1849.
Until World War I, the largest and principal use of mercury in the United
States was in the amalgamation process of recovery of gold. Significant quantities
of mercury were used also in fulminate, drugs, and antifouling paint.
In 1944,'production began on the mercury dry-cell battery, and the electrical
apparatus category became the principal use for mercury. The mercury-cell
process to produce caustic soda and chlorine became widespread following World
War II, and a number of such plants have been installed in the United States.
In recent years, the relighting of many streets with mercury lamps, the production
of hearing-aid batteries, and other uses in electrical equipment have contributed
to the high usage in this category.
The mercury industry throughout the world is comparatively small in terms
of quantity produced, value of production, and number of producers. Major
mercury producing countries include the Soviet Union, Spain, Italy, Mexico,
Canada, Peoples Republic of China, and Yugoslavia.
143
-------
Domestic Production and Capacities. In 1972, there were 37 producing mines
in the United States, down from 56 mines in 1971. The number of producing mines
dropped to 24 in 1973, and to two mines in November of 1974. Table 66 gives
the amount of mercury ore treated in the United States by year.
TABLE 66
MERCURY ORE TREATED IN THE UNITED STATES*
Ore treated
Year (metric tons)
1969 392,440
1970 385,109
1971 241,121
1972 74,915t
1973 23,820
Source: Reference 1.
* Revised.
t Excludes mercury produced from old surface ores, dumps, and as a by-
product.
Production of mercury was 251 MT (7,286 flasks) in 1972, down from 942 MT
(27,296 flasks) in 1970 (Table 67). These figures were further-reduced in 1973
when only 76 MT (2,200 flasks) were produced. Although data are not available
for 1974, it is assumed that mercury production would be considerably below
the 1973 production level. These declining trends in the mercury industry are
a direct result of restrictions in the use of mercury in cosmetics, and EPA
standards for plants producing and using mercury and its compounds. With the
decline in the U.S. primary production, the nation will be dependent on
foreign sources for its supply. This dependence increased from 72 percent
of primary consumption in 1972 to almost 100 percent in 1974. Demand for
mercury is expected to increase at an annual rate of less than 1 percent
through 1983.
144
-------
TABLE 67
MERCURY PRODUCTION - 1968-1973
(METRIC TONS)-
Year
1968
1969
1970
1971
1972
1973
1974t
United States
996
1,023
942
617
251
76
69
Rest of
world
7,963
8,957
8,856
9,683
9,391
9,246
Total
8,959
9,980
9,798
10,300
9,642
9,322
--
Source: Reference 2.
* Calculated to metric tons.
t Estimated.
Number. Location, Size, and Age of Mines and Mills. In 1974, there were 10
mercury mines and mills in operation, but all except two of them shut down
during the year (Table 68). Both of these operations are open-pit mines located
in Napa County, California (Figure 22). While production data are not available,
the two mines are very small, with an estimated employment of about 25 workers
for the combined operations. Both mines are relatively new, having been
established since 1967. However, in both cases there were underground mercury
mines located on the site, which operated for a number of years and had been
closed down for at least 10 years.
Employment. The two active mercury mines employed fewer than 50 workers in
1974.
By-Products and Coproducts. Domestic zinc producers recover coproduct mercury
from smelter operations. In 1974, mercury mines produced no by-products of
value in the mining operation.
Waste Generation and Characterization
Waste Generation. There are two main sources of waste from the mining of
mercury ores: overburden, and waste rock. Overburden, composed of soil, gravel,
clay, and rock, is the material covering ore deposits which is stripped during
surface mining operations. Waste rock, usually containing chert, iron pyrite,
and sulfur is the "non-ore" refuse material associated with the ore mined.
145
-------
TABLE 68
NUMBER, LOCATION, AND SIZE OF ACTIVE MERCURY MINING
AND MILLING OPERATIONS (1974)
State
California
U.S. total
EPA regions
Region IX
Employees
(1-19)
2
2
1
2
Total
operations
2
2
2
Total No.
No. of mills of mines
2 2
2 2
2
Source: Reference 3.
146
-------
Figure 22. Location of mercury (SIC 1092) mines for 1974.
-------
Table 69 contains the production data for mercury mining for 1974. The scale
of operation of mercury mines in.the United States ranged from about 5,000 to
12,000 MT (5,512-13,228 tons) of ore mined per year per mine. The two operating
mines accounted for the production of 22,000 MT (24,000 tons) of ore in 1974.
The amount of waste generated from the mining of mercury ore at these mines
in 1974 was 1,416,000 MT (1,560,000 tons), consisting of 908,000 MT (1 million
tons) of waste rock and 487,000 MT (537,000 million tons) of overburden.
Table 70 also contains the data for ratio of waste to ore. The amount of waste
associated with the two mines was very similar, with ratios of waste rock-to-ore
and overburden-to-ore of about 41:1 and 22:1, respectively.
Mining Processes.
Open-Pit Mining.
Description of a Typical Process, A flow sheet for mercury ore open-pit
mining operations is shown in Figure 23. At the time we started our mine visits
and receipt of questionnaires, there were only two operating mercury mines in
the United States. Both of these mines were originally underground mines, with
the initial mining taking place over 100 years ago. The mines have been converted
within the last 10 years to open-pit mines and are less than 80 m (262 ft) deep.
Conventional earth-moving equipment is used at both mines, and the overall
mining methods are very similar. The mines drill and blast to loosen overburden
and waste rock. Ore is loaded by shovel into trucks and hauled to a concentrator
located on the site.
The waste from mercury mining consists of waste rock and overburden. These
materials are generally mixed and used in land reclamation for filling gulleys
near the two operating mines. One facility leaves part of the waste in the
pit where it is being mined. They also stockpile top soil for use as a cover
over the waste dump to aid in revegetation.
Groundwater intrusion is an operating problem in these mines. Mine water
is pumped out of the mines and discharged to nearby streams. Both mines operate
only during the dry season, usually from April to November, and stop mining
when the rainy season begins.
Mass Balance of Materials. The mass balance for mercury mines shows that
1,395,000 MT (1,537,724 tons) of overburden and rock are mined to recover
22,000 MT (24,000 tons) of ore.
Description of Individual Waste Streams. The waste streams from the open-pit
mining of mercury consist of overburden and waste rock. The overburden is
comprised of soil, sandstone, and clay. The waste rock varies widely in
composition, but usually contains chert, and may contain iron pyrite, and
sulfur. In most cases, waste materials contain sufficient plant nutrients to
support vegetation on the surface of the waste piles.
148
-------
TABLE 69
PRODUCTION STATISTICS BY STATE AND EPA REGION, SIC 1092.
FOR MERCURY (METRIC)
State
California
Ratio of
Total Total
Wastes (10' TPY) Products (TPY) waste waste 7-
Ore mined Waste Concentrator Total Primary By-product Hlscel- Total to to Total Metal
Region ( 10 J TPY) rock Overburden wastes waste metal metal laneous products ore product waste ore
IX 22 908 487 21 1.416 69 0 0 69 64 20,000 - Hg
Source: Reference 5.
TABLE 70
RATIO OF TOTAL WASTE ROCK--OVERBURDEN--CONCENTRATOR WASTE TO ORE MINED. SIC 1092.
FOR MERCURY ORES (METRIC)*
Ratio of
Total total waste
Ore waste rock rock to
State Region (103 TRY) (10 TPY) ore
California IX 22 908 41
Ratio of
Overburden total overburden Tailings
( 103 TPY) to ore (103 TPY)
487 22 21
Ratio of
tailings
to ore
0.95
* Compiled from data In Table 69.
-------
OVERBURDEN & WASTE ROCK
OPEN PIT MINE
DRILL, BLAST, LOAD t,
HAUL TO MILL
TO LANDFILL
CRUSHING-
BOAST/MG-
ROTARY FURNACE
(t//VO£K NEGATIVE PKESS)
OFF-
OASES
CALCINED
OK£
EXHAUST GAS
TO STACK
DUST COLLECTOR
(CYCLONE)
MIST REMOVAL
(EXPANSION TANK)
GAS
Hg
WATER
MERCURY CONDENSING
SYSTEM
©
MEPCURV PACKAGING
BURNT ORE BIN
ME&ATIVE PKESS)
LIQ. SFFLUENT
RECYCLED
W. BLEED
i
CALCINE
SOLIDS
LANDFILL
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEM
Quontity.TPY
M.lric TPY
Anoy. Hg%
Ft$2%
F«%
OVERBURDEN
& WASTE
ROCK
1. 538. 000
1 395.000
ORE
24.000
22.000
0.615
1.6
CALCINED
ORE
23.000
21.000
0
0
1
MERCURY
76
69
100
Soorc* : Rrfsrenc* 5.
Figure 23, Mercury mining and concentrating process.
150
-------
Identification of Potentially Hazardous Waste Streams. Due to the small
scale of the mercury mining industry in the United States, which involves
negligible amounts of hazardous material in the mining waste, it is concluded
that no potentially hazardous wastes are generated by the domestic mining of
mercury ore.
In common with all open-pit mining operations, the inactive surface dumps
require protection by a cover of vegetation in order to prevent erosion by
rain and wind.
Concentrator Process.
Description of Process. All mercury presently produced in the United States
is recovered by heating the mercury ore in furnaces and retorts. This roasting
process is an efficient method for distilling the mercury from the ore.
A typical flow diagram for the concentrating of mercury ore is shown in
Figure 23. A discussion of the operating steps is given in the following
subsection.
Crushing. Ore is hauled from the mine to the mill where it is either spread
on the ground for sorting and drying or placed directly in storage bins. The
mercury ore is crushed to about 2.5 cm (1 in.) in size by jaw crushers. One
mine has a screening operation, and over-sized particles are returned to the
crushing operations. The fines are stored in a fine ore bin and then fed
continuously by a feeder into a roasting furnance. One of the mines has no
covered storage, and ore that is mined must be processed within' a day or two.
When the rainy season starts, the entire operation closes down.
Roasting. Mercury ore is fed continuously into a direct-fired rotary
furnace where it is heated to a temperature of about 750 C by combustion gases.
The furnace is oil fired using a Harick burner and is fired countercurrently
with flame impingement on the ore. The mercury vaporizes at about 443 C. The
cinnabar is oxidized into mercury and sulfur dioxide, and these gases are
withdrawn by an exhaust fan along with the combustion gases.
Calcined ore is discharged into a burnt-ore bin where it is held for a
soaking period and then batch-discharged into trucks and hauled to waste piles.
The furnace and burnt-ore bin are kept under negative pressure to prevent
escape of any gases.
Cyclone. The mercury and other exhaust gases are passed through a cyclone
separator where dust is removed.
151
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Condensing System. After dust is removed, the gases are passed through
a series of 40-cm (16-in.) cast iron pipes connected at the top and bottom
with "U" type bends and hoppers. Here the gases are cooled and condensed to
form a liquid. The condenser hoppers are equipped with spouts which extend
into water-immersed mercury collection pots. The water acts as a seal to prevent
the escape of any mercury vapors. Here, mercury is recovered and packaged into
flasks.
Gas Scrubber. The gas leaving the condensing system is processed through
a Venturi scrubber to remove particulates and dissolved S02, and to cool the
gas and condense any remaining mercury vapor so that the gases can be emitted
to the atmosphere. The gas leaving the scrubber passes into an expansion tank
for mist removal. The gas from the expansion tank is emitted to the atmosphere.
The gas leaving the scrubber passes into an expansion tank for mist removal.
The gas from the expansion tank is emitted to the atmosphere.
Hoeing. Sludges from the gas scrubber are processed by rake and hoeing
operations to recover metallic mercury.
Retort. After the metallic mercury is removed from the sludges by hoeing,
the remaining sludges are retorted to recover any mercury still present. The
retorting of the sludges is a batch operation, a very small operation compared
to the continuous furnace roasting operations. The gaseous product of the
retorts is passed through the condensing system, and the calcined solids are
removed.
Mass Balance of Materials. Figure 23 shows a material balance for one mercury
concentrating plant. There are no chemicals added in mercury concentrating.
Description of Individual Waste Streams. The only waste streams from the
concentrating of mercury are the calcined ore from the furnace and retort
operations. -
Identification of Potentially Hazardous Waste. Mercury concentration waste
consists of calcined material that contains no hazardous materials. It is
concluded that no potentially hazardous waste is generated by the domestic
concentrating of mercury ore.
Waste Treatment and Disposal
t
No potentially hazardous wastes result from the treatment and disposal of
wastes generated in the mining and concentrating of mercury ores at the only
two operating facilities. This conclusion was reached as a result of the
site visits.
152
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REFERENCES
1. U.S. Bureau of Mines. Metals, minerals, and fuels. 1973 Minerals Yearbook,
v.I. p. 757-767.
2. U.S. Department of the Interior, U.S. Bureau of Mines. Unpublished
Commodity Data Report, Mercury (V. Anthony Cammarota, Jr.), February
1974.
3. Engineering and mining journal, 1973-1974. International Directory of
Mining and Mineral Processing Operations. Published by Engineering &
Mining Journal. McGraw-Hill, New York, New York.
4. Unpublished company data.
5» MRI communication with mercury mining companies.
153
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SECTION V
URANIUM-RADIUM-VANADIUM ORES (SIC 1094)
1
Industry Characterization
History of the Industry.
Uranium. Uranium and vanadium are frequently found in the same mineral ore,
and radium has always been obtained from uranium ores, in which it is formed
as a radioactive decay product. The uranium-radium-vanadium ores are,
therefore, grouped into one industrial segment.
Uranium was discovered in 1789 by Marten Klaproth in pitchblende from a
mine in Germany.—' The element was first isolated in 1842.
Radioactivity was discovered in 1396, and radium, a daughter of uranium
decay, was discovered by the Curies and Bemont in 1898 in pitchblende from
Czechoslovakia.— In the early 1900's radium became important in medical
therapy; this led to the search for uranium ores as a source of radium«-i'
The first important sources of radium outside Czechoslovakia were the
uranium-vanadium sandstone deposits in western Colorado and eastern Utah
from which about 275,000 MT (303,137 tons) of ore were produced during 1898
to 1923.-' This ore yielded about 200 g (0.44 Ib) of radium, 2,000 MT
(2,205 tons) of vanadium, and a small amount of uranium; most of the uranium
went into the tai lings J^'
In 1936, mining of the uranium-vanadium ores increased markedly because of
increased demand for vanadium.— Carnotite ores were mined from about 1915
until World War II for recovery of their vanadium content. Since World War
II, the ores have been mined for both the uranium and the vanadium values.
Before the discovery of fission in 1939, the known uses for uranium and
its compounds were relatively unimportant. The first plant to treat uranium
ore was built in Czechoslovakia in L906; the product sought was radium, not
uranium. A plant was built in Denver, Colorado, in 1913 to process handpicked
carnotite ore from the Colorado plateau; the radium was recovered in this
plant, and most of the uranium was wasted.—
Uranium did not achieve prominence until the successful development of the
"atom bomb" during World War !!»=' Before this, interest in uranium ore
centered on the radium content, and much of the early work on radium was
done by Madame Curie*—
155
Preceding page blank
-------
The United States remained the major world producer of radium from 1911 to
.2/ From about 1912 until the early 1940rs, the uranium industry was
small, and the uses for carnotite ores were few. The requirements for uranium
were dramatically changed by events during World War II. Military developments
in the early 1940 Ts demanded large quantities of uranium, and ores from the
Belgian Congo, the Republic of South Africa, Australia, and Portugal were
imported. Prior to 1946, annual domestic production of uranium was small and
did not reach 50 percent of procurement until 1953. The purpose of nuclear
developments up to 1945 remained military, even though the peacetime
possibilities of nuclear energy became apparent when the first chain reaction
was confirmed. In 1946, the Atomic Energy Commission was established, and it
was given the full burden of nuclear responsibility formerly held by the
Army's Manhattan District. As an independent civilian agency of the Federal
Government, the Atomic Energy Commission (AEC) in 1946 assumed the
responsibility for the development, use, and control of atomic energy for the
defense and security of the United States and for peaceful applications.—'
Within recent years, a new federal agency, the Energy Research and
Development Administration, has assumed the functions of the AEC.
There has been no domestic production of radium since 1950^' The small
domestic demand was met by imports, withdrawal from dealers' stocks, and
reprocessing. Radium Chemical Company, Inc., New York, was the main dealer
in the United States*^' During 1974, radium was used mostly in therapeutic
treatment of cancer; however, in medical and industrial applications, cheaper
and less hazardous isotopes have continued to replace radium.4' For these
reasons, radium processing is not discussed in this report.
The uranium industry in the United States consists of many privately-owned
firms engaged in exploring, drilling, mining, milling, fabricating, and
chemical processing of uranium.—
Major free world producers of uranium include the United States, Canada,
Republic of South Africa, Niger, France, and Gabon,-'
Vanadium. Vanadium, when discovered in Mexico in 1801, was thought to be
a form of chromium. The metal was not recognized as a new element until 1831.
The vanadium content of the earth's crust has been estimated at about 1,500
g/MT (48 oz/ton)-?-more than copper, lead, or zinc. Vanadium, however, is
widely distributed throughout a great variety of rocks and does not tend to
concentrate and form ore deposits as readily as do the base metals. Recovery
is principally as a coproduct of other metals; rarely are deposits mined for
the vanadium content alone. One notable exception in the United States is
the vanadium mine near Hot Springs, Arkansas.
156
-------
The first commercial use for vanadium developed about 1860 as vanadium
salts for making ink and in coloring fabrics, leather, glass, and pottery.
Its use as an alloying element in steels began about 1905 by the U.S.
automobile industry. However, its beneficial effects in armor plate, cutlery,
tools, and constructional steels were recognized in France and England as
early as 1896. The manufacture of specialty steels has continued to supply
the principal market for vanadium. Vanadium has important catalytic
properties, but this application has remained small. Since 1960, the use of
vanadium in nonferrous alloys for jet engines and airframes has become
increasingly important. Major world producers of vanadium include the
Republic of South Africa, U.S.S.R., the United States, Finland, and Norway.
Domestic Production and Capacities.
Uranium. Production data, by state, for domestic mining of uranium ore
are shown in Table 71. These data show the recoverable contents of U Og
in the mined ore. During the period of 1970 through 1974, New Mexico and
Wyoming were the dominant producing states.
In 1970, the distribution of total mine output was 52.9 percent (6,061 MT
of U30g from underground mines, 44.8 percent (5,133 MT of ^Og) from open
pits, and 2.3 percent (264 MT of U^Og) from other sources (leaching, mine
water treatment, and raffinate)J^'
During 1971, underground mines accounted for 45.3 percent (5,108 MT of
U-jOg) of total mine output, open-pit mines produced 53.3 percent (6,009
MT of U30g), and miscellaneous activities provided the remaining 1.4 percent
(158 MT of U30g) J^/
In 1972, 40 percent (4,768 MT of ^Og) of mining output was from underground
and 2 percent (238
operations,
mines, 58 percent (6,914 MT of ^Og) was from open pits,
MT of UoOo) was' from miscellaneous operationsri2'
During 1973, open-pit mines contributed 62.5 percent (7,423 MT of U-jOg)
of the total mine output, while 36.1 percent (4,287 MT of UoOg) came from
open-pit mines, and miscellaneous operations (recovery of ^Og from mine
waters, heap leaching, solution mining, and raffinate) represented 1.4
percent (166 MT of U-jOg) of the total output^'
Mining production during 1974 amounted to 11,430 MT (12,600 tons) of U30g,—'
underground mining contributed 40 percent (4,664 MT of U-jOg) of total mine
output, open-pit mines accounted for 58 percent (6,762 MT of ^Og), and 2
percent (233 MT of U-jOg) was from miscellaneous operations (heap leaching,
mine water treatment, solution mining, and low-grade stockpiles)*^'
157
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TABLE 71
URANIUM ORE MINING PRODUCTION IN THE UNITED STATES, BY STATE*, t ,
(RECOVERABLE CONTENT U308 METRIC TONS)
Colorado
New Mexico
Utah
Wyoming
Other states
Miscellaneous sources
all states **
19705
1,237
5,249
741
2,878
1,089
264
1971§
1,150
4,792
655
3,168
l,352t
158
1972§
851
4,902
678
3,875
l,376t
238
19?35
871
4,145
880
4,562
l,252t
166
1974
-
4,899tt
-
3,719tt
2,812tt,*
229
Total 11,458 11,275 11,920 11,876 11,659
Average ore grade,
7. U308f 0.202t 0.205t 0.213t O^OS1" 0.2*
* The other states include Colorado, Utah, Washington, and Texas.
t Source: Reference 5. Includes Alaska, South Dakota, Texas, and
Washington, in 1971 and 1972. The other states for 1973 include Alaska,
Texas, and Washington.
$ MRI estimate based on listing of average grade from 1970 to 1973 and
responses to MRI written questionnaires.
§ Source: Reference 4. Data are based on concentrator recovery factors.
tt Source: Reference 6.
t$ Includes uranium recoverable in miscellaneous operations (treatment of
mine waters, solution mining, heap leaching, and raffinate).
158
-------
Domestic U-jOg production is expanding to meet growing uranium fuel demand.
Nuclear power development, delayed by environmental opposition and other
sources, is expected to be expedited because of the energy crisis.—
Salient statistics for domestic production of uranium concentrate from 1969
through 1974 are presented in Table 72. The annual production of ore increased
from 5,356,000 MT (5,904,000 tons) in 1969 to 5,930,000 MT (6,536,700 tons) in
1973; the average grade of ore during this period varied from 0.20 to 0.213
percent U^Og.
Annual domestic production of uranium concentrate (UjOg) during the period
1969 through 1974 ranged from 10,531 to 12,007 MT (11,600-13,240 tons).
Capacity data for domestic concentrator plants which were operated during
the period of 1969. through 1974 are shown in Table 73. During this period,
the number of active concentrator plants varied from 15 to 19, and the total
nominal capacity ranged from 21,591 MT/day (23,800 tons/day) of ore in 1969
to 28,941 MT/day (31,900 tons/day) of ore in 1972. In 1974, the processing
capacity of the concentrator plants ranged from 408 to 6,350 MT (450-7,000
tons) of ore per day.
The utilization of total concentrator production capacity was 69 percent
during 1972 and 1973 and 65 percent in 1974.
As shown in Table 72, the total amount of UoOg concentrate produced
domestically in 1973 was 12,007 MT (13,236 tons) of U30g. Demand for uranium
concentrate is.expected to increase at an average annual rate of 18 percent
through 1980.— MRI estimates that this 18 percent annual increase will
continue from 1980 through 1983. At this rate of increase in demand, U-jOg
production could reach 23,280 MT (25,660 tons) of U30g in 1977 and 62,840
MT (69,270 tons)1 of l^Og in 1983.
The AEC has terminated its role as a lessor of enriched uranium to industry.
All government-owned special nuclear materials (leased from the AEC for use
in facilities licensed under Sections 103 and 104b of the Atomic Energy Act)
were returned to the AEC or converted to private ownership before July 1973.
The present mining and concentrating sectors are dominated by subsidiaries
or affiliates of major mining and petroleum companies. Many operations are
integrated, and the trend is toward further vertical integration into
uranium refining, enrichment, fuels manufacture, and reprocessing of spent
nuclear fuels.
159
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TABLE 72
SALIENT URANIUM CONCENTRATE (U30g) STATISTICS FOR THE UNITED STATES*
(METRIC TONS U^Oa UNLESS OTHERWISE SPECIFIED)
Production
Mine
Ore (thousand MT)
U-jOg content of ore
Average grade of ore, % ^Og
Recoverablet , $
Mill concentrate}
Deliveries of concentrate
Atomic Energy Commission
Quantity
Value (thousands)
Price per pound
Private industrytt
Imports , concentrate
* Principal source (exceptions
t Estimate.
1969
5,356
11,141
0.208
10,768
10,531
5,610
$72,336
$5.85
5,625
1,364
are noted) :
1970
5,737
11,583
0.202
11,059
11,707
2,286
$28,078
$5.59
8,437
603
Reference 5.
1971
:>,696
11,709
0.205
11,122
11,134
-
-
-
11,612
855
1972
5,822
12,398
0.213
11,684
11,703
-
-
-
10,523
2,113
1973
5,930
12,327
0.208
11,703
12,007
-
-
-
10,977
5,085
1974
5,715**
ll,430tt
0.20**
-
10,614tt
.
-
-
-
-
t Receipts at mills; excludes uranium from leaching operations, mine waters, and refinery residues.
§ Includes marketable concentrate from leaching operations.
** MRI estimate based on history of average grade from 1969 to 1973.
tt Source: Reference 6.
-------
TABLE 73
ACTIVE URANIUM ORE CONCENTRATOR PLANTS*
State and company
New Mexico
The Anaconda Company
Kerr - McCee Corporation
United Nuclear - Home stake Partners
United Nuclear - Home stake Mining Company
Texa£
Susquehanna - Western, Inc.
Susquehanna - Western, Inc.
Continental Oil Company - Pioneer Nuclear, Inc.
Subtotal
Colorado
American Metal Climax, Inc.
r''f-rcr Corporation
Union Carbide Corporation
Union Carbide Corporation
South Dakota
Mines Development, Inc.
Utah
Atlas Corporation
.Rio Algom Mines, Ltd.
Wyoming
Federal - American Partners
Petrotomlcs Company
Union Carbide Corporation i
Utah Construction and Mining Company—'
. Utah Construction and Mining Company**/
Federal Resources Corporation - American Nuclear Corporation
Western Nuclear, Inc.
Exxon Company
Concentrator
location
Bluewater
Grants
Grants
Grants
Falls City
Ray Point
Karnes County
Grand Junction
Canon City
Rifle
Ura va n
Edgemont
Moab
La Sal
Gas Hills
Shirley Basin
Gas Hills
Gas Hills
Shirley Basin
Gas Hills
Jeffrey City
EPA Nominal
region 1969
2.7?2
5,443
3.1/5
VI
907
12.247
Powder River Basin
454
363
1,361
590
1,361
» VIII
816
907
907
1,089
1,089
capacity, metric tons
1970
2,722
6,350
3,175
907
907
14,061
408
|l,814
'
590
1,361
862
1,361
907
1,089
1,089
1971
2,722
6,350
3,175
907
907
14,061
408
K.814
1
590
1,361
1,361
907
1,089
1,089
862
1,089
1972t
2,722
6,350
3,175
907
907
1.588
15,649
408
Il,8l4
'
590
1,361
454
1,361
907
1,089
1,089
862
1,089
1.814
of ore
1973*
2,722
6,350
3,175
1.588
13,835
408
1,814
1.361
454
862
1,361
907
1,089
1,089
1,089
1.814
per day*
1974?
2,722
6.350
3,175
1.588
13.835
408
1,814
1.361
454
862
1,361-
907
1,089
1,089
1,089
1.814
Subtotal
Washington
Dawn Mining Company
Total
No. of concentrator plants
8,936 9,481 10,570 12,838 12,248 12,248
408
454
454
454
454
454
21,591 23,996 25,085 28.941 26,537 26.537
15 15 16 19 16 16
* Sources: References 5. 6, 8, 9, 10. 12, 13, and 14.
t Utilization of total concentrator capacity in 1972 was 697. (Reference 8).
* Utilization of total concentrator capacity in 1973 was 697. (Reference 9).
$ Utilization of total concentrator capacity In 1974 was 65% (Reference 6).
** The company name in 1974 was changed to Utah International, Inc.
ft The Petrotomlcs concentrator in Shirley Basin, Wyoming, was shut down in 1974, but is Included in the table since It Is part of a oajor production
center and can be reopened.
-------
Phosphate rock containing U^Og is a potential source of uranium. In the
fertilizer industry, the rock is reacted with sulfuric acid to form phosphoric
acid, and most of the uranium present goes into the product acid. One of the
recovery processes owned by Uranium Recovery Corporation makes a concentrated
strip solution that is transported to a central processing plant in Florida JJa'
Initial operation of conversion plants for production of UoOg from the solution
is expected in 1976.—'
Vanadium. The principal domestic sources of vanadium are the Colorado Plateau
uranium-vanadium ores, the Arkansas vanadium ore, and the phosphate rock in
Colorado. These include Union Carbide's mine at Wilson Springs, Arkansas, and
Union Carbide's uranium-vanadium properties on the Colorado Plateau. The Wilson
Springs, Arkansas, operation is the only one producing vanadium from vanadium
ore on a large scale. Table 74 gives the mine production of vanadium produced
in the United States by year. In 1973, the production of vanadium ore from this
mine was 454,919 MT (501,462 tons). U.S. vanadium production in 1974 is
estimated at 8,763 MT (9,660 tons) (recovered vanadium concentrate products)
(Table 75).
New sources of vanadium being considered for commercialization include
evaluation of a property in southeastern Idaho containing about 20 million
metric tons (22 million tons) of ore assaying 1,35 percent V20^.
U.S. consumption of vanadium for all end uses probably reached 6,300 MT
(6,930 tons) in 1973, about equal to the 1969 record. The sheer volume of
steel production was responsible for the high level of vanadium use; 80 to
85 percent of all vanadium produced is consumed by the steel industry. U.S.
demand for vanadium is expected to increase at an annual rate of about 5
percent through 1983. Although domestic reserves are large enough to meet
expected demand, cheaper foreign material will probably meet about 25 percent
of the demand.
The problem of oversupply of coproduct vanadium from uranium production
no longer exists and will probably not recur. The United States can no
longer depend on coproduct vanadium from uranium and phosphate mining for
an adequate supply. An increase in supply from these sources will be limited
by the demand for uranium and phosphorus. To maintain a domestic production
capability in vanadium, it will be necessary to (1) discover and develop
other deposits such as those at Wilson Springs, Arkansas; (b) exploit the
vanadium deposits of the Colorado Plateau and the phosphatic shales of Idaho
and Wyoming with little or no recoverable uranium or phosphate; (c) recover
vanadium from titaniferous magnetic operations.
162
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TABLE 74
MINE PBDDUCTION OF VANADIUM PSDDUCED
IN THE UNITED STATES
(MT)
Year Mine production*
1969 5,205
1970 5,255
1971 5,032
1972 4,263
1973 3,735
Source: Reference 5.
* Measured by receipts of uranium and
vanadium ores and concentrates at mills,
vanadium content.
163
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TABLE 75
U.S. VANADIUM CONCENTRATE PRODUCTION, 1974*
(RECOVERED VANADIUM CONCENTRATE PRODUCTS)
State • MT
Arkansas
Colorado
U.S. total
EPA regions ; MT
Region I
Region II
Region III
Region IV
Region V
Region VI 4,536
Region VII
Region VIII 4,227
Region IX —
Region X
Source: Reference 15.
* Engineering & Mining Journal and unpublished
mining company data.
164
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Number. Location. Size, and Age of Mines and Concentrators,
Uranium. Data for uranium mining operations conducted during the period
of 1969 through 1974 are presented in Table 76.
The total number of mining properties in the United States declined from
310 in 1969 to 175 in 1973. From 1970 to 1973, the share of total production
from open-pit mines increased from 44.8 to 62.5 percent, while the
percentage output from underground mines decreased from 52.9 to 36.1 percent.
The output from miscellaneous sources, which included recovery of UoOg from
mine water, heap leaching, solution leaching, and raffinate, ranged from 1.4
to 2.3 percent of total output.
Estimates for 1074 include 122 underground mines, 33 open-pit mines, and
20 miscellaneous output sources. The majority of the mines are located in
New Mexico, Colorado, and Wyoming, with small concentrations in Texas, Utah,
and Washington. Data Cor the number, location, and size of concentrator
plants are shown in Table 73.
Uranium mines vary in size from very small operations employing fewer
than 20 workers to some that employ more than 500 workers. Most uranium
mines are recent developments; practically all have opened since 1950, and
most of the active mines in 1974 have opened since 1960. Nearly all of the
concentrator plants were placed in operation after 1955; many of the plants
were built in the 1960's and the 1970's.
Vanadium. The only mine producing vanadium from vanadium ore on a large
scale is located in Wilson Springs, Arkansas. The mine has a production
capacity of 762,000 MT/year (839,961 tons/year) and employs about 177
workers.
Employment.
Uranium. Total employment in active uranium mines was 5,202 workers in
1974. Employment was concentrated (81 percent of all uranium mine workers)
in New Mexico, Wyoming, and Colorado (Table 77).
Data concerning the total employment for uranium mines and concentrators
are presented in Table 78.
Vanadium. In 1974, approximately 177 workers were employed at the only
vanadium mine operating in the United States at Wilson Springs, Arkansas.
165
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TABLE 76
DOMESTIC URANIUM MINE OPERATIONS AND PRODUCTION DATA*
Operating
year
1969
1970
1971
1972
1973
1974
* Sources
Total
producing
properties
310
263
240
190
175
175*'
: References
Underground mines
No. ?c
-
216
193
141
122
122^
5, 6, 10,
, of Output
-
52.9
45.3
40.0
36.1
40.0
12, 13, and
Open-pit mines
No:
-
27
29
37
33
33b/
14.
% of Output
-
44.8
53.3
58.0
62.5
58.0
Other output sources
No. 7» of Output
20 2.3
18 1.4
12 2.0
20 1.4
20^ 2.0
Estimated by MRI to be the same as for 1973.
-------
TABLE 77
EMPLOYMENT IN ACTIVE URANIUM MINE OPERATIONS IN 1974*
State Employment
Colorado
New Mexico
Texas
Utah
Washington
Wyoming
U.S. total 5,202
EPA regions
Region I
Region II
Region III
Region IV
Region V
Region VI 2,608
Region VII
Region VIII 2,344
Region IX
Region X 250
* Source: Engineering & Mining Journal, 1973-1974. Mining Enforce-
ment and Safety Administration. Unpublished mining company data.
167
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TABLE 78
TOTAL EMPLOYMENT AT DOMESTIC URANIUM MINES AND CONCENTRATORS^
1969 1970 1971 1972 1973 1974
Employees, for mines 9,059 8,165 7,373 6,403 6,500 6,000*
and concentrators
at year end
* Estimate by MRI based on data in Reference 15.
By-Product and Coproduct Relationships. A large part of the U.S. vanadium
output is a coproduct of uranium at two mills on the Colorado Plateau.
Scandium has been recovered in the United States as a by-product of uranium
processing, but has not been recovered from domestic residues in recent years.
Small quantities of molybdenum occur in the uraniferous lignites of the
Dakotas. Copper, zinc, gold, and silver may be recovered during uranium
processing.
168
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Waste Characterization
Uranium-Vanadium. Production statistics by state and region for uranium-
radium-vanadium ores in 1974 are shown in Table 79. The total ore mined
amounted to 5,715,000 MT (6,300,000 tons). The estimated total solid wastes
from mining included 93,6 million metric tons of overburden (103 million tons)
and 2.2 million metric tons (2.5 million tons) of waste rock. The estimated
total dry weight of concentrator wastes amounted to the same quantity as the
ore mined and the total wastewater discharged to disposal in tailings ponds
was 11.7 million metric tons (12.9 million tons). On a national basis, the
ratio of dry weight of total waste to ore was 18 to 1, and the weight ratio
of total dry waste to total concentrator products was 7,401 to 1. Region VIII
accounted for about 75 percent (75.7 million metric tons) of the total dry
wastes and Region VI generated 23 percent (23.9 million metric tons); the
remainder of the dry wastes (2 percent (2 million metric tons)) was contributed
by Region X.
Data on the ratio of total waste rock, overburden and concentrator waste
to ore mined are shown in Table 80. On a national basis in 1974, the weight
ratio of waste rock to ore was 0.39, the ratio of overburden to ore was 16.38,
and the ratio of concentrator waste (dry solids) to ore was 1.0.
Descriptions of representative domestic processes, mass balances of materials,
individual waste 'streams, and potentially hazardous wastes generated in
uranium ore mining and concentrating operations are given in the following
subsections.
Open-Pit Mining Waste. Open-pit mines can be operated in conjunction with
any type of uranium ore concentrating process. Open-pit mine operations of
Wyoming are typical.
Description of Representative Processes. Open-pit, or surface mining,
methods are usually applied where the uranium ore deposits are located near
the surface and are covered with loosely consolidated soil. In other cases,
the choice of method depends basically on the relative mining costs for the
required output. The factors which must be evaluated before determining the
mining method are the size, shape, grade, depth, and thickness of the ore
deposits.i£/
Some mining companies use open-pit methods to depths of more than 152 m
(500 ft), but as the depth of overburden above the ore zone increases beyond
91 m (300 ft), underground mining methods usually are more feasible.—' The
equipment used in surface mining varies widely. For mining small deposits to
depths of 9 m (about 30 ft), conventional construction-type equipment such as
bulldozers, front-end loaders, trucks, gasoline, and diesel-powered shovels
are used..i2/ For deeper deposits, much larger earth moving equipment is utilized,
169
-------
TABLE Ti
TOTAL PRODUCTION STATISTICS BY STATE AND EPA REGION FOR SIC 1094 (URANIUM-VANADIUM ORES)
IN 1974 (METRIC)
State
New Mexico
Texas
Total
Colorado
Utah
Wyoming
Total
Washington
National
Source :
t~4
•xl
O
Ore
gpA mined
Renlon (103 TPV)
2,449
367
VI 2,816
514
420
1,860
VIII 2,794
X 105
5,715
Reference 36.
Waste
rock
(103 TPY)
1,127
484
1,611
0
38
353
391
221
2,223
Over-
burden
(103 TPY)
12,172
7.263
19.435
0
332
72,224
72,556
1,644
93.635
wastes
dry weight
(103 TPY)
2,449
367
2,816
514
420
1,860
?,794
105
5,715
wastes
wet weight
(103 TPY)
6.930
2,276
9,206
1,092
932
5,889
7.913
263
17,382
Conci
Uranium
concentrate
5,442
706
6,148
987
807
3,649
5,443
203
11,794
Vanadium
concentrate
0
0
0
lv?87
0
0
1,787
0
1,787
Other
metals
39
0
19
i04
0
0
104
0
143
Total
5,481
706
6 , 18?
?.B78
807
3,649
7,334
203
13,724
Ratio of dry
weight waste
to ore
6.43
JJ.l)
8.47
1.00
1.88
40.02
27.11
18.76
17.77
Ratio of dry Z Total
weight waste dry waste
to products in region
2,873
11,493
3,857 23.49
179
979
20,399
10,327 74.57
9,704 1.94
7,401 100.00
Z Total
wet waste
In region
52.96
45.53
1.51
100.00
-------
TABLE 80
RATIO OF TOTAL WASTE ROCK—OVERBURDEN—CONCENTRATOR WASTE TO ORE MINED
FOR SIC 1094 (URANIUM-VANADIUM ORES) FOR 1974 (METRIC)*
State
New Mexico
Texas
Total
Colorado
Utah
Wyoming
Total
Washington
National
EPA Ore
Region (103 TPY)
2,449
367
VI 2,816
514
420
1.860
VIII 2,794
X 105
5,715
Waste
rock
(103 TPY)
1,1.27
484
1,611
0
38
353
391
221
2,223
Ratio of
total waste
rock to
ore
0.46
1.32
0.57
0
0.09
0.19
0.14
2 . 10
0.39
Over-
burden
(103 TPY)
12,175
7.263
19,43i"
0
332
72.224
72,556
1,644
93,635
Ratio of
total O.B.
to ore
4.97
19.79
6.90
0
0,79
38.83.
25.97
15.66
16.38
Concentrate
waste .
(103 TFY)
2,449
367
2,816
S14
420
1.860
2,794
105
5,715
Ratio of total
concentrator
waste to ore
1.00
1.00
1.00
1.00
1.00
1.00
1.00
1.00
1.00
* Compiled Crom data in Table 79.
-------
Since control of ore grade is important, each mine develops a method best
suited for its particular ore body. Ore grade is determined by radiation
detection equipment. Most mines resort to ore stockpiling and blending}
chemical analysis is used to control the blended material.^./
Stripping of the overburden to expose the top layer of the deposit may be
a large part of the cost of mining^' The depth of overburden varies from a
few meters to 30 m (100 ft) or moreJLZ' The weight ratio of overburden to
uranium ore removed in the mining operations is high compared to other types
of mining; the ratios can range from about 0.8:1 to 39:1, as shown in Table 80.
Since uranium mining is carried out principally in low rainfall regions,
relatively few mines discharge any water. Wherever feasible, mine water is
used as process water in the ore concentrating operations. In these cases,
it becomes a component of concentrator effluent. A detailed description of
the water balance for domestic uranium mines and concentrators-1 is given in
the technical literature.i§/
A representative flow diagram for open-pit mining operations is shown in
Figure 24, p. 185. Generally, the mining method consists of drilling, blasting,
loading, the broken ore by power shovel into dump trucks, and transporting the
ore to a concentrator.
The stripped overburden and waste rock are usually hauled out of the pit
and land-disposed on open surface dumps located near the mine. At some mines
a portion of the overburden and waste rock is used in mine backfill operations.
Mined ore is checked for grade and hauled to a collection area which serves to
supply feed ore to a concentrator plant.
Mass Balances of Materials. Mass balance data for a representative model
of an open-pit mining operation are shown in Figure 24. The values for mining
wastes were calculated from data shown in Table 79, assuming that all mines in
Wyoming, Texas, and Washington are open-pit mines and that one plant (The
Anaconda Company) in New Mexico operates an open-pit mine. The weighted
average weight ratios (overburden to ore and waste rock to ore) for these
selecte^ data were calculated and reported for the model mine. The assumed
production rate for the model plant was suggested by the literature*^'
The data (.dry weights) in Figure 24 apply for a model open-pit mine which
theoretically produces 662,200 MT (730,000 tons) per year of uranium ore (0.20
percent l^jOg). The corresponding annual quantity of overburden mined is
19,418,000 MT (21,405,000 tons); the weight ratio of overburden to ore is
29 to 1. The annual amount of waste rock mined is 425,700 MT (469,300 tons) and
the weight ratio of waste rock to ore is 0.6 to 1.
172
-------
Industry data presented previously for uranium industry characterization
show that in 1974, open-pit mines accounted for about 58 percent of total mine
production in terms of 11303 values. The estimated ore production from open-pit
mines in 1974 is 3,152,000 MX (3,470,000 tons). The data for dry wastes in
Table 79 indicate that a total of 93,635,000 MT (103,215,000 tons) of overburden
was generated in 1974; all of this waste is produced by open-pit mining. The
estimated total quantity of waste rock mined by open-pit operations during
1974 is 2,026,000 MT (2,233,000 tons). The estimated values are based on the
assumptions and data described in the two preceding paragraphs and the data
in Table 79.
Only very meager data were obtained concerning the moisture contents of the
open-pit mined ore; the moisture content ranges from about 6 to 10 percent.—'
No information was obtained concerning the moisture content of overburden and
waste rock.
Description of Individual Waste Streams. The waste streams from open-pit
uranium mines are overburden material and waste rock. The overburden generally
consists of particles of soil, sandstone, clay, and shale. The waste rock can
vary widely in composition, but usually contains chert, feldspar, quartz, and
small amounts of radioactive minerals (e.g., natural uranium and its decay
products). Usually these waste materials contain sufficient plant nutrients
to support some vegetation on the surface of the waste piles.
Identification of Potentially Hazardous Waste Materials. The only
potentially hazardous waste generated in open-pit mining is the waste rock,
which contains above background levels of radioactive materials (natural
uranium and its decay products). All waste rock is considered to be potentially
hazardous. The decay products of uranium include several potentially hazardous
radioactive substances such as radium, radon gas, and thorium.ii/ The total
potentially hazardous waste generated by domestic open-pit mines in 1974
consisted of about 2,026,000 MT (2,233,000 tons) of waste rock.
Underground Mining Wastes. Underground mines can be used with any type of
uranium ore concentrating process. Underground mines of New Mexico are typical.
Descriptions of Representative Processes. Since the uranium ore deposits
vary widely in shape, size, altitude, and grade, underground mines use a
variety of mining methods .is/
173
-------
In the Colorado Plateau area many mines are simple adits or inclined
entries made in a canyon wall or sloping ground to recover small deposits of
ore.iS.' The small ore bodies are usually mined by open-cast methods, which
may be unsupported or supported by roof bolting, and in wide areas pillar
supports may be used. In the small mines transportation equipment varies from
hand-tramming to mechanical haulage.!^./
The larger and deeper ore deposits require more complex mining methods.^'
The mining techniques include the room-and-piliar, long wall retreat, and panel
methods, along with modifications of each of these procedures. The haulageways
accommodate either electric locomotives on tracks or large, trackless haulage
equipment .!£/
A representative flow diagram for underground mining is shown in Figure
25, p. 191. Generally, the mining method consists of drilling, blasting,
moving broken ore to the surface and transporting the ore to a concentrator.
Mine ventilation is a problem, because of the presence of radon gas and
its radiation hazard.—The decay products of uranium contained in the ore
and waste rock include radon gas and radon daughters. Bodily harm (e.g., lung
cancer), can occur from inhalation of radon gas and radon daughter products
which can attach themselves to dust particles in the air.±i' Significant
improvement has been achieved in the mine environment through increased mine
ventilation and better distribution of fresh air to all working areas.—
Fans are used to circulate the air, and large diameter vent shafts are
installed to provide sufficient fresh air for the mine.—'
Groundwater pumped from the mine to the surface is commonly used as
processing water in the mill. This water frequently contains significant
amounts of dissolved uranium values.
Some companies also conduct special mining operations such as solution
mining, heap leaching, and treatment of mine water to recover the uranium
values. These special mine production activities accounted for about 2 percent
of the total mine output of U30g values in 1974.
174
-------
Mass Balances of Materials* Mass balance data which apply for a
representative model of an underground mining operation are shown in Figure 25.
These values were calculated on the basis of data shown in Table 79, assuming
all mining operations in Colorado, Utah and all but one plant (Anaconda
Company) in New Mexico are typical of underground mining. The weighted
average ratio of waste rock to ore for these selected data was calculated
and reported for the model mine. The assumed production rate was suggested by
the literature.—' These data (dry weights) apply for a model mine which
conceptually produces 662,200 MT (730,000 tons) per year of crude ore (0.20
percent UjOg), The corresponding annual amount of waste rock mined is 51,000
MT (56,200 tons) per year. The weight ratio of waste rock ore is 0.077 to 1
for the model mine; for individual operating mines the ratio can range from
0 to 1 to about 0.1 to 1.
Data indicate that in 1974, underground mines contributed about 40 percent
of total mine output in terms of U-jOg values. Mining production statistics by
state and region for uranium-vanadium ores in 1974 are given in Table 79. For
the operations during 1974, the estimated total ore production by underground
mines was 2,563,000 MT (2,825,000 tons). The corresponding estimated total
quantity of waste rock mined from underground mines was 197,000 MT (217,000
tons). These estimates are based on data in the preceding paragraph and the
data in Table 79.
Only very limited data were found concerning the moisture contents of ore
recovered from underground mines; the moisture content varies from about 8 to
20 percent JP-/
Description of Individual Waste Streams0 The amount of solid waste (waste
rock) generated in underground mining operations is small in proportion to ore
mines. In several mining operations the waste rock is stored in surface piles
for future use as a low-grade uranium ore. The waste rock contains small
amounts of natural uranium and its radioactive decay products.
Identification of Potentially Hazardous Waste Materials. All of the waste
rock mined in underground mining operations is potentially hazardous because
it contains above background levels of radioactive substances such as uranium,
radium, and thorium.
In 1974, the total potentially hazardous waste generated by domestic
underground mines was about 197,000 MT (217,000 tons) of waste rock.
175
-------
Concentrating Operation Wastes. In the uranium ore concentrators, the
uranium values are recovered from the crude ore and concentrated to yield an
1 Q / J
intermediate, semirefined product containing 11303 or Na2U207*i^' This product,
which is commonly referred to in the industry as yellowcake, is shipped to
refineries and reprocessed to obtain either uranium metal, UO?, or UFg for use
in the nuclear industry. This study is concerned only with the processes and
wastes involved in mining and concentrating the ore to produce yellowcake.
Basic Technology of the Ore Concentrating Industry. The basic steps
required in all uranium ore processing consist of: (1) ore preparation; (2)
ore concentration; (3) yellowcake recovery*iS.' Ore preparation includes
crushing and grinding and possibly drying or roasting to improve handling or
solubility properties. Concentration is carried out by hydrometallurgical
leaching methods, using either dilute sulfuric acid or alkaline carbonate
solutions.!^./ The uranium concentrate is recovered from solution by chemical
precipitation, dewatered, ground, and packaged for shipment.
The uranium ore dressing industry can be divided into two basic types of
concentrating operations which are characterized by the leaching process used.
* Concentrators which use acid leaching or combined acid-and-alkaline
leaching operations.
* Concentrators which use alkaline leaching only.
Acid leaching is used for ores containing less than 12 percent lime, and
alkaline leaching is especially useful for the high-lime ores that would
consume excessive amounts of
Concentration and purification of the uranium values from dilute and impure
leach solutions are required for production of a final high-grade, uranium
concentrate product. — The two principal techniques utilized for concentration
and purification are solvent extraction and resin ion exchange, used either
individually or In a two-stage series. Solvent extraction is used to treat only
clarified acid solutions, while the ion exchange process is applicable to the
treatment of both ore pulps and clarified solutions in either acid o.r alkaline
solution. Alkaline leaching operations are frequently sufficiently selective
to permit elimination of the purification step — '
L76
-------
The uranium ore concentrating industry is highly diversified. The processes
used vary among the domestic plants due largely to differences in chemical
composition of the ore, the cost and availability of leaching chemicals, and
the presence of other metal constituents to be recovered as by-products. A
tabulation of the producing companies and an identification of the concentrator
operations used bv the active domestic ore concentrator plants during 1974, are
shown in Table 81.
The data in Table 81 show that the daily feed ore processing capacity of the
concentrator plants ranged from about 408 MT (450 tons) to 6,350 MT (7,000 tons)
in 1974. There were 16 active concentrator plants in 1974. The ores processed
commercially in the United States generally contain 0.1 to 0.5 percent 11303;
the average concentration is about 0.2 percent.—'
An analysis of the data in Table 81 indicates that the percentage distribution
of types of concentrator plants active in 1974 was that shown in Table 82.
The data presented in Table 82 show that 12 of the 16 plants used an acid
leach process and represented 79.6 percent of total concentrator capacity.
Two concentrator plants utilized an alkaline leach process and accounted for
13.7 percent of production capacity, while another two plants used a
combination acid-and-alkaline leach process and contributed 6.7 percent of
the total concentrator capacity.
The acid leach facilities active in 1974 can be further subdivided, as
shown in Table 82, into four plants using solvent extraction (41.8 percent of
total concentrator capacity), three plants using solvent extraction plus ion
exchange (11,5 percent of total concentrator capacity), and five plants using
ion exchange (26.3 percent of total concentrator capacity). Thus, the entire
industry is represented by either acid leach or alkaline leach processes with
a number of variations in the concentrating and purification steps in the acid
leach processes. The acid leach process dominated the industry in 1974 and
represented about 80 percent of the total concentrator capacity.
The 16 active concentrator plants in 1974 were sited in six western states.
Three plants were located in New Mexico, seven plants in Wyoming, two in
Colorado, two in Utah, and single plants in Texas and Washington. The combined
operations in the States of Wyoming and New Mexico accounted for 77.1 percent
of total concentrator capacity (20,460 MT of ore per day).
The data base for the uranium ore mining and concentrating industry
collected by Midwest Research Institute (MRI), by site visits and written
questionnaires, did not represent a sufficiently wide cross-section to make
it truly representative of the industry category. Therefore, in preparing a
characterization of representative concentrator processes and their waste
generation, it was necessary to use technical data available in the current
published literature concerning model plants for processing uranium ores.—'
177
-------
TABLE 8)
ACTIVE U.S. URANIUM ORE CONCENTRATOR P1JWT!; IN 197A
Company t/
EPA region
Anaconda Co./How Mexico/VI
United Nuclear -Homos take Partners/
New Mexicn/Vl
Kerr-McCee Corporation/New Mexico/VI
Conoco-Ploneer H-Jclear/Tex«s/VI
Cotter Corp./Colorafio/VIIt
Union Carbide Corp. /Colorado/Yin
A: las -..;rp.;;'La::/v:;.i
Rio Algoni/Utah/VIII
Union Carbide Corp./Wyptnlng/VIII
Utah International, Inc./Wyomlng/VIII
Cfls Pills
UtiVi International, Tnc ./Wyoming/VIII
Shirley Basin
F«i;rf.'t crutch Co. /Wyoming/ VIII
Dawn Mining Co./Hashington/X
f Data taken from References 6 and
i Data taken from References 19 and
Da tit
started*
19.58
1958
197',
1958
1936
t'j'jS
1972
1960
1958
1972
1962
toe 7
173 /
Vlt'i
1968
19.
20.
Nominal Average
capacity Type ot r>'re* Of-
metric tons Imder- Open- grade T.
of ore/day* ground pit U30g'
i,722 X 0.3'i
3,175 X 0.21
6,350 X
1,588 X
408 X
1,814 X 0.15
1,361 X
454 X
907 X
1,089 X 0.20
1,089 X
1,361 X
i ORP y y
1 , UO7 A A
I,bl4 X
86? X
454 X 0.298
26,537
5 Data taken from data sources 19 and 20. The nomenclature for abbreviations listed
AcH
Alkal'
Ac Id
Acid
Aclri
Alkaline
Acid
ActJ
Alkaline
Alkaline
Acid
Acid
Acid
Acid
Acid
Ac Id
Acid
Uquld-
•.oUd
•-en
:;t
ceo
DF
DF
CCD
COD
DF
CCD
CCD
CCD
CCD
CCD
under table* headings
Operations used?
Yellow Overall
KIP MgO 87
N.-.0-1. 95 ,„
H2S04.
Amlne Nllj > 97
Amlne M1I3 97
** 80 95-96
tl'iOM. Dual process
H2S04,
NH}
IX Nil, 00 95-96 Vr.
A.iilnett
RIP NH-j Dual process
NnOH ,
B2S04.
H202
RIP H113 > 90
IX Amtne Yes NH3 90 95
IX K!l3 or
HRO
Amlne NH3 97 97
Amlne NHj 95
IX
are: CCD - Countercurrent Decantatlon RIP - Resin In P«ilp
" DF - Drum Filters IX - Column Ion Exchange
** In the dual process, utilizing both acid and alkaline leach steps, solvent extract Ion Is conducted using an alkyl phosphoric acid extractant, dl(2-ethylhe>t}'l) phosphoric acid.
hft loaded strip solution is combined with the filtered alkaline leach solution from the alkaline concentrator process for precipitation.
ft In this dual process, the loaded strip solution is combined with the loaded strip solution from the alkaline RIP circuit for precipitation.
** A combination Ion exchange—solvent extraction process (sulfuric acid elutlon of resin followed by solvent extraction).
-------
TABLE 82
DISTRIBUTION OF URANIUM ORE CONCENTRATOR PLANTS OPERATED IN 1974 BY TYPE
OF PROCESS AND PERCENT OF TOTAL FEED ORE CAPACITY*
% of Total
Number concentrator capacity
Plants using acid leach process
Plants using acid leach and alkaline
leach processes (dual process)
Plants using alkaline leach process
(with no purification step)
Total
12
2
2
16
79.6
6.7
13.7
100.0
Acid leach plants:
Plants using solvent extraction 4 41.8
Plants using solvent extraction
plus ion exchange 3 11.5
Plants using ion exchange _5 26.3
Total 12 79.6
* Compiled from data in Table 81.
179
-------
Descriptions of Representative Processes. Detailed descriptions of the acid
leach process and the alkaline leach process for concentrating uranium ore are
presented in the following subsections.
Acid Leach Process. A model acid leach concentrator plant described in
the technical literature was selected as being representative of this type of
process.12/ An analysis of 1974 data (Tables 81 and 82) shows that there are
two major subcategories for this process: acid leach solvent extraction,
representing about 53 percent of total acid leach concentrator capacity, and
acid leach-ion exchange accounting for about 33 percent of total plant capacity.
A third subcategory is a dual process combining acid and alkaline leach
systems. The two major categories taken together, represent most (86 percent)
of the acid leach processes.
Description of Representative Process. A detailed discussion of a
model plant reported in the technical literaturai^ for the acid leach solvent
extraction process is given in this section. The use of this type of concentrator
plant appears to be the trend of the future since four out of five of the new
acid leach plants constructed in the western world in the last 5 years were of
this typeJ:2' Also, many of the processing variables in acid leach plants
affecting the radiological waste are the same for both solvent extraction and
ion exchange systems.—' The common processing items are:
* The system for ore size reduction and chemical leaching.
* The quantity and composition of the solid wastes generated.
* The quantity of radionuclides in the liquid effluent wastes.
* The treatment and disposal methods used for solid, liquid, and
gaseous wastes.
* The quantity and composition of airborne effluents.
Therefore, it is considered that the acid leach solvent extraction
process also basically represents the operations and waste generation for
acid-leach with ion exchange.
A representative flow diagram for the acid leach solvent extraction
process is shown in Figure 24. Uranium ore from either an open-pit mine or
an underground mine can be used in this process. An open-pit mine is shown
as an example in this study.
180
-------
OPEN PIT MIHC
DRILL, BLAST, LOAO 4
HAUL ORE TO MILL
OVERBURDEN®
L WASTE ROCK®
i
MINE WATCH
useo in
P8OCC3S
CRUSHING
WCT G-RINDIHQ-
DECANTAT/OM
(CCD)
AMINE.
KCffGSeH
ALCOHOL
,,(*>
TAILING-3-
SAND, SU*1E
IJQUID WASTES TO
TAILIN&S POKP
sot.ve/vr
exTVAcrtoN
SArFINATC
TO LfACHIMtr
STRIPPING-
SULFUHIC,
ACID
f SODIUM
*
LC ACHING-
AMMONIA
PRECIPITATION
FILTRATION
DQYINO-
YCLL.OW CAKE
PACKAGING-
r
PRODUCT
MATERIALS BALANCE AND COMPOSITION
DATA ITEMS
Annuol Quantity. Metric TPY
Partial Analyiti;
1)303. vVt.%
Radium. >>Ci/g UjOg
Thorium. ^ Ci/g UjOg
U f P^' ''9
226 Ra. pCi/g
230Th. pCi/g
234 Th. pCi/g
Un.LpCi/lir.r
226 Ra. pG/ liter
230TK pCi/liter
Calcium (Co )g/li!er
Iron (Fe)s/l'"r
Aluminum (Al)g/ liter
Ammonia (NHjJg/liler
Sodium ( No) g/ liter
Anenic (Aj) g/liter
Fluoride (F) g/liter
Vonodium (V) g/liter
Sulfot. (SO42-) g/liter
Chloride (CI-) g/liter
Total Oiuolved Solid,.
PH
-~~^L>
OVER-
BURDEN
19,418.000
"\>A~~ "
WASTE
ROCK
425.700
-~^ji
C«JD6
ORE
(DRY)
662.200
0.20
TAILINGS
DRY
SOLIDS
667.200
0.018
52
567
567
52
LIQUID
993. 300
6.700
500
1«,000
0.5
1 0
20
0 5
0 2
0.0002
0.005
J.0001
30
0 3
35.0
2
"~~\£J
YELLOW
CAKE
PRODUCT
1.339
90
5.5K 10-*
1.4x ID"2
j
1
PROCESS ADDITIVES 1
NAME
Sulfuric Acid
Sodium Chlorate
Ammonia
Amine*
Alcohol'
Kerosene
Flocculonr
Total
IB/TON
OF ORE
PROCESSED
73
2.2
2.3
0.035
0.06
0.64
0.16
78.40
G/MT '
OF ORE '
PROCESSED !
36. 500 :
1.100 '
1.150 !
17.5 j
30.0 !
320.0
80.0 i
39. 198 j
Alamin* 336 and Adog«n '364 ore uwd int«rchang«ibly
liodecanot and Trid«conol or* used interchangeably.
Source: Reference 36.
Figure 24. Mining and concentrating of uranium ore—acid leach process.
-------
The model concentrator was selected to have a daily capacity of
1,814 MT (2,000 tons) of feed ore containing 0.20 percent U30g. The
concentrator is assumed to operate continuously for 364 days/year.ii/ A
discussion of the process used in the model concentrator follows.
Ore delivered by dump trucks is passed through a bar screen (grizzly)
to a primary crusher which reduces the ore to -1.25 cm (1/2 in.) size.
Oversize material is recycled to the crusher. The crushed ore is elevated and
moved on a rubber belt conveyor to storage bins which are vented through a
dust collection unit (orifice dust collector) to the atmosphere. Air exhaust
hoods are located on the crusher, at the screens, and at each ore transfer
point. The air is passed through a dust separator unit (orifice dust
collector) and then discharged to the atmosphere through a roof vent.—-'
Collected dust is sent to processing; detailed data on dust collection equipment
and dust quantities are given in Reference 19.
Crushed ore, drawn from storage bins, is wet-ground in ball mills as
mixture containing about 65 percent solids (by weight)._' Ore ground to pass
a 28-mesh screen is discharged into a. chemical leaching system consisting of
several mechanically agitated tanks in series with a total residence time of
about 7 hr. In some other plants, air agitation is employed in tall vessels
(called Pachucas) as a substitute leaching system. Sulfuric acid and an oxidant,
sodium chlorate, are added continuously to the slurry discharged from the ball
mills. To obtain the desired high leaching efficiencies, it is necessary to
oxidize any tetravalent uranium present in the ore to the hexavalent state.
Sodium chlorate (NaC103) is a typical oxidant. Manganese dioxide is also a
suitable oxidizer. A typical leach reaction is:
3 H2S04 -fr NaC103 + 3 U02 ^ 3 U02S04 + NaCl + 3 H20
The ore is leached continuously at atmospheric pressure and at slightly above
room temperature.. The pH of the leach slurry is usually maintained between
005 and 2.0.—' In this step, most of the uranium values in the ground ore
are dissolved along with other elements which may be contained in the ore such
as some radionuclides, iron, aluminum, and many other impurities. The ore is
ground prior to leaching to permit a much faster and more efficient extraction
of uranium because of the greatly increased surface area in the ground ore
particles.
182
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Following the leaching step, the leach liquor (containing undissolved
suspended solids) is sent to a countercurrent decantation (CCD) thickener
operation.12/ Flocculant is added to the thickener to promote settling of
solids. The CCD system serves to effect a separation of the uranium-bearing
liquid from the undissolved solids. Water is added to wash uranium solution
from undissolved solids8 Underflow (slurry) from the CCD system is passed
through liquid cyclone separators (hydroclones) to separate the coarse sand
fraction, and the sand is washed in a series of classifiers. The overflow
from the classifiers is combined with the hydroclone overflow, and the
resulting materials (slimes) are washed with fresh water and recycled raffinate
in a series of thickeners using flocculant to promote settling. The washed
slimes and sands are pumped as a tailings slurry to a tailings pond. The
weight ratio of dry solids in the tailings slurry to ore is 1 to Ij the
composition of the dry solids is 70 percent sands and 30 percent slimes. The
total weight of waste solution accompanying the sands and slimes to the tailings
pond is 1.5 times the weight of the ore processea»iz/
Leach solution discharged by the countercurrent decantation system is
sent to a solvent extraction (SX) step. Uranium is recovered from the leach
liquor by countercurrent contact in several extraction stages in series (e.g.,
four stages) with a long-chain amine dissolved in kerosene*iZ/
This process is based on the property of the organic solvent, which is
immiscible in water, to form a complex with the uranium. This complex is
soluble in an excess of the solvent.—' When aqueous leach solution containing
solubilized uranium compounds is contacted with the organic solvent, the
uranium is distributed between the aqueous and organic phases. Under properly
controlled process conditions, the uranium is extracted almost quantitatively
into the organic phase, while most of the other constituents (impurities) of
the leach liquor remain in the aqueous phase. After mixing to provide adequate
contact, the mixture of solutions is removed to a settling tank, where the
lighter, uranium-loaded layer rises to the surface and is separated by
decantation. These operations are conducted on a continuous basis in a series
of mixer and settler tanks. Aqueous raffinate from the solvent extraction
operation is recycled to the countercurrent decantation step.!2'
The uranium loaded organic solvent discharged from the solvent
extraction is treated in a stripping operation to separate the uranium from
the organic solvent.—' The uranium is stripped from the solvent in several
stages (e.g., four stages) with an aqueous solution of ammonium sulfate. The
recovered organic solvent is recycled back to the solvent extraction step.—'
Uranium loaded aqueous solution from the stripping operation is
treated by chemical precipitation to separate uranium values from the
The uranium is precipitated by addition of gaseous anmonia to the strip
solution, concentrated, and partially washed in thickenersJt2'
183
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The uranium-bearing aqueous slurry discharged from the precipitation
step is collected and washed on filters.!!/ The wet uranium-bearing solids
discharged from the filters are dried in a continuous steam-heated dryer. The
dried uranium precipitate (yellowcake) is packaged in 55-gal steel drums for
shipment to a refinery. The yellowcake product contains about 90 percent U308,
and the overall recovery of uranium as product is 91 percent of that contained
in the feed ore.!i/ The literature indicates that the thorium content of the
yellowcake is about 1.4 x 10"2 u.Ci/g l^Og (i.e., 5 percent of the total thorium
in the ore), and that the radium content is 5.5 x 10 p.Ci/g l^Og (0.2 percent
of total radium in the ore).—' No other significant radionuclides are contained
in the product.
The air streams from the l^Og precipitate dryer and the hoods over the
packaging area are combined and passed through a dust collector (wet impingement
dust collector). !27 Collected dust is recycled to the process.
Any liquid spillage or leakage throughout the concentrator plant is
collected in floor sumps and then returned to the appropriate process step.
The only liquid waste stream is that leaving with the sands and slimes to the
tailings area.12'
Mass Balances of Materials* Data for a mass balance of materials
involved in the acid leach concentrating process are shown in Figure 24. These
data apply for a model plant (not an existing facility) reported in the current
technical literature.—' This model is considered to be representative of all
domestic acid leach processes*!!/
The data for this model show that on an annual basis, the consumption
of 662,200 MT (729,950 tons) of feed ore corresponds to an identical quantity
of tailing solids, 1,339 MT (1,476 tons) of product, and 993,300 MT (1,094,900
tons) of waste tailing liquid.
Process additives used include sulfuric acid, sodium chlorate, ammonia,
aiuine, alcohol, kerosene, and flocculant. The amine, alcohol, and kerosene are
recovered and reused in the process, while the remaining materials and their
reaction products are distributed among the waste materials. The consumption
of additives amounts to 39,200 g/MT of feed ore (78.4 Ib/ton).
The composition of the dry waste solids is 30 percent slimes and 70
percent sands (by weight).!!/ The slimes fraction contains 85 percent of the
insoluble radioactive materials. The uranium distribution (originating with the
feed ore) is 91 percent in the yellowcake product, 7.6 percent in the slimes,
and 1.4 percent in the sands.!!/ Fifty percent of the thorium dissolves during
concentrating and ultimately crystallizes in the slimes fraction, while 50
percent is insoluble in the tailing.!!/ Other radionuclides are insoluble.
184
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Total production statistics for uranium-vanadium ores in 1974 are shown
in Table 79. The estimated total quantity of concentrator dry tailings wastes
were 5,715,000 MT (6,299,700 tons). Data in Table 82 show that 79.6 percent of
the total domestic concentrator capacity was represented by concentrator plants
using an acid leach process. The amount of solid waste generated is directly
proportional to the quantity of feed ore consumed regardless of the type of
concentrating process. Therefore, assuming an equal utilization of processing
capacity by each concentrator plant, the quantity of land-disposed solid wastes
from acid leach processes can be estimated as 79.6 percent of 5,715,000 or
4,549,000 MT (5,014,400 tons) in 1974. Based on wastewater data from written
questionnaires and estimates by MRI, the total wastewater discharged to
tailings ponds (including some direct discharge of mine wastewaters) in 1974
for all acid leach processes was 10,201,000 MT (11,245,000 tons).
The other land-disposed wastes for the industry in 1974 were attributable
to the operation of dual concentrator processes using both an acid leach and an
alkaline leach. In 1974, there were two active dual processes which accounted
for 6.7 percent of the total feed ore consumed. The estimated quantity of solid
wastes from dual processes during that year is: 6.7 percent of 5,715,000 or
382,900 MT (422,100 tons); the estimated amount of liquid wastes accompanying
the solid wastes to land disposal is 688,000 MT (758,000 tons).
Description of Individual Waste Streams. The only concentrator waste
which is land-disposed is the tailings slurry containing sands, slimes, and
liquid. Airborne effluents containing ore dust, radon gas or yellowcake dust
are discharged to the atmosphere from various operations in the concentrator
plant, such as ore crushing and screening and product drying.i2'
Data on the composition of waste generated in a model acid leach-solvent
extraction plant are presented in Figure 24. In this model plant, the dry solids
fraction of the tailings contains 0.018 percent of U30g, 52 pCi/g of natural
uranium, 567 pCi/g each of radium-226 and of thorium-230, and 52 pCi/g of thorium-
234.
As shown in Figure 24, the liquid fraction of the tailings slurry (for
the model plant) contains various metals dissolved from the uranium ore along
with chemical additives (or reaction products) introduced in the concentrator
process.
Identification of Potentially Hazardous Waste Materials. The entire
concentrator tailings, which are land-disposed, are potentially hazardous
because they contain radioactive materials in concentrations above the
background level.
185
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Alkaline Leach Process. A model alkaline concentrator plant reported
in the technical literature was selected as representative of this type process.~'
Description of Representative Process. A representative flow for the
alkaline leach process is shown in Figure 25. Uranium ore from either an open-pit
mine or an underground mine can be used with this process. An underground mine
is shown as an example in the study.
The model concentrator plant has a daily capacity of 1,814 MT (2,000
tons) of feed ore containing 0.20 percent U^Og.12/ The concentrator is assumed
to operate continuously for 364 days/year.JL2/ A discussion of the process used
in the model concentrator follows.
The operations involved for ore receiving, conveying crushing and ore
storage facilities are the same as those described previously for the acid leach
mill including dust separation from air streams^' Therefore, the description
of these activities is not repeated here.
The crushed ore is ground to provide a finely divided ore suitable for
efficient leaching in subsequent processing. The wet-grounding system consists
of a ball mill operated in closed circuit with a classiferJL2' The grinding
is carried out at a solids content of 65 percent (by weight) in a sodium
carbonate-bicarbonate solution. The ore discharged from the system is finely
ground and the average particle size is about 95 percent -48 mesh, and 35 percent
-200 mesh.
The ground ore is leached in a two-stage system consisting of a 5-hr
leach at a pressure of 65 psig (in autoclaves) and 200 F, followed by an 18-hr
leach at atmospheric pressure and 185 F.i2' The alkaline leaching system
requires the oxidation of any tetravalent uranium contained in the ore to the
hexavalent state using oxygen available in air supplied to the leaching
vessels. Hexavalent uranium dissolves in the presence of carbonate alkalinity
to form a uranyl tricarbonate complex ion. Leaching is carried out in an
aerated solution containing sodium carbonate and sodium bicarbonateJ£/
Undissolved solids are separated and washed free of uranium by three
stages in series of countercurrent filtration (i.e., a sequence of filtering
and repulping operations). The solids (filter cake), which are separated from
the final filtration step, consist solely of equal parts (by weight) of sands
and slimes; these solid wastes are repulped with fresh water and pumped to a
tailings pond. The weight ratio of solid wastes to ore is 1 to 1. The weight
of waste solution sent to the tailings ponds is 1.05 times the weight of the
feed ore processed.—'
186
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nOCCULANT
UNDERGROUND MINE
DRILL. BLAST, LOAD t
HAUL ate TO MILL
o
WASTE
ROCK
MINE WATCH
USED IN
PROCetS
ORE
CRUSHING-
WATER «V
LIQUOR FROM
I
GRINDING
LEACHING-
FILTRATION
SODIUM
HYDROXIDE
TAILING-
SAND,
LIQUID WASTES TO
TAILIN&S POND
PRECIPITATION
&AS
FILTRATION
LIQUID
CARBONATED LIOUOR-
TD GRINDING/ LEACH ING-
2HYIN0- g PACKAGING-
OF rcu.ow
Li 6ALANCE AND COMPOSITIONS
/TV
DATA ITEMS
Annual Quantify. Metric TPY
Po/tial Anaiyiii:
U30B. Wt.ro
Radium. ^iC!/g vJ3Os
22Tlia. pCVg
230 Th. pCi/g
234 Tn. pCi/g
Unat. pCi/ liter
226 Ra. pCi/liter
230 Th. pCi/ liter
Iron (Fe) g/ liter
Aluminum (Al)i)/l:ter
Sodium (No) g/liter
Artenic ( AI) g/liter
Fluoride (F)9/liter
Vanadium (V)g/ liter
Sulfote (SO42-) g/liter
Chloride(Cr)g/lite/
Carbonate (CO^~) g/liter
Total Ditiolved Solidt.
g/liter
pH
v^
WASTE
ROCK
51.000
— \^s
CRUDE
ORE
662.200
0.20
1 ® 1
TAILINGS
DRY
SOLIDS
ii2.200
O.OU
SM
565
40
LICUID
695 300
10.000
too
20
0.0005
1.0
3.C
0.0002
0 002
0.0001
2.0
1.0
6.0
12.0
10
YELLOW
CAKE
PRODUCT
1.44?
85
5:5 « 10-3
PROCESS ADDITIVES
NAME
Sodium Hydroxide
Sodium Carbonate
Total
L8/TON
OF ORE
PROCESSED
15
12
•JT
G/MT
OF ORE
PROCESSED
7.500
6.000
13. OM
Source: Reference 36.
Figure 25. Mining and concentrating of uranium ore--alkaline leach process
187
-------
Uranium is recovered from the leach solution by addition of sodium
hydroxide, which forms insoluble sodium diuranate. JL2/ In the clarified
solution, the uranium is present as a complex uranyl tricarbonate ion which
is stable at a pH less than 11. To accomplish precipitation of the uranium
values, the pH of the solution is raised to above 12 by the addition of sodium
hydroxide. At this pH level, the complex ion decomposes to form sodium diuranate
and sodium carbonate. The decomposition reaction is:
2 Na4U02(G03>3 + 6 NaOH - ^ Na2U207 + 6 Na2C03 + 3 H20
The precipitation is carried out at atmospheric pressure and 180 F in a series
of agitated tanks.12/
The caustic barren solution from the precipitation circuit contains
sodium carbonate and a small amount of sodium hydroxide. To permit reuse of
this barren solution, the caustic must be converted to sodium carbonate and
sodium bicarbonate. This required conversion is carried out in packed towers
(recarbonators) in which caustic barren solution is contacted with boiler flue
gas. The carbon dioxide in the flue gas neutralizes the sodium hydroxide and
converts some of the carbonate to bicarbonate. This recarbonated barren
solution is sent to the filtration circuit for use as a wash Iiquor.i2/
The precipitated sodium diuranate (yellowcake) is filtered, washed, and
dried in a steam-heated dryer.J^/ The dried product is packaged in 55-gal. drums
for shipment. Each drum contains a maximum of 455 kg (1,000 Ib).
Off -gas from the drying and packaging area is passed through a dust
separation system (wet impingement dust collector) before discharge to a roof
stack. The overall recovery of uranium is 93 percent of that contained in the
feed ore, and the U^Og content in the product is 85 per cent. ±2.' The estimated
radium content in the yellowcake is 5.5 x 10" p.Ci/g ITjOg, which represents
about 1.8 percent of that in the feed ore.il/ No other significant radionuclides
are present in the yellowcake.
For alkaline leach processes which involve the recovery of by-product
vanadium values, some minor process modifications are required in order to
maximize the production of vanadium values as well as yellowcake. For example,
removal of the vanadium and the carbonate from the initial yellowcake product
is accomplished by a roasting treatment at 1600 F followed by water leaching.
The water leaching serves to dissolve the carbonate and vanadium in the
yellowcake. The vanadium is usually recovered as 'a solution containing
values. and sold to vanadium producers.
188
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In order to meet specifications set by some purchasers for sodium
content in yellowcake, an additional process is required. After removal of
vanadium, yellowcake containing about 7.5 percent sodium is dissolved with
sulfuric acid at a pH of 2.2. Ammonia is then added to this leach solution
to maintain a pH of 7.4 and precipitate yellowcake containing less than 0.5
percent sodium. This product is then filtered, dried, and packaged.
Mass Balance of Materials. Data for a mass balance of materials
involved in the alkaline leach concentrating process are shown in Figure 25.
These data apply for a model plant described in the technical literature.—'
It should be noted that this model concentrator does not represent the design
for any particular existing facility; however, the model is considered to be
representative of all domestic alkaline leach processes.12'
The data for this model show that on an annual basis, the consumption
of 662,200 MT (729,950 cons) of feed ore corresponds to 1,449 MT (1,597 tons)
of product, 662,200 MT (729,950 tons) of tailing solids and 695,300 MT (766,440
tons) of .tailing wastewater.
Process additives include sodium hydroxide and sodium carbonate. The
total consumption rate of additives amounts to 13,500 g/MT of ore processed
(27 Ib/ton ore).ii/ .
The dry waste solids for the model plant are composed of equal parts
by weight of slimes and sands. The slimes fraction contains 85 percent of the
insoluble radioactive materialsJL2' The recovery of uranium in product
(yellowcake containing 85 percent l^Og) is 93 percent of that in the feed ore;
1 percent remains in the sands and 6 percent in the slimes.12' The yellowcake
product contains 2 percent of the radium originally in the feed ore, and 98
percent of the radium is present as an insoluble form in the tailings. The
other radionuclides remain insoluble during leaching and leave the concentrator
with the tailings solids»12/
Total production statistics for uranium-vanadium ores in 1974 are
shown in Table 79, p. 170. From data in Table 82 showing that 13.7 percent
of the concentrator capacity was from alkaline leach plants, it is possible
to estimate the quantity of land-disposed solid,wastes from alkaline leach
processes in 1974 as 13.7 percent of 5,715,000 or 783,000 MT (863,100 tons).
The quantity of total wastewater generated by alkaline leach processes during
1974 is estimated to be 777,000 MT (857,000 tons).
Description of Individual Waste Streams. The only concentrator waste
which is land-disposed is the tailings slurry containing sands, slimes, and
liquid. Airborne effluents containing ore dust, radon gas or yellowcake dust
are discharged to the atmosphere.
189
-------
Data on the composition of waste generated in a model alkaline leach
concentrator plant are shown in Figure 25. For the model plant, the dry solids
fraction of the tailings contains 0.014 percent l^Og, 40 pCi/g of natural
uranium, 560 pCi/g of radium-226, 565 pCi/g of thorium-230, and 40 pCi/g of
thorium-234.
As shown in Figure 25, the liquid portion of the tailings slurry
contains natural uranium and radionuclides (radium-226 and thorium). This
liquid also contains a number of metals including iron, aluminum, sodium,
arsenic, and vanadium as well as fluoride, sulfate and chloride compounds.
The total dissolved solids content in the liquid is about 12 g/liter and the
liquor has an average pH of about 12.
Identification of Potentially Hazardous Waste Streams. The entire
concentrator tailings, which are a land-disposed waste, are potentially
hazardous because they contain radioactive materials in concentrations above
the background level.
Table 83 shows the quantities of potentially hazardous wastes generated
by mining and concentrating uranium-vanadium ores in 1974. The national total
of potentially hazardous dry waste was 7,938,000 MT (8,750,000 tons) consisting
of 2,223,000 MT (2,450,000 tons) of waste rock and 5,715,000 MT (6,300,000
tons) of concentrator tailings solids. The total potentially hazardous dry
waste accounted for 7.8 percent of the total dry waste. In addition, the
estimated quantity of potentially hazardous wastewater discharged for disposal
in tailings ponds was 11,667,000 MT (12,861,000 tons). Therefore, the total
quantity of potentially hazardous wastes generated by this industry in 1974
was 19,605,000 MT (21,611,000 tons).
Modifications of Concentrator Processes. Several modified recovery
methods not discussed in the previous process descriptions can be employed,
following the leaching operations, to yield a uranium concentrate. These
modifications are:
* Ion exchange circuits.
* Alternate solvent extraction systems (Dapex, Amex, Eluex processes).
* Alternate precipitation methods, using sodium hydroxide, gaseous
ammonia, hydrogen peroxide, or magnesia.
* Secondary metal recovery.
The waste products, and potentially hazardous wastes, from these modifications
have been treated in previous subsections which present information as the
basic acid and alkaline leach processes, and therefore will not be discussed
in this subsection.
190
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TABLE 83
TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING URANIUM-VANADIUM ORES FOR SIC 1094 IN 1974
(METRIC)
Dry waste weight
State
New Mexico
Texas
Total
Colorado
Utah
Wyoming
Total
Washington
National
Total
process
EPA waste
Region (103 TPY)
15,743
8.114
VI 23,862
514
790
74.437
VIII 75,741
X 1. , 970
101,573
Total
potentially
hazardous
waste
generated
(103 TPY)
3,576
851
«,427
514
458
2.213
3,185
326
7,938
Waste rock
Total
(103 TPY)
1,127
484
1,611
0
38
353
391
''21
2,223
Potentially
hazardous
(103 TPY)
1,127
484
1,61?
0
38
353
39)
221
2,223
Overburden
Total
(103 TPY)
12,17?
7.261
19,435
0
332
72.224
72,556
1,644
93.635
Potentially
hazardous
(103 TPY)
0
0
0
0
0
0
0
0
0
Dry
weight
Concentrator tailings
Potentially
Total hazardous
(101 TPY) (103 TPY)
2,449
367
2,816
514
420
1.860
2,794
105
5,715
2.449
367
2,816
514
420
1.860
2,794
105
5,715
Wet
weight
Concentrator tailings
Total
(103 TPY)
6,930
2 . 276
9,206
1,092
932
5.889
7,913
263
17,382
Potentially
hazardous
(103 TPY)
6,930
2.276
9,206
1,092
932
5.889
7,913
263
17,382
Source: Reference 36.
-------
Ion Exchange Systems. These systems are used to concentrate and
purify the uranium leaching solution. Strong and intermediate base anion-type
resins are loaded from either a sulfuric acid or a carbonate leach feed
solution. The loaded resin is stripped with a chloride, nitrate, bicarbonate
17 /
or an ammonium sulfate-sulfuric acid solution.—'
Four types of ion exchange circuits are used by uranium mills. They
are: (1) fixed bed, (2) moving bed, (3) continuous resin-in-pulp (RIP)
process, (4) basket resin-in-pulp process.il' The fixed-bed type consists of
stationary columns packed with resin. As the feed solution passes through the
columns, uranium is absorbed on the resin. The resin is washed and the uranium
desorbed. The moving bed column circuit has stationary columns, but the resin
is transferred to different columns to perform loading, washing, and eluting
operations. In the continuous resin-in-pulp (RIP) process, sorption, washing,
and desorption are conducted by contacting the resin and process solutions
in a series of tanks. The resin and solution flow countercurrently in the
tanks and are separated by screens. Forced air is used for agitation. The
basket resin-in-pulp process employs a series of resin-filled cubical baskets
that are jigged up and down in the process solution contained in tanks. Feed
slurry, eluant, and wash solution are passed through the tanks to extract the
uranium. The RIP process represents an exception to the separation procedure
in that suspended solids do not have to be removed from the leach slurry..iZ/
In 1974, there were four concentrator plants which used column ion
exchange units. Taken together, these process units were representative of
about 17 percent of the total domestic concentrator feed ore capacity and the
same percentage of the total solid wastes generated by concentrators and
land-disposed.
During 1974, five concentrator plants used the RIP process. These five
plants represented about 26 percent of the total concentrator feed ore capacity
and the same percentage of the total solid waste which was land-disposed by
concentrators.
Alternate Solvent Extraction Systems. Two solvent extraction processes,
the Dapex and the Ainex, are currently used to concentrate uranium in about
one-half of the concentrators+LL' The processes are efficient but require that
the leach solution be essentially free of solids.
The uranium is extracted from clarified feed solution into an organic
solvent, followed by stripping operations to yield a uranium-bearing aqueous
phase.^2./ The circuit consists of a series of tanks in which feed solution
flows countercurrent to the organic solvent. Phase separation is improved by
optimizing mixing rates of solutions and by installing settlers in the circuit.
192
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The Dapex process uses the alkyl phosphoric acid extractant,
di(2-ethylhexyl) phosphoric acid (EHPA) at a 4 percent concentration in
kerosene with tributyl phosphate (TBP) added as a modifier to improve phase
separation.^' Long-chain alcohols, such as isodecanol, are also used as
modifiers. The loaded organic phase is stripped of uranium with a sodium
carbonate solution and recycled to the extraction stage of the circuit for
reuse.
The Ainex process consists of an amine extraction followed by stripping
with ammonium sulfate, chloride or sodium carbonate solutions.—' A 6 percent
concentration of a tertiary amine, such as alamine-336, in a kerosene diluent
is employed as the organic extractant. Isodecanol is added as a modifier.
Stripping with ammonium sulfate at a pH of 4.0 to 4.3 eliminates sodium
impurities.
The Eluex process is a combination ion exchange-solvent extraction
17 /
process.—' The eluate produced by sulfuric acid elution of the ion exchange
resin is fed to either the Dapex or the Amex solvent extraction process. The
Eluex process eliminates the need for nitrate and chloride reagents and yields
a purer end product. In 1974, this process was used in three concentrator
plants representing about 11.5 percent of total concentrator feed ore capacity
and about 11.5 percent of the total concentrator solid wastes.
Alternate Precipitation Methods. Uranium is precipitated from solution
by the addition of sodium hydroxide, gaseous ammonia, hydrogen peroxide or
magnesia«iZ' Several stages of precipitation at controlled pH are often used.
In most concentrators (12 of 16, representing about 80 percent of the total
concentrator feed ore capacity in 1974), gaseous ammonia is utilized as the
precipitating agent.
Secondary Metal Recovery. In some instances, other metals are present
in uranium ores in sufficient concentrations to permit economic extractions.
For example, a mill in Utah extracts copper as a by-product by means of a
flotation circuit. Vanadium is extracted at one mill in Colorado and one mill
in New Mexico. Concentrations of molybdenum in ores mined in Texas may be
sufficient to justify by-product recovery.
Potentially Hazardous Land-Disposed Waste in 1977 and 1983. Data from the
U.S. Bureau of Mines Mineral Yearbook!!' show that for 1973 the total production
of uranium concentrate was 12,007 MT (13,236 tons) of ^Og. Demand for U30a is
expected to increase at an average annual rate of 18 percent through 1980^=-t'
Midwest Research Institute has estimated that this 18 percent annual increase
will continue through 1983. It is not known whether future production levels
will match the demand for uranium concentrate. However, on the basis of this
estimated increased demand the domestic production of U^Og could reach 23,280
-MT (25,660 tons) of U308 in 1977 and 62,840 MT (69$270 tons) of U30g in 1983.
193
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Based on these estimated production rates, calculations were made to
provide projections of the total waste from domestic mining and concentrating
of uranium-vanadium ores in 1977 and 1983. Previously developed data (Table 79)
on mining and concentrating wastes for 1974 were used as a basis for these
projections. It was assumed that no changes would occur in the regional
distribution of mine and concentrator plants or the ratio of waste generated
to ore mined and processed. Also, it was considered that the projected waste
quantities would be directly proportional to the estimated l^Og demand in a
given year. Thus, factors based on relative U30g demand were applied to the
1974 waste data to obtain projections of waste quantities for 1977 and 1983.
Calculated data for total uranium-vanadium ore mining and concentrating
wastes in 1977 are presented in Table 84. The estimated national potentially
hazardous dry waste is 13,042,000 MT (14,300,000 .tons) consisting of 3,652,000
MT (4,026,000 tons) of waste rock and 9,390,000 MT (10,351,000 tons) of
concentrator tailing solids. Also, the estimated quantity of potentially
hazardous wastewater in the concentrator tailings is 19,169,000 MT (21,130,000
•tons). Thus, the total quantity of potentially hazardous wastes in 1977 is
estimated to be 32,211,000 MT (35,507,000 tons).
Data concerning the uranium-vanadium ore mining and concentrating wastes
in 1983 are shown in Table 85. The estimated total potentially hazardous dry
waste is 35,207,000 MT (38,809,000 tons) consisting of 9,860,000 MT (10,869,000
tons) of waste rock and 25,347,000 MT (27,940,000 tons) of concentrator tailings
solids. The estimated amount of potentially hazardous wastewater discharged
to tailings ponds is 51,743,000 MT (57,037,000 tons). The total'quantity of
potentially hazardous wastes for 1983 is estimated to be 86,950,000 MT
(95,846,000 tons).
Waste Characterization—Vanadium Ores.
Mining Wastes. The only operating vanadium mines in the United States are
open-pit mines located near Hot Springs, Arkansas. At these open-pit mines,
the vanadium-bearing component of the crude ore is vanadinite, Pb^VO-Kd,
a reddish-brown material. Other minerals associated with the ore body are:
illite, geothite, montroselite, pyroxene, quartz, potash feldspar, and
kaolinite. This crude ore contains an average of about 1 percent
194
-------
TABLE 84
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING URANIUM-VANADIUM ORES FOR SIC 1094 IN 1977
(METRIC)
Dry waste weight
State
New Mexico
Texas
Total
Colorado
Utah
Wyoming
Total
Washington
National
Total
process
EPA waste
ReRlon (103 TPY)
25.874
13,331
VI 39,205
845
1,298
122.300
VIII 124,443
X 3,237
166,885
Total
potentially
hazardous
waste
generated
MO3 TPY)
5.875
1,398
7,273
845
752
3,636
5,233
536
13,042
Dry weight
Waste rock
Total
do3 TPY)
1,852
795
2,647
0
62
580
642
363
3,652
Potentially
hazardous
dO3 TPY)
1,852
795
2,647
0
62
580
642
363
3,652
Overburden
Total
do3 TPY)
19,999
11,933
31,932
0
545
118,664
119,209
2,701
153,842
Potentially
hazardous
(103 TPY)
0
0
0
0
0
0
0
0
0
Concentrator tailings
Total
do3 TPY)
4,023
603
4,626
845
690
3,056
4,591
173
9,390
Potentially
hazardous
do3 TPY)
4,023
603
4,626
845
690
3,056
4,591
173
9,390
Wet weight
Concentrator tailings
Total
do3 TPY)
. 11,385
3,739
15,124
1,795
1,531
9,676
13,002
433
28,559
Potentially
hazardous
do3 TPY)
11,385
3,739
15,124
1,795
1,531
9,676
13,002
433
28,559
Source: Reference 36.
-------
TABLE 85
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FUOM MINING AND CONCENTRATING URANIUM-VANADIUM ORES FOR SIC 1094 IN 1983
(METRIC)
Dry waste weight .
State
'New Mexico
Texas
Total
Colorado _
Utah
Wyoming
Total
Washington
National
Total
process
EPA waste
Region (103 TPY)
69,842
35 ,986
VI 105,828
2,280
3,504
330.128
VIII 335,912
X 8,737
450,477
Total
potentially
hazardous
waate
generated
(103 TPY)
15,859
3,775
19,634
2,280
2,032
9,815
1.4,127
1,446
35,207
Dry weight
Waste rock
Total
(103 TPY)
*,998
2,147
7,145
0
169
r.566
1,735
980
9,860
Potential ly
hazardous
(103 TPY)
4,998
2,147
7,145
0
169
1,566
1,735
980
9,860
Overburden
Total
(103 TPY)
53,983
32,211
86,194
0
1,472
320,313
321,785
7,291
415,270
Potentially
hazardous
(103 TPY)
0
0
0
0
0
0
0
0
0
Concentrator tailings
Total
(10? TPY)
10,861
U628
12,489
2,280
1,863
^8,249
12,392
466
25,347
Potentially
hazardous
(103 TPY)
10,861
1,628
12,489
2,280
1,863
8,249
12,392
466
25,347
Wet weight
Concentrator tailings
Total
(103 TPY)
30,734
10,094
40,828
4,843
4,134
26,118
35,095
1,167
77,090
Potentially
hazardous
(103 TPY)"
30,734
10,094
40,828
4,843
4,134
26,118
35,095
1,167
77,090
Source: 'Reference 36.
-------
Description of a Typical Process. At the Arkansas plant, three open-pit
vanadium mines supply crude ore to a nearby mill* The total tonnage of ore
mined annually is about 362,874 MT (400,000 tons) per year. The present
operations involve the use of power shovels and scrapers, and the ore is moved
to the mill by trucks. The overburden is drilled and blasted, and the solid
waste is moved to nearby surface waste dumps, which are contoured and revegetated.
The solid waste disposed of on land amounts to 907,185 MT (1,000,000 tons) per
year. No mine backfilling is currently being done. A flow diagram for the mining
process is shown in Figure 26.
Mass Balance of Materials. At this mine, an average of 2.38 MT (2.62 tons)
of overburden and 0.125 MT (0.138 tons) of waste rock must be handled for every
ton of crude ore produced. The production rate for crude ore is about 362,874
MT (400,000 tons) per year. Mass balance data are shown in Figure 27 and Table 86.
Data on the ratio of wastes to ore and of wastes to product are presented in
Table 87.
Description of Individual Waste Streams. About 95 percent of the solid
waste generated in domestic vanadium mining operations consists of overburden
material containing soil, clay, and various minerals associated with the ore
body, such as quartz and feldspar. The remaining 5 percent consists of waste
rock containing minerals such as illite, geothite, montroselite, pyroxene,
quartz, feldspar, and kaolinite.
Concentrating Operations Wastes.
Roast Leach Process for Vanadium. As mentioned, there is only one production
facility in the United States for the mining and concentrating of vanadium
ore. As shown in the simplified process flow diagram (Figure 27), the basic mill
circuit consists essentially of the following process steps:
1. Drying and grinding of the ore.
2. Blending the fine ore with sodium chloride and extruding the blended
mixture to form agglomerated feed material for roasting.
3. Roasting the mixture to yield a material containing water-soluble
vanadium values.
4. Continuous water leaching, with warm water at about 45 C, at atmospheric
pressure, and a pH of 6 to 8, of the roast in tanks to dissolve the contained
vanadium. This leaching process requires several hours to complete.
197
-------
Drill Ore Body
I
Blast Ore Body
Overburden 864,000 MT
Load & Haul Ore
Waste Rock 45,000 MT
'i 363,000 MT
Surface
Waste Dump
Crush & Grind Ore
I
Ore to Concentrator
Source: Reference 36.
Figure 26. Vanadium mining.
198
-------
ORE FROM MINE
CKUSH AND GRIND
DRYING 6 GRINDING-
NaCl(,
BLENDING- WITH SALT
i EXTKUDING
PREHEATING, ROASTING
& COOLING-
WATEK LEACH &
WASHING- CIRCUITS
PRECIPITATION
SEPARATION, OWING,
PACKAGING I SHIPPING
TAILINGS
SAND PLANT
FINES
N£UTRALIZED
EFFLUENT LIQUOR
AND SOLIDS FKOM GAS SCRUBBER
WATER
SOLIDS
COARSE TO
BUILD 0AM
TAILINGS
POND
EFFLUENT SUKGE
POND
SOLIDS
WATER
EFFLUENT LAKE
DISCHARGE
(LIQUID TO STREAM)
MATE9IA.S BALANCE Ar;D COMPOSITIONS
DATA ITEMS
Quontiry. TPY
M«lric TPY
V^C^ *"o
F«$2 °°
Pb%
-vv •••
ORE
400.000
362.874
1.0
-5
Pr*1""1
- W -
V205
CONC.
5.000
4.536
65 - 7i
0
0
-• -w
TAILINGS
394. 000 (Dry)
NA '
NA
NA
TAILINGS
. A. 000
5.443
NoCI
5.000
4.536
Figure 27. Vanadium mining and processing.
199
-------
TABLt Oo
TOTAL PRODUCTION STATISTICS, BY STATE AND EPA REGION FOR SIC 1094
(URANIUM-VANADIUM* ORES) FOR 1974 (MKTRIC)
Wastes
EPA Ore Over-
state Region mined burden
Arkansas VI 363 862
National 363 862
generated (103 TPY)
Tailings
Waste concentrator
rock wastes
45 368
45 368
Concentrator Products, TPY Ratio of:
Total wet waste Other Total Total waste Total waste Wet waste
wastes (103 TPY) V Cone. metals product to ore to product to ore
1,275 1,587 4,536 0 4, 5.36 3.52 281
1,275 1,587 4,536 0 4,536 3.52 281
4.37
4.37
Source: References 37 and 38.
* Arkansas Is the only state producing vanadium ores.
TABLE 87
RATIO OF TOTAL WASTE KOCK-OVERBURDEN-CONCENTRATOR WASTE TO ORE MINED FOR SIC 1094
(URANIUM-VANADIUM* ORES) FOR 1974 (METRIC)
S3 —
O
0
EPA Ore
State Region CIO TPY) '
Arkansas IV 363
National 363
Ratio of Ratio of
Ratio of total over- Concentrator total concen-
Waate rock total waste Overburden burden to wastes trator waste Concentrator
(103 TPY) rock to ore (JO3 TPY) ore (103 TPY) to ore wet waste
45 0.12
45 0.12
862 2.38 368 1.01 1,587
862 2.38 368 1.01 1,587
Ratio of
wet waste
to ore
4.37
4.37
Source: References 37 and 38.
* Arkansas ic the only state producing vanadium ores.
-------
5. Processing of the leach solution is accomplished by solvent extraction
to produce a modified vanadium oxide* Tailings from this recovery system are
sent to a tailings pond.
6. Separating, drying, packaging, and shipping of product vanadium oxide
concentrate to a sister plant for further processing*
Vanadium ore obtained from a stockpile is dried and then ground. The drier
and mill are equipped with dust collectors and wet scrubbers for treatment
of off-gases before they are sent to a stack.
The ground ore is blended thoroughly with crushed salt (NaCl), and the
blended material is then processed through an extruder to prepare agglomerated
material suitable for use as feed for a roasting operation.
The agglomerated mixture of ore and salt is then preheated and roasted at
about 850 C to convert the contained vanadium to a water-soluble sodium
metavanadate (NaVO,)* The roaster is equipped with an off-gas scrubber. The
vanadium is extracted by leaching with warm water (45 C) and precipitated as
sodium hexavanadate, Na^V^Oiy (referred to as redcake) by the addition of
sulfuric acid to adjust the pH to between 2 and 3.
The precipitated vanadium compound is separated from the liquid, dried,
packaged, and shipped to a sister plant for further processing to a higher
grade vanadium oxide product. The solids are sent to a tailings pond. The
liquor is sent to an effluent pond where it is treated and discharged to a
stream. This method of disposal has been judged satisfactory and complies with
effluent limitations guidelines.
Mass Balance of Materials. In the Arkansas vanadium facility 362,874 MT
(400,000 tons) per year of ore are processed in the concentrator by the roast-
leach process. The total quantity of waste or solids discharged to a tailings
pond is about 357,000 MT (394,000 tons) per year (dry weight). In addition,
about 5,400 MT (6,000 tons) per year (dry weight) of waste solids from the
effluent surge pond are sent to a tailings impoundment.
Description of Individual Waste Streams* The tailings discharged from
the water-backing step of the concentrator process contain gangue material
originally present in the feed ore. In 1974, there were 368,000 MT (405,651
tons) of tailings concentrator wastes. The ratio of total concentrator waste
to ore was 1.01. There are no potentially hazardous wastes destined for land
disposal from the mining and concentrating of vanadium ore.
Identification of Potentially Hazardous Wastes. There are no potentially
hazardous wastes resulting from the treatment and disposal of wastes generated
in the mining and concentrating of vanadium cres.
201
-------
Waste Treatment and Disposal--Uranium
The potentially hazardous wastes which are land-disposed in the mining and
concentrating of uranium ores are the mined waste rock located adjacent to
the ore deposit and the concentrator tailings (a slurry of sand, slime, and
liquid).
All domestic waste treatment and disposal operations practiced in the
uranium mining and concentrating industry are conducted on the plant sites
by the operating companies, and no waste materials are sold. A discussion
of the treatment and disposal practices for these potentially hazardous
wastes follows.
Open-Pit Mining Wastes. Open-pit mining accounted for about 58 percent
(6,629 MT of U30g) of the total uranium mine output in 1974r^ The estimated
total number of active open-pit mines during 1974 was 122*=^ Open-pit mining
operations in Wyoming are typical.—
In 1974, the. estimated national total of waste rock mined in open-pit
uranium mine operations amounted to 2,026,000 MT (2,233,300 tons). This
tonnage represented about 91 percent of the total waste rock mined during
1974 from both open-pit and underground mines
Waste Rock Treatment and Disposal. Waste rock, the only potentially
hazardous waste involved in open-pit mining, is disposed of in either surface
waste dumps or in a very few cases in mine backfilling operations. Most
plants dispose of the waste rock along with the overburden waste (nonhazardous)
which is also removed from the pits and comingled with the waste rock. Some
mines segregate the waste rock from the overburden and store the waste rock
in open surface piles for future use as a low grade feed ore; the estimated
amount stored was 0.5 to 1 percent (33-66 MT of U-jOg) of the total ore output
of U-jOg values in 1974 for open-pit operations.
Levels I, II, and III Technology. Figure 28 is a flow diagram showing the
representative waste disposal and treatment operations for Level I technology
for disposal of potentially hazardous wastes based on 10,000 MT ore, waste
rock, and overburden combined. This level of technology applies for about
all of the operating mines. The waste rock, which represented on the average
about 2.1 percent (208 MT) of the total material mined in the open-pit
operation, is disposed of in either a surface dump or as backfill (along
with comingled overburden which is a nonhazardous waste) in abandoned areas
of open-pit mines. About 93 percent (193 MT) of the waste rock is permanently
disposed of in surface dumps along with overburden material. The dump surface
is contoured and stabilized by natural vegetation. The remainder of the waste
rock (7 percent or 15 MT) is used along with overburden as backfill material
for abandoned open-pit mine areas, following temporary storage in surface
dumps.
202
-------
W.R.193 MT*
Chemical
Additives
to
o
i
= WASTE ROCK* 208 MT Plus
= Overburden 9.469 MT Non-
hazardous
O.B.7JOOMT
.B; 2.369MT
Ore
323 MT**
W.R. 15 MT*
Concentrator
(Acid Leach Process)
TAILINGS'
323 MT
Earthdike
Tailings Pond-
Evaporation,
Seepage
Natural Vegetation
on Dam and Dry
Tailing Surfaces
Zero Discharge of
Wastewater
Surface Waste Dump
Overburden Plus Waste Rock
Dump Surface Contoured
and Stabilized by Natural
Vegetation
Mine Backfill of Overburden
Plus Waste Rock
Contoured Surface with
Natural Vegetation
Source: References 36 and 37.
Legend:
*Potentially hazardous waste.
**The numbers represent the estimated weighted average distribution of total
materials mined, based on 10,000 metric tons.
Figure 28. Level I technology for treatment and disposal of potentially hazardous wastes in open-pit
uranium ore mining and in ore concentrator operations.
-------
In Level I technology, the surface waste dump method for disposal of a
mixture of waste rock and overburden is not considered to be adequate for
environmental protection. Under these conditions, the exposed surface areas
of the pile not stabilized by natural vegetation are vulnerable to erosion
by wind and water which could cause pollution problems.
The disposal method of using the waste rock and overburden mixture for
mine backfilling operations is also considered to be unsatisfactory as a
means of environmental protection because of potential erosion problems.
No significant effects on the future adequacy of Level I technology are
anticipated as a result of changes in the volume and composition of wastes
as new air and water pollution control facilities are installed or existing
facilities are improved to meet more stringent control standards. It is
unlikely that there will be any significant change in the volume or
composition of waste rock due to air and water pollution enforcement actions.
Level I technology is adaptable for use with any domestic open-pit uranium
mining operation. Either mode or both modes of disposal (i.e., surface waste
dump or mine backfill) can be readily adapted at a given mine, and the
deciding factor is usually the economics involved in the disposal operation.
Backfilling is more expensive than disposal on a surface pile.
Energy requirements differ considerably between surface dump disposal and
mine backfilling; the latter method consumes by far the greater amount of
energy. Except for the energy requirements for waste disposal, there are no
significant nonland related impacts involved in Level I technology.
Waste rock contains higher than background levels of radioactive materials
consisting of uranium and its radioactive decay products (e.g., radium,
thorium, etc.). These radioactive materials are the principal environmental
hazards in the waste.
Factors which affect the potential hazards presented by the waste include
the prevailing climate and the mode of disposal. For example, very windy
locations pose an increased hazard that windborne particles of tailing
solids may be carried outside the perimeter of the mining area.
204
-------
The potential problems with Level I control technology center around the
risk of erosion (i.e., wind and rainwater transport) of radioactive particles
of waste rock. This risk applies particularly to the surface waste piles and,
to a reduced degree, to backfilled areas. In instances where the waste rock
is covered by a thick layer of overburden, this hazard probably ranges from
negligible to very low. However, in cases where the waste rock is present on
exposed surfaces of waste piles, and therefore is vulnerable to erosion, the
environmental hazard is significant. In some geographic areas which have low
rainfall levels (e.g., arid states such as New Mexico), the annual rainfall
is insufficient to support natural vegetation on the surface piles. There is
considerably less risk of erosion in the northwestern mining states (e.g.,
Wyoming and Washington), where natural vegetation is more easily maintained.
Because most of the waste rock consists of large lumps and coarse particles,
the potential hazard of windblown particulates (i.e., small particles)
probably ranges from low to negligible depending upon climatic factors such
as prevalence and velocity of winds in a given locality. Areas of high annual
precipitation have a relatively high risk of water erosion.
Level I technology can be applied at a mine site without any significant
implementation time requirements or equipment purchases. Regular mining
equipment is used in the waste disposal operations. This technology is not
expected to have a significant impact upon noise pollution problems or to
eliminate all potential problems concerning environmental pollution. A
potential intermedia pollution problem, which has not been thoroughly
investigated, involves the possibility of surface water pollution by
airborne contaminants transported off the plant property.
It should be especially noted that Level I technology contains no provisions
for minimizing the dissemination of potentially hazardous wastes into air and
water during the active lifetime of a waste dump, i.e., during the period of
time (which may be several years) when the dump surface is being built up to
the final contour which can then be allowed to stabilize with natural vegetation.
In Level I technology, monitoring is desirable to determine whether potentially
harmful pollutants are entering the environment, and if so, to what extent. This
activity should include periodic chemical analysis of samples of air, soil,
vegetation, and surface water at the perimeter and near the plant site. In
conjunction with these tests, natural background levels of radioactivity would
be useful for comparison with the test samples.
Level II technology for disposal of potentially hazardous materials, which
is also Level III technology, is practiced by about 10 percent of open-pit
mines in areas where the state reclamation laws require application of
advanced technology for waste disposal»
205
-------
A flow diagram showing the Level II/III technology is presented in Figure
29. In this disposal method, waste rock segregated from overburden is
deposited in a surface dump. After the dump becomes inactive, it is contoured,
covered with a layer of compacted topsoil or approved subsoil, vegetated by
seeding, and fertilized if necessary, to maintain a stabilizing plant cover
which minimizes the erosion of the waste rock by wind and runoff water. The
depth of soil cover used varies from one plant locality to another, depending
on factors such as inherent differences in climate, topography, and
availability of soil. No data were found which establish the required
thickness of topsoil in different localities for long-term stabilization.
f
The period of greatest environmental risk in Level II/III technology is the
active lifetime of the waste dump. During this period, which may under current
practice last several years, the surface of the dump is in a state of flux,
and is susceptible to water/wind erosion, and thus transport of potentially
hazardous waste into water and the atmosphere. This risk can be reduced by
shortening the effective active lifetime of the dump, i.e., by constructing
a series of small dumps rather than a single large one or by periodically
finishing and stabilizing segments of a large dump. Another alternative,
particularly applicable to open-pit mines, consists of periodically covering
layers of potentially hazardous waste rock with layers of nonhazardous
overburden, i.e., operating the dump in a manner similar to a sanitary
landfill in which each day's fill is covered with soil. The Level II/III
technology will in the overall sense be adequate only if measures such as .
the above are, as appropriate in a specific mining operation, conscientiously
implemented. The preferred measure is periodic covering of potentially
hazardous waste rock with nonhazardous overburden or with available on-site
soil materials. Added costs are expected to be essentially zero when overburden
is abundantly available and must in any event be disposed in waste dumps. When
overburden is not available, there will be an added expense associated with
costs of moving and emplacing protective covers, or with scheduling the dump
construction/stabilization protocol to reduce the effective active lifetime
of the dump. The added:cost has been estimated, in the following section on
costs, for one assumed case in which an annual accumulation of wastes is
covered with on-site generated cover (not overburden).
In arid regions (e.g., New Mexico) where watering would be required on a
continuing basis to support vegetation, and where water is expensive, an
acceptable alternate control method is available. This alternate method
consists of contouring the surface of the inactive waste pile to conform
to the landscape and then applying a layer of gravel or crushed rock onto
the surface as a stabilizing cover.
206
-------
Open Pit
Mine
ro
o
-j
Chemical
Additives
WASTE ROCK*
208 MT
Overburden
9.469MT
Ore**
323 MT
Concentrator
(Acid Leach Process)
TAILINGS'
323 MT
Surface Waste Dump
Waste Rock
Contoured, Covered with Topspil
or Approved Subsoil, Vegetated
by Seeding, Watering and Fert-
ilized, if Necessary, to Maintain
Stabilizing Plant Cover
Earthdike
Tailings Pond-Evaporation, Seepage
Stabilization of Dam and Dry
Tailing Surfaces with Seeding,
Watering, and Fertilization,
if Necessary, to Maintain
Stabilizing Plant Cover
Source: References 36 and 37.
Legend:
* Potentially hazardous waste.
**The numbers represent the estimated weighted average distribution of
total materials mined, based on 10,000 metric tons.
Figure 29. Level II/Level III technology for treatment and disposal of potentially hazardous wastes in
open-pit uranium ore mining and in ore concentrator operations.
-------
Several mines segregate waste rock from the overburden during mining
operations and store the waste rock in surface piles for future use as a
low-grade feed ore; for. 1974, the estimated amount stored was 0.5 to 1
percent (33-66 MT of U-jOg) of the total output in U^Og values for open-
pit mines.
The Level II/III control technology, if conscientiously applied and
maintained, should provide adequate protection against environmental
pollution. The risks associated with this technology are concerned with
the application of sufficient soil cover and seeded vegetation to ensure
a uniform and complete plant cover which thoroughly stabilizes the surface
and eliminates erosion problems, and with the care taken to minimize erosion
when the dump surface is still in the active stage. For example, an inadequate
depth of original soil1 cover on the waste rock pile could result, in very
windy areas, in the soil being scoured off before the plant growth is able
to stabilize the surface. The required depth of topsoil or subsoil varies
from one locality to another depending on climatic conditions,, availability
of soil, and other factors.
No significant changes in volume or composition of the waste rock are
expected as a result 'of future air and water pollution enforcement actions.
Also, the new pollution control measures would probably not adversely affect
the efficiency of Level II/III technology. All existing installations could
be brought up 'to Level II/III technology without additional equipment or any
serious retrofit problems.
With the exception of increased energy and chemical requirements (for
applying soil covering and fertilizing), no significant nonland-related
environmental impact is anticipated as a result of the implementation of
this technology.
The technical discussion of the radioactive properties of waste rock given
for Level I technology also applies here. Factors affecting the hazards
presented by the wastes include climatic conditions and operating care
exercised in maintaining a uniform and complete' cover of plant growth over
the entire surface of the waste rock pile. The principal problems concerned
with Level II/III technology are the potential difficulties to be encountered
in maintaining an adequate growth of vegetation cover on waste dumps in
certain arid states where the climatic conditions are not amenable to
revegetation as a reclamation method. For example, in some uranium ore
processing areas (e.g., New Mexico), the annual1 rainfall is insufficient
to sustain some seeded vegetation. As stated earlier, in arid regions of
this type, it is believed that a suitable alternative procedure would involve
depositing a covering layer of crushed rock or gravel upon the contoured
surface of waste dumps as a means for minimizing or possibly eliminating
erosion of waste rock by wind and water runoff. Such an alternative method
should be completely effective in eliminating any windborne transport of
waste rock particles off of the plant site.
208
-------
No significant amount of implementation time would be required to apply
the Level II/III technology, which can be largely accomplished using regular
mining and concentrating equipment.
For Level II/III technology, monitoring is desirable to establish that
no hazardous pollutants enter the environment because of wind and water
erosion of waste piles. This monitoring should provide for periodic sampling
and chemical analysis of air, soil, vegetation, and surface water at the
plant perimeter and in the immediate plant vicinity.
Underground Mining Wastes.
Waste Rock Treatment and Disposal. Underground mining represented about
40 percent (4,572 MT of U30g) of the total uranium mine output in 1974^'
The estimated total number of active underground mines during 1974 was 33 .•=-£'
Underground uranium mining operations in New Mexico are considered to be
typical.
Waste rock is the principal potentially hazardous waste generated in
underground mining. Mine water containing some dissolved uranium compounds
is also potentially hazardous. Generally, this water is treated to recover
the uranium values or is used in a concentrator process. Some waste rock
is usually left in the mine. Some mines store the waste rock in surface
piles for use as a low-grade feed orej for 1975, the estimated amount stored
was 0.5 to 1 percent (23-46 MT of U^Og) of the total ore output for underground
mines.
The disposed or stored waste rock contains above background levels of
potentially hazardous radioactive materials such as uranium and its radioactive
decay products, A description of these radioactive properties is given in the
previous section on open-pit mines. Both the prevailing climate and the
topography in the disposal area are critical factors affecting the hazard
presented by the waste rock.
Levels I, II, and III Technology. Level I technology for disposal of
potentially hazardous waste (Figure 30) in underground mining of uranium
ore consists of depositing waste rock onto open surface dumps, terracing
the surface to conform to the natural landscape, and relying on natural
vegetation to provide a cover of native plants or grass. This level of
technology is practiced by almost all of the operating plants.
The Level I technology is not considered to be environmentally adequate
because exposed surfaces, not protected by natural vegetation on the dump,
are subject to wind and water erosion of the waste reck, which could result
in environmental pollution problems.
209
-------
Ni
H-*
O
Chemical
Additives
Underground
Mine
Ore**
9.280MT
Concentrator
(Acid or Alkaline
Leach Process )
WASTE ROCK*
720 MT
TAILINGS*
9.280MT
.Surface Waste Dump,
Contoured, Natural
Vegetation
Earthdike Tailings Pond-
Evaporation, Seepage
Natural Vegetation on Dam
and Dry Tailing Surfaces
Zero Discharge of
Wastewater
Source: References 36 and 37.
legend:
* Potentially hazardous wastes.
**The numbers represent the estimated weighted average distribution of total
materials mined, based on 10,000 metric tons.
Figure 30. Level I technology for treatment and disposal of potentially hazardous wastes in underground
uranium ore mining and in ore concentrator operations.
-------
The future installation of new air and water pollution control facilities
is not expected to materially affect either the volume or the composition
of the land-disposed wastes.
Since all mines are applying at least Level I technology, there are no
problems of retrofitting this technology to existing installations. With
the exception of the energy requirements for contouring the waste dumps,
there are no significant nonland-related environmental impacts. Level I
technology can be implemented readily with little or no time delay at any
new mine site. Application of this technology does not significantly affect
the existing noise pollution problems at a given site.
Monitoring activities comprised of periodic sampling and analysis of
samples of air, soil, vegetation, and surface water at the perimeter of
the plant site are desirable, particularly since it has been concluded
that Level I technology may not adequately protect the environment.
As shown in Figure 31, Level II technology, which is also considered
to be Level III technology, includes depositing waste rock from underground
mines on open surface dumps, contouring the surface to match the natural
landscape, spreading a layer of topsoil or approved subsoil onto the
contoured surface, and revegetating by seeding for stabilization against
erosion. Fertilization and watering are carried out, if necessary, to
maintain plant growth. It is estimated that about 10 percent of the mines
use this technology on inactive dumps. The Level II/III technology described
above for inactive dumps must include measures designed to protect the dump
surface during the active stage as discussed earlier for open-pit mining
wastes. These measures include providing interim covers of nonhazardous soil
materials or scheduling dump construction so that dump surfaces are in the
active stage for relatively short periods of time.
In arid regions where rainfall is inadequate to support plant growth
and irrigation is expensive, an alternative stabilization method, as for
open-pit mine wastes, which involves application of a layer of crushed rock
to the surface of contoured piles, is considered by the authors of this
report to be an acceptable and effective procedure.
The Level II/III control technology should be environmentally adequate,
if it is conscientiously applied. Risks associated with this control method
are concerned with the application of sufficient soil cover and revegetation
(or crushed rock) to provide a complete stabilizing cover which prevents
problems with erosion of the waste rock.
The small anticipated changes in the volume and composition of the waste
rock as future air and water pollution control facilities are installed
would not adversely affect the adequacy of the Level II/III technology.
211
-------
Underground
Mine
ro
\->
IS)
Chemical
Additives 1
WASTE ROCK.*
720 MT
Ore**
9,280 MT
Concentrator
(Acid or Alkaline
Leach)
TAILINGS*
9.280MT
Surface Waste Dump, Contoured,
Covered with a Layer of Topsoil
or Approved Subsoil, Vegetated
by Seeding, Watering and Fert-
ilized, if Necessary, to Maintain
Stabilizing Plant Cover
Earthdike Tailings Pond-Evaporation,
Seepage, Stabilization of Dam and
Dry Tailing Surfaces with Seeding,
Watering, and Fertilization,
if Necessary, to Maintain Plant
Growth
Zero Discharge of Wastewater
Source: References 36 and 37.
Legend:
* Potentially hazardous wastes.
**The numbers represent the estimated weighted average distribution of total
materials mined based on 10,000 metric tons.
Figure 31. Level II/III technology for treatment and disposal of potentially hazardous wastes in
underground uranium ore mining and in ore concentrator operations.
-------
No problems are anticipated in retrofitting the Level II/III technology
to installations having Level I technology, aside from the increased cost
for the improved control system. The nonland-related environmental impact
would be primarily the increased energy and fertilizer requirements for
Level II/III operation.
The Level II/III technology could be planned and implemented at any Level
I site or new facility within 6 to 12 months. It is not anticipated that
this technology would have any significant impact on the existing noise
pollution problems at a given site.
A discussion of the potentially hazardous constituents in the waste rock
is given in the description of Level I technology for open-pit mining.
Factors affecting the hazards posed by the waste rock include climatic
factors (e.g.} wind and rain storms and other severe weather) and the
degree and continuity of operating care used in maintaining an adequate
stabilizing cover on the waste dump. Thus, the reliability of the
environmental protection would depend in large measure on continuing
maintenance of properly stabilized waste dumps and on the adequacy of
measures adopted for protecting surfaces of active dumps.
The activities necessary to adequately monitor the disposal operations
involve periodic collection and analysis of samples of air, soil, vegetation,
and surface water at the plant perimeter and in the vicinity of the mine
site.
Concentrator Operation Wastes.
Waste Treatment and Disposal. The potentially hazardous land-disposed
wastes in the uranium ore concentrator operations are the tailings which
consist of a slurry of sand,- slime, and liquid. Large quantities of
potentially hazardous tailings wastes have been generated by domestic
plants. For example, from 1948 through 1972, the domestic uranium
concentrator industry processed 93,510,300 MT (103,078,000 tons) of ore
containing 221,100 MT (243,700 tons) of U308.— This operation represents
an accumulation of about 93 million metric tons (102.5 million tons) of
solid waste (tailings),^'
All uranium ore concentrator plants in the United States use the same
method for disposal of tailings, i.e., the use of a tailings pond
(impoundment basin) for containment of the solid wastes. Liquid disposal
is accomplished by natural evaporation and seepage from the pond. No
evidence has been found that pond seepage migrates beyond the plant.
perimeter or causes any pollution problems in current operations.—
The size of these tailings ponds varies from a few hectares to over
100 hectares (250 acres), and from 1 to 10 ponds may be used at individual
concentrator sites.—'
213
-------
The tailings slurry is commonly spigotted into the impoundment area from
a peripheral pipeline. The dam height is increased periodically as needed
by application of coarse sand separated from the tailings slurry by a
portable cyclone; the separated slimes are discharged away from the
embankment areas into the pond. The tailings pond construction and operation
are basically the same for all types of concentrator plants.
Of the 16 active domestic concentrator plants in 1974, 12 plants used an
acid leach process and two plants used a similar dual process which involves
both an acid leach and an alkaline leach step. These concentrators
represented about 85 percent (10,143 MT) of the total production of 11303
concentrate and generated 80 to 90 percent (4,572,000-5,144,000 dry MT) of
the total concentrator tailings waste during 1974.
In concentrating processes using the acid leach process, the average
quantity of solid tailings waste is .estimated to be 1 MT/MT of uranium
feed ore consumed in the process.-" Also, the total quantity of waste
solution accompanying the solids (consisting of sands and slimes) in the
tailings slurry is estimated as 1.5 MT/MT of the feed ore processed*^'
Sears et al^/ estimate that the slime fraction (< 200 mesh particles)
contains 85 percent of the insoluble radioactive materials originally in
the feed ore.— The weight ratio of sands to slimes in the dry solids
fraction of tailings is 2:33 to 1. Detailed data on the volume and
composition of the tailings for a model plant using the acid leach process
are shown in Figure 24, p. 181. Potentially hazardous radioactive materials
are contained in both fractions of the solid waste and in the liquid waste.
During 1974, two domestic concentrator plants used the alkaline leach
process for production of uranium concentrate. These plants contributed
about 14 percent (1,651 MT) of the total concentrate production and
generated about 10 to 20 percent (572,000-1,143,000 dry MT) of the total
tailings in 1974.
For concentrator plants using the alkaline leach process, the average
quantity of solid tailings waste is estimated to be 1 MT/MT of feed ore
processed.—' The total waste solution contained in the tailings slurry
is about 1.05 MT/MT of the feed ore consumed.— The slime fraction contains
about 85 percent of the insoluble radioactive materials originally present
in the feed ore.~ The dry solids fraction of tailings consists of equal
parts by weight of sands and slimes. Detailed data on the volume and
composition of the tailings for a model alkaline leach plant are given
in Figure 25, p, 187,/All fractions of the tailings waste contain potentially
hazardous radioactive materials. '
214
-------
Level I Technology. The Level I technology presented in this report deals
with active tailings ponds. For a discussion of methodologies applicable to
inactive tailings ponds, the reader is referred to a cost benefit study!!'
which encompasses inactive tailings ponds.
Level I technology for the waste treatment and disposal of the potentially
hazardous wastes from uranium concentrating processes involves discharging
all the tailings waste from the concentrator to an earthen-dike tailings
pond as shown in Figure 28, p. 203. The residual wastes finally disposed in
the tailings ponds include the sand, slime, and liquid contained in-.;the'
tailings slurry discharged to the tailings pond. Natural evaporation and
seepage are the principal means for disposal of liquid contained in the
tailings slurry.— In some plants, a portion of the settled clear liquor
in the tailings pond is recycled to the concentrator process, especially
in areas where water is in short supply. Waste liquor from concentrators
using an acid leach contain various heavy metals and are generally unsuitable
for recycling without special treatment. The wastewaters from alkaline
leaching are more amenable for reuse and are, therefore, commonly recycled
in part to the process.—'
All of the domestic concentrator plants use Level I disposal technology.
Most of the concentrators (about 80-90 percent) have zero discharge of liquid
effluent from their tailings ponds. The other plants carefully treat and
monitor the effluent liquor to remove all potentially hazardous pollutants
prior to discharge into water courses,^'
Stabilization of the surfaces of the impoundment dam against erosion is
accomplished in Level I technology by growth of natural vegetation. In some
localities, natural plant growth provides good surface stabilization. Some
natural vegetation also occurs on the exposed dry surface areas of the
tailings ponds, and this aids in stabilization of the surface against wind
erosion. In some arid localities, however, the natural rainfall is inadequate
to support a suitable cover of vegetation. Natural vegetation, thus, is not
in the general sense considered to be adequate environmental protection.
In the Level I technology for concentrator wastes, plants which have zero
discharge of liquid effluent (80-90 percent of total Level I plants) are
considered to be fully adequate from the standpoint of protection against
pollution by wastewater. Also, the practice in some Level I plants of
discharging some treated wastewater (i.e., treatment to remove potentially
hazardous materials) is considered to be environmentally adequate; however,
this practice involves risk of careless handling which could result in
improperly treated effluent and pollution of surface streams. Therefore,
zero discharge is considered to be a safer method of waste treatment and
disposal than treatment and discharge of wastewater.
215
-------
The only risk of environmental pollution in the Level I technology is
concerned with the possibility of windborne losses of tailings from exposed
areas of the tailings dam and dry beaches in the tailings pond. Depending
upon climatic conditions and topography in a given concentrator locality,
there is a significant risk that potentially hazardous tailings solids
can be carried beyond the limits of a plant site by prevailing wind conditions.
In some geographic areas (e.g., Colorado), natural vegetation is reported to
be an effective method for stabilizing the surface of tailings dams. However,
in some arid climates, natural vegetation is not adequately supported by the
existing rainfall in the region, and insufficient natural groundc'over is
present to prevent erosion.
The future installation of new air and water pollution control or
improvement of the efficiencies of existing facilities is not expected to
have a significant impact on either the volume or composition of concentrator
plant wastes. These changes would probably have a negligible effect on the
volume of wastes and their composition since nearly all of the existing waste
materials are presently being collected and disposed. Therefore, the adequacy
of the Level I technology would be essentially unaffected by future changes
in air and water pollution control equipment. There are no problems of
retrofit of waste treatment and'disposal technology since all existing plants
utilize at least the Level I technology. Level I technology can be applied at
a concentrator site well within the time required to plan and construct the
uranium ore concentrator facility. The technology for construction and
operation of tailings ponds has been well established in the mining industry
for many years; implementation of this technology does not involve any
special problems in engineering studies, equipment, delivery^ and startup.
The nonland-related environmental impact is concerned principally with the
energy requirement for installation and operation of the tailings ponds.
This requirement includes the energy required to construct the initial
impoundment dam to pump the tailings slurry from'the concentrator plant to
the tailings pond and to periodically build up the height of the dam. This
level of technology does not create any significant noise pollution problems.
However, it is conceivable that this technology could, in some instances,
cause problems concerning air and water pollution in the immediate vicinity
of the operating site. For example, windborne particles of tailings waste
could be transported beyond the perimeter of the plant. A possible, but
unproven, intermedia problem is concerned with the possible pollution of
surface water by contaminants transported from a tailings pond area (e.g.,
dry tailings areas or exposed dam surfaces).
216
-------
Regularly scheduled monitoring of the treatment and disposal facility
operations is desirable. This activity would comprise periodic chemical
analysis of air, soil, vegetation, surface water, and groundwater at the
plant perimeter and in the plant vicinity. These monitoring tests would
serve to establish whether any tailings waste was being transported by
wind or water erosion into the environment. Data on natural background
levels of radiations would be needed for comparison with the test sample
data.
Level II/III Technology. Level II technology for disposal of potentially
hazardous wastes which is also considered to be Level III technology consists
of tailings ponds having zero discharge of effluent and special treatment to
minimize erosion losses of tailings. The exposed surfaces on the impoundment
dam and dry tailings in the pond area are stabilized with seeded vegetation,
fertilized, and watered, if necessary, to maintain plant growth. About 10
percent of the concentrator plants use the Level II/III disposal technology.
The impoundment dam is periodically built up with separated coarse tailings
sand by the same method as that described previously for the Level I
technology. Since all plant operations at this level of technology have
zero discharge of effluent wastewater from the ponds, there is no risk of
environmental pollution due to discharge Of any liquid waste.
Stabilization with seeded vegetation of the exposed surfaces of the
tailings dam and the dry tailings beaches serves to minimize the risk of
windborne transport of tailings dust beyond the perimeters of the plant
site. The conscientious application and maintenance of vegetation cover
should serve to adequately protect the environment against pollution by
wind or water erosion of finely divided tailings particles. Improper
maintenance of a vegetative cover could result in significant risks of
airborne pollution, particularly in windy locations.
It is not anticipated that the installation of new or improved air and
water pollution control facilities would significantly affect the volume
or composition of wastes generated by the uranium ore concentrator plants.
Therefore, it is believed that the future adequacy of Level II/III technology
will remain essentially unchanged after new control facilities are installed.
With one exception, there are no serious problems of retrofitting the
Level II/III technology to existing installations having the Level I
technology. Level I plants which own insufficient land to develop the
required size of tailings pond may have to continue to treat and discharge
a portion of their tailings wastewater. For Level I plants having no liquid
discharge from thre tailings ponds, the changeover to Level II/III is
relatively easy, i.e., a change from the use of natural vegetation to
seeded vegetation as a means of stabilizing the surface of the active
tailings ponds.
217
-------
The principal nonland-related environmental impact of Level II/III
technology is the energy requirement for installing and operating the
tailings ponds and the chemicals requirement for fertilization.
Under certain conditions, Level II/III technology could be rendered
partially ineffective in regard to possible air pollution problems. If
the seeded vegetative cover on the dam and the dry areas of the active
tailings pond are not continuously maintained by periodic reseeding,
fertilizing, and watering as necessary, then barren areas would develop,
and air pollution problems could result.
Some treatment of liquid waste discharge would be required in the event
that the volume of concentrator operations was greatly increased without
a corresponding increase in the available tailings pond disposal area.
The estimated implementation time to convert a Level I installation to
Level II/III including engineering studies, equipment delivery, and
installation for applications and maintenance of seed vegetative cover
is 6 months to 1 year0 The application of Level II/III technology at a
given site would not create any new problems concerning air, water, and
noise pollution*
The activities necessary to properly monitor the waste disposal operations
involve a regularly scheduled system for collection and analysis of test
samples of air, soil, vegetation, and surface water at the plant perimeter
and the immediate vicinity of the site.
218
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Cost Analysis--Uranium-Radium-Vanadium Ores
Disposal Costs--Open-Pit Mine Wastes.
The rationale and assumptions used in developing a representative model
for open-pit uranium mining operations are explained on page 172 under "Mass
Balances of Materials." The mass balance data for a representative plant
model and for waste treatment and disposal practices of open-pit mining
operations are shown in Figures 24 (p. 181), 28 (p. 203), and 29 (p. 207).
Waste Treatment Costs - Surface Waste Dump, Level I Technology. The waste
treatment costs of uranium mining and concentrating will be broken into two
major categories. First, the costs of depositing overburden and waste rock
in surface waste dumps and mine backfill will be examined. Next, the cost of
treating the tailings from the ore concentrator will be assessed.
The representative open-pit uranium mine creates 1.942 by 10 MT of
overburden per year of which 1.46 by 10 MT (75 percent) is deposited in
the surface waste dump. The rest is used as mine backfill. The mine also
produces 425,700 MT of waste rock per year. Approximately 395,000 MT (93
percent) of the waste rock is deposited in the surface waste dump. The
Level I disposal technology does not separate overburden and potentially
hazardous waste rock.
The first step in this calculation is to determine the capital costs of
the surface dump. A total of 1.5 by 10 MT/year of waste and overburden are
deposited on the dump. We will assume a bulk density of both materials equal
to 1.96'MT/m3 (1.5 MT/yd3).3-^ Therefore, 7.62 by 106 m3/year (9.97 by 106
yd /year) are deposited. Over 20 years, the dump will contain 1.52 by 10 m3
(1.99 by 108 yd3) of rock. If the dump is 7.62 m deep (25 ft), the land
required would be 1.99 by 107 m2 or 1,995 hectares (4,929 acres). Assuming
a land value of $12,350/hectare ($5,000/acre) in 1973 dollars, the total
land cost will be approximately $24.05 by 10° for the surface dump.
The $24 million cost of the surface dump can enter the capital cost
calculation in at least three distinct ways. First, one can assume all 1,995
hectares are bought during the 1st year of the analysis. This land may have
the same value at the end of the period of analysis (20 years) or have a
reduced value. On the other hand, the land can be purchased incrementally
over the period. Approximately 100 hectares (250 acres) are needed per year.
At $12,350/hectare, the annual acquisition cost of land would be $1.2 by 10 .
Again, the land may have the same or reduced value at the end of the period.
Last, the land may be considered as part of the conventional costs of
operating the mine and not attributable only to potentially hazardous waste
disposal. In this case, the value of the land is jiot relevant to the cost
calculation and should be omitted.
219
-------
The answers obtained using these alternative approaches are quite different.
The difference occurs because time is being discounted at a 10 percent rate.
The present value of alternative capital cost outlays will therefore vary
significantly. To give the broadest meaning to the results, the cost of waste
disposal is calculated using each of the alternative capital cost outlay
patterns. These alternatives "cases" are carried throughout the analysis.
Other major cost assumptions are presented in Table 88.
The land cost calculations presented above and the other assumptions in
Table 88 were used to estimate the cost of the surface waste dump for the
disposal of uranium. Table 89 presents the results. The dollar costs in
Table 89 are expressed as annual amounts, except where otherwise specified,
expressed in fourth quarter (Q4) 1973 dollars. Table 89 indicates that the
cost per metric ton of potentially hazardous waste disposal using a surface
waste dump (Level I technology) can range from $0.23 to $0.003, depending
on which method is used for evaluating land costs. Total industry costs
associated with Level I technology would be $0.5 million to $8,000 or
less than 0.5 percent of the total value added in mining uranium ores.
Waste Treatment Cost—Mine Backfill, Level I Technology. In addition to
surface waste dumping, potentially hazardous waste materials can be disposed
by mine backfilling. Level I technology for mine backfilling is a process
which does not separate overburden and waste rock. This rock is temporarily
piled relatively close to the pit and later trucked back to the edge. The
material is contoured, and natural vegetation occurs. The cost of potentially
hazardous waste disposal by mine backfilling (Level I technology) includes:
(1) land rental or purchase; (2) hauling the material back to the pit; (3)
contouring the surface.
As mentioned earlier, the representative open-pit uranium mine must dispose
of 1.94 by 10 MT of overburden per year of which 4.86 by 10 MT (25 percent)
is used as mine backfill. The mine also produces 4.257 by 10^ MT of potentially
o
hazardous waste rock per year, of which 30.7 by 10J MT (7 percent) are used as
backfill. The total amount of mine backfill material equals 4.89 by 10 MT/year.
Assuming a bulk density of 1.96 MT/m3 (1.5 MT/yd3), 2.49 by 106 m3/year will be
piled for later mine backfill. If the pile is 7.62 m deep (25 ft), the pile will
cover 3.27 by 105 m2/year (80.8 acres) or 32.7 hectares/year. We will further
assume that the temporary pile is allowed to accumulate for 5 years before
backfilling begins. Once backfilling begins^ it is assumed to proceed at a rate
greater than the rate that material is being added. In other words, the maximum
land area needed is 163.5 hectares. At a price of $12,350/hectare ($5,000/acre)
in 1973 dollars, the initial land cost would be $2.C2 by 10 . Reselling the
land after 20 years would yield a total rental cost of $1.69 by 10 . Applying
the capital recovery factor (20 years at 10 percent) yields a levelized land
rent of $1.98 by 105/year.
220
-------
TABLE 88
MAJOR ASSUMPTIONS OF COST ANALYSIS, OPEN-PIT URANIUM MINE-
SURFACE WASTE DUMP—TECHNOLOGY LEVEL I
I. Capital Cost (other than land)
One bulldozer is assumed to work full-time at surface waste dump
a. Bulldozer (Cat-977) Capital cost = $56,000*
Useful life = 5 yearst
b. Present value calculations
Year purchased
or replaced
1
6
11
16
Real dollar
cost (1973)
56,000
56,000
56,000
56,000
Present value
factor*
1.0
0.621
0.386
0.239
, .v'4--
Present value v" . -
,- •> -'i * CJ'' '
of cost . ••;«;:?,:-
. -. .;;•.! ^o><-'
$56,000 SVVV'
34,776
21,616
13,384 ,_:M
Total $125,776
Capital recovery factor (20 years at 10%) = 0.117
Levelized annual capital cost of bulldozer = $14,716
II. Labor Cost
One bulldozer operator required at $8.97/hr (1973 dollars, 40)§
Labor cost = $8.97 x 8 hr/day x 260 days/year
= $18,658/year
III. Supervision
Assumed to be 25% of labor cost**
Supervision = 0.25 x $18,658/year = $4,664
IV. Maintenance
Assumed to equal 6% of nominal cost of bulldozer per yeartt
Maintenance = 0.06 x $56,000 = $3,360/year
221
-------
TABLE 88 (CONCLUDED)
V. Insurance and Taxes
Assumed to equal 2% of nominal capital cost (land and equipmen
Insurance and taxes = 0.02 x ($24.05 x 106 + 56.0 x 103) = $4.82 x 105
VI. Energy and Power
Bulldozer assumed to use 60 gal. diesel fuel per day^t
Energy plus power cost = 260 days/year x 60 gal/day x $0.50/gal$*
= $7,800/year
* U.S. Environmental Protection Agency. Assessment of industrial haz-
ardous waste practices, inorganic chemical industry. Contract No. 68-01-
2246, 1975. p. 7-33.
t U.S. Environmental Protection Agency. Assessment of industrial haz-
ardous waste practices, inorganic chemical industry. Contract No. 68-01-
2246, 1975. p. 7-33.
$ Assuming a 10% discount rate (r), the present value factor equals
i , i where i - the relevant year 1 in column 1.
§ U.S. Environmental Protection Agency. Assessment of industrial haz-
ardous waste practices, inorganic chemical industry. Contract No. 68-01-
2246, 1975. p. 7-5 and C-2.
** U.S. Environmental Protection Agency. Assessment of industrial haz-
ardous waste practices, inorganic chemical industry. Contract No. 68-01-
2246, 1975. p. 7-34.
tt Gruber, G. I. Assessment of industrial hazardous waste practices,
organic chemical pesticides and explosives industries. U.S. Environmental
Protection Agency, Contract No. 68-01-2919, January 1976. p. 7-2.
$$ U.S. Environmental Protection Agency. Assessment of industrial haz-
ardous waste practices, inorganic chemical industry. Contract No. 68-01-
2246, 1975. p. 7-34.
222
-------
. TABLE 89
DISPOSAL COSTS--POTENTIALLY HAZARDOUS WASTE FROM OPEN-PIT URANIUM
MINING--SURFACE WASTE DUMP, TECHNOLOGY LEVEL I
(EXPRESSED IN ANNUAL COSTS--1973 (Q4) DOLLARS)
Capital
Labor
Supervision
Maintenance
Materials
Insurance + taxes
Energy -1- power
Total cost of surface dump
Case "A"*
$2.83 x 106
$18,658
$4,664
$3,360
None
S4.82 x 105
$7,800
$3.34 x 105
Case "B"t
$2.36 x 106
$18,658
$4,664
$3,360
None
$4.82 x 105
$7,800
$2.87 x 106
Case "C"$
$1.2 x 106
$18,658
$4,664
$3,360
None
$2.4 x 104
$7,800
$1.25 x 106
Case "D
$1.47 x
$18,658
$4,664
$3,360
None
1,120
$7,800
$5.03 x
"§
104
104
Percent of metric tons of
hazardous waste rock
(in total waste) 2.7 2.7 2.7 2.7
Cost of hazardous waste
disposal 90,360 77,650 33,800 1,360
Metric tons of potentially
hazardous waste to sur-
face dump each year 395 x 103 395 x 103 395 x 103 395 x 103
Cost per metric ton of
potentially hazardous
waste in surface dump
per year $0.23 $0.20 $0.09 $0.003
Total cost if entire
industry adopted
($106/yr) $0.47 ' - - $0.008
*
Total cost as a percent
of value added in mining 0.29% ' 0.07.
* Case "A" assumes all land purchased in Year 1 and has 0 value after 20 years.
t Case "B" assumes all land is purchased in Year l' and is resold at the same value
after 20 years.
* Case "C" assumes land is purchased yearly (247 acres) and has 0 value after 20 years
§ Case "D" assumes land would be purchased even if no waste disposal was required
and, therefore, land costs do not enter cost calculation.
223
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Other capital equipment for mine backfilling includes bulldozers, front
loaders, and trucks. The total amount of material to be hauled back to the
open pit will be 4.98 by 107 m (i.e., 2.49 by 106 m3/year by 20 years). If
mine backfilling begins after 5 years and continues until the end of the
period of analysis (20 years), approximately 3.32 by 10 m will have to be
moved per year. We will assume 77-MT capacity (85 tons or 39.31 ttr) trucks
will be used, and a round trip from the temporary pile to the edge of the
pit takes 20 min. Therefore, 4.5 trucks will be needed on a full-time basis
(24 hr/day, 260 days/year). In addition, one 7.6-m3 (10 yd ) front loader
would be required full time. Finally, one bulldozer would be required 33
percent of the time. The cost and useful life of this equipment is shown in
Table 90. Table 90 also presents the labor, operation, maintenance,.taxes,
and energy costs assumptions for the mine backfilling operation.
The assumption presented in Table 90 and the-land costs discussed above
were used to generate the estimates of the cost of disposal of potentially
hazardous uranium wastes through mine backfilling (Table 91). The four
cases of land valuation used in Table 91 are similar to the four cases
presented in the discussion of disposal using surface waste dumps. However,
Case "C" is slightly different. Because the mine backfill dump is only
temporary, material will only accumulate for 5 years. For the next 15
years, the material is assumed to be backfilled to the mine at a faster
rate than it is being placed on the pile. In Case "C", 32.7 hectares are
assumed to be purchased each year for 5 years. The land is then held for
15 additional years and is assumed to have no resale value.
The results of the cost analysis show that the valuation of land is less
important to the cost of mine backfill than in the case of surface waste
dumping. The other capital equipment (trucks, loaders, and bulldozer) is a
much more important cost component in the mine backfill operation. The
associated labor supervision, maintenance, and energy also become more
important. Table 91 shows that the cost per metric ton of potentially
hazardous waste in mine backfilling is between $0.21 and $0.16 for the
representative mine (Level I technology). These costs expressed as total
cost to the industry represent $0.023 to $0.030 million per year or less
than 0.5 percent of the value added in mining.
Waste Treatment Costs—Open-Pit Uranium Mine Waste, Levels II and III Technology.
Surface Waste Dump Disposal Costs—Levels II and III Technology. Disposal
Costs associated with Levels II and III technology differ from Level I
technology in two important ways. First, the potentially hazardous waste
rock from open-pit uranium mining is separated from overburden in Levels
II and III technology. Therefore, a much smaller amount of material is
deposited on the potentially hazardous waste dump. Second, the waste dump
is contoured, covered with top soil (4 or 12 in. in the analysis), seeded,
and fertilized, if necessary. The major cost assumptions used in the
calculation are presented in Table 92.
224
-------
TABLE 90
ASSUMPTIONS OF COST ANALYSIS, OPEN-PIT URANIUM MINE-
MINE BACKFILLING--TECHNOLOGY LEVEL I
I. Capital Cost (except land)
A. Trucks
77.1 MT (85-ton) trucks are assumed to be used
Temporary pile is 1.6 km (1 mile) from edge of pit.
Hauling time equals 5 min to load, 6 min in transit (loaded) 3 min
to dump, 4 min back to pile (unloaded) or % 20 min/trip
Each truck hauls 2,832 cu meters/day (24 hr)
3.32 x 10 cu meters (4.34 x 10^ cu yd) must be moved each year.
3.32 x 10^/year-s. 260 working days/year = 1.28 x 10 cu meters/day
Therefore, 4.5 trucks are required full time (24 hr/day and 260 days/year)
85-ton truck cost (in 1973 dollars) = $47,000*
Useful life = 5 yearst
B. Front loaders
One caterpillar (Cat-992) with 7.6 cu meter capacity (10 cu yards)
is required full time
Front end loader cost = $265,000 (1976 dollars)* or $203,000 (1973
dollars)§
Useful life = 5 years**
C. Bulldozer
From previous analysis it was assumed that one bulldozer (Cat-997)
could contour 1.49 x 10' MT/year of overburden and waste rock.
Therefore, the bulldozer is required 0.33 of the time for 4.89 x
10^ metric tons/year.
Bulldozer cost (Cat-977^ » $56,000 (1973 dollars)t
Useful life = 5 years*
225
-------
TABLE 90 (CONTINUED)
D. Present value calculation
Trucks - 4.5 at $47,000 = $2.115 x 105
Loader - 1.0 at $203,000 = $32.03 x 105
Bulldozer - 0.33 at $56,000 = $1.848 x 104
Total nominal cost (1973 dollars) = $4.330
Year purchased Nominal cost Present value
or repaired of equipment factor at 10%
1 $4.33 x
6 .$4.33 x
11 $4.33 x
16 $4.33 x
Total present value
105
105
105
105
1.0
0.621
0.388
0.239
x 105
Present value
of cost
$4.33 x
$2.69 x
$1.68 x
$1.03 x
$9.73 x
105
105
105
105
105
Capital recovery factor (20 years at 10%) = 0.117
Levelized annual cost of capital equipment = $1.14 x 10
II. Labor cost
Truck drivers and equipment operators—1973 wages assumed to be
$8.97/hr*
A. Labor cost (trucks) = $8.97/hr x 4.5 drivers x 24 hr/day x 260
days/year
= $251,878/year
B. Labor cost (loader) = $8.97/hr x 1 operator x 24 hr/day x 260
days/year
r $55,973/year
C. Labor cost (bulldozer) = $8.97/hr x 0.33 operators x 8 hr/day x
260 days/year
••=> $6,157/year
D. Total labor cost = $314,008/year
226
-------
TABLE 90 (CONTINUED)
III. Supervision
Assumed to be 25% of labor cost***
Supervision = 0.25 x $314,008/year = $78,502/year
IV. Maintenance
Assumed to equal 67» of nominal cost of equipment
Maintenance = 0.06/year x $4.33 x 105 = $25,980/year
V. Insurance and taxes
Assumed to equal 2%/year of nominal capital cost (land and equipment)
Insurance and taxes = 0.02/year x ($1.69 x 106 + $4.33 x 105) =
$4.2 x 104/year (Case B)
VI. Energy and power
A. Trucks
85-ton trucks assumed to consume 15 gal/hrt
Diesel fuel assumed to cost $0.50/gal (1973 dollars)*
Energy plus power (trucks) =4.5 trucks x 24 hr/day x 260 days/year x
15 gal/hr x $0.50/gal.
- $210,600/year
B. Loader
7.65 m^ (10 cu yd) front loader is assumed to consume 18 gal/hr§
Energy plus power (loader) = 1 loader x.24 hr/day x 260 days/year x
18 gal/hr x $0.50/gal
= $56,160/year
C. Bulldozer
1 bulldozer is assumed to use 60 gal/8 hr*
Energy plus power (bulldozer) - 0.33 bulldozers x 60 gal/8 hr x 260 days/
year x $0,50/gal.
= $2,574/year
227
-------
TABLE 90 . (CONCLUDED)
* U.S. Environmental Protection Agency. Assessment of industrial
hazardous waste practices, inorganic chemical industry. Contract No.
68-01-2246, 1975. p. 7-33. Estimates were obtained for the 10-ton truck
and the "0.6 power factor" was used to estimate the cost of the 85-ton
truck.
t U.S. Environmental Protection Agency. Assessment of industrial
hazardous waste practices, inorganic chemical industry. Contract No.
68-01-2246, 1975, p. 7-33. Estimates were obtained for the 10-ton truck
and the "0.6 power factor" was used to estimate the cost of the 85-ton
truck.
$ MRI communication with Caterpillar Corporation, Kansas City, Missouri,
June 3, 1976.
§ Using M+S Equipment Cost Index, Chemical Engineering. May 24, 1976
and February 18, 1974.
** U.S. Environmental Protection Agency. Assessment of industrial
hazardous waste practices, inorganic chemical industry. Contract No.
68-01-2246, 1975, p. 7-33. Estimates were obtained for the 10-ton truck
and the "0.6 power factor" was used to estimate the cost of the 85-ton
truck.
228
-------
TABLE 91
DISPOSAL COSTS— POTENTIALLY HAZARDOUS WASTE FROM OPEN-PIT URANIUM
MINING— MINE BACKFILLING, TECHNOLOGY LEVEL I
(EXPRESSED AS ANNUAL 1973 (Q4) DOLLARS)
Case "A"* Case "B"t Case "C"j Case "D"§
Capital
Land
Equipment
Labor
Supervision
Maintenance
Insurance plus taxes
Energy and power
$2.36 x 105
$1.14 x 105
$3.14 x 105
$7.85 x 104
$2.60 x 104
$4.90 x 104
S2.70 x 105
$1.98 x 105 $1.87 x 105
$1.14 x 105 $1.14 x 105 $1.14 x 105
$3.14 x 105 $3.14 x 105 $3.14 x 105
$7.85 x 104 $7.85 x 104 $7.85 x 104
$2.60 x 104 $2.60 x 104 $2.60 x 104
$4.20 x 104 $4.20 x 104 $8.66 x 103
$2.70 x 1Q5 $2.70 x 105 $2.70 x 105
Total cost of mine backfilling $1.09 x 106 $1.01 x 106 $1.03 x 106 $8.10 x 105
Ratio of potentially hazardous
waste rock to overburden
0.6%
Cost of hazardous waste disposal $6.5 x 103
0.6% 0.6% 0.6%
$6.2 x 103 $6.2 x 103 $4.9 x 103
Metric tons of potentially
hazardous waste to mine
backfill per year
30.7 x 10
Cost per metric ton of poten-
tially hazardous waste in
mine backfill $0.21
Total cost if entire industry
adopted ($106/yr) $0.30
Total cost as percent of
value added 0.02%
30.7 x 10
$0.20
30.7 x 10
$0.20
30.7 x 10
$0.16
0.01%
* Case "A" assumes all land required is purchased in year 1 and. has no resale value
after 20 years.
t Case "B" assumes all land required is purchased in year 1 and has the same value
when resold after 20 years.
* Case "C" assumes 32.7 ha are bought per year for 5 years and then held for 15
years, and have 0 value after 20 years.
§ Case "D" assumes land has no rent or cost to the company because it is owned by
the firm for future mining (i.e., the land has no opportunity cost).
229
-------
TABLE 92
MAJOR COST ASSUMPTIONS: TECHNOLOGY LEVELS II AND III,
URANIUM OPEN-PIT MINE, SURFACE WASTE DUMP
Note: In technology Levels II and III, the potentially hazardous waste
rock is segregated from the overburden.
I. Land requirement--representative open-pit mine
425,700 MT (469,300 tons) of waste rock produced per year by
representative plant
395,000 MT (435,400 tons) of waste rock per year are deposited
in surface dump
2.01 x 10 cu m deposited at dump per year at bulk density of
1.96 MT/cu m (1.5 MT/cu yard)
After 20 years, 4.03 x 10 cu m will be deposited
Dump assumed to be 7.62 m deep (25 ft)
Therefore, 5.28 x 10^ sq m or 52.8 hectares (131 acres) are required
II. Capital cost
A. Land cost = 52.8 hectares (130.5 acres) at $12,350/hectare = $6.52x 105
Levelized annual cost (assuming land is purchased in year 1 and
has no value after 20 years) = $76,725
B. Bulldozer
Levelized annual cost of one bulldozer = $14,716*
Bulldozer assumed to be required 3% of time
Levelized annual cost of 3% of one bulldozer = $441/year
III. Materials and revegetation costs
All revegetation (i.e., topsbil cover, seeding and fertilizer) assumed
to be done when surface dump becomes inactive (i.e., after 20 years)
$742 - $1,4847
4 in. soil and seeding (includes materials and labor)
hectare ($300 - $600/acre)—
12 in. soil cover and seeding (includes materials and labor)
$4,325/hectare ($750 - $l,750/acre)li/
$1,854 -
4 in. soil
Total revegetation cost
(52.8 ha, 131 acres)
Present value (20 years at 10%)
Levelized annual costt
Low
$39,300'
6,445
754'
High
$78,600
12,890
1,508
12 in. soil
Low High
$98,250 $229,250
16,113
1,885
37,597
4,399
230
-------
TABLE 92 (CONCLUDED)
IV. Labor cost
Bulldozer operator = $18,658* x 0.03 = $560/year
V. Supervis ion
Assumed to be 25% of labor cost or $140/year
VI. Maintenance
Bulldozer maintenance = $3,360/year* x 0.03 = $100/year
Maintenance on revegetated land is minimal
VII. Energy and power
Bulldozer energy and power costs = $7,800/year* x 0.03 = $234/year
VIII. Insurance and taxes
39 /
Assumed to. equal 2% of nominal capital cost (land and equipment)—
Insurances and taxes = 0.02 x ($6.52 x 105 + $1.7 x 103) = $1.3 x 104
* See Table 89, p. 223.
f Capital recovery factor of 0.117.
231
-------
The disposal costs of potentially hazardous waste from open-pit uranium
mines in surface waste dumps are presented in Table 93. A low and high cost
estimate is presented for 4 and 12 in. of top soil covering. Fertilizer
costs are not included. Fertilizer costs would not alter the cost per metric
oo/ r
ton estimate because these costs are relatively smallr^' The table indicates
that the different cost estimates of revegetation have little impact on the
disposal cost per metric ton. Differences in the depth of the top soil only
change the cost per metric ton by $0.01.
Added Costs for Protective Cover on Active Surface Waste Dumps—Level
I/II Technology. The preceding cost analysis was based on a 20-year life for
a dump, and with stabilization being affected at the end of the 20-year
period when the dump becomes inactive. If overburden is available for •
periodic emplacement on potentially hazardous waste rock, no added expense
should be incurred. If, however, the cover material must be generated on-
site, added costs will be incurred for bulldozing adjacent soil materials
and emplacing them periodically over exposed potentially hazardous wastes.
A variety of schedules and dump construction procedures may in principal
be selected. The following cost analysis is based on the assumption that a
protective layer 0.3-m (1-ft) thick is spread over an annual accumulation of
waste rock. It is further assumed that one-fourth of the area needed for a
20-year period is active at any one time. The assumptions are summarized
below.
Total area required for 20 years - 52.8 hectares (131 acres)
Area active in a given year - 13.2 hectares (32.7 acres)
Waste rock per year - 395,000 MT
' • *}
Waste rock volume per year <- 201,000 nr
Waste rock depth per year - 1.52 m (5 ft)
Cover depth per year - 0.305 m (1 ft)
Cost of annual cover per unit area - $3,950/hectare ($l,600/acre)
Since the costs will be incurred annually, no levelization is necessary.
The added costs are thus calculated as follows*
232
-------
TABLE 93
DISPOSAL COSTS - POTENTIALLY HAZARDOUS WASTES FROM OPEN-PIT URANIUM
MINING - SURFACE WASTE DUMP,. TECHNOLOGY LEVELS II AND III
(EXPRESSED IN EQUIVALENT ANNUAL 1973 Q4 DOLLARS).
Assuming 4 in. of Topsoil on Dump
Low Estimate High Estimate
Capital (land and equipment)
Materials and Revegetation
Labor
Supervision
Maintenance
Insurance and Taxes
Energy and Power
$76,725-^ (63, 773) ^
754
560
140
100
13,000
234
$76,725
1,508
560
140
100
13,000
234
Assuming 12 in. of Topsoil on Dump
Low Estimate
$76,725
1,885
560
140
100
13,000
234
High Estirate
$76,725
4,399
560
140
100
13,000
234
Total Cost of Surface
Dump (Technologies II and III)
$91.513 (78,561)
$95,158
Metric tons of potentially
hazardous waste to .
surface dump each year 395x10
Cost per metric ton of
potentially .hazardous waste $0.23 (0.20)
Total cost if entire industry
adopted ($106/yr) $0.38
Total cost as a percent of
value added 0.237.
395xl03 395xl03 395xl03
$0.23 $0.24 $0.24
$0.45
0 . 28%
a/ This value assumes the land is all purchased in the first year and has no resale value after 20 years.
b/ This estimate, in parenthesis, assumes the land is purchased in the first year and retains 1007, of its value wher
resold after 20 years.
-------
Total costs for annual addition of cover = 13.2 hectares by $3,950/
hectare = $52,140
Cost per metric ton of potentially hazardous waste = $52,140 -r 395,000
MT = $0.13/MT
The costs will decrease in direct proportion to a decrease in the fraction
of the total dump which is active at a given time and will increase in proportion
to a shortening of the period between applications of cover. Thus, the cost will
be $0.26/MT if one-fourth of the dump is active in a given interval and the
surface is covered semiannually. Semiannual coverage of one-eighth of the total
dump area will cost $0.13/MT.
The added cost of annual coverage of one-fourth of the total dump area
will increase Level II/III costs estimated in Table 93 by about 55 percent
and by about 57 percent of Level I costs. The total cost, if the entire
industry used this technology would vary from $0.38 to $0.45 million per
year and this would amount to less than 0.5 percent of the value added in
mining.
Disposal Costs--Underground Uranium Mining, Surface Waste Dump.
The rationale and assumptions used to develop a representative model for
underground mining operations are discussed on page 175 under "Mass Balances
of Materials." The mass balance data for a representative model and for waste
treatment and disposal practices of underground uranium mining operations are
given in Figures 25 (p. 187), 30 (p. 210), and 31 (p. 212).
Level I Technology. The disposal of potentially hazardous waste from
underground uranium mining is quite similar to open-pit uranium mining
disposal techniques. Potentially hazardous waste rock is deposited on a
surface dump, contoured (under Level I technology), and natural vegetation
is allowed to occur. Major assumptions used in the calculation of these
costs are presented in Table 94. Many of the assumptions used in the open-
pit mine waste disposal calculations are also used in the analysis of
underground mine waste disposal costs.
The results of the analysis of Level I technology are presented in
Table 95. The cost per metric ton of potentially hazardous waste treated
range from $0.25 to $0.04, depending on how land is acquired and valued.
Total cost if the entire industry adopted the technology level would not
exceed $50,000/year. The total cost would be less than 0.05 percent of the
value added.
234
-------
TABLE 94
MAJOR COST ASSUMPTIONS--UNDERGROUND URANIUM MINING SURFACE
WASTE DUMP--TECHNOLOGY LEVEL I
I. Land requirements
51,000 MT/year (56,200 tons/vr) of potentially hazardous waste rock
are deposited on dump in the representative plant
26,020 cu m/year (34,000 cu yd/yr) are deposited assuming a bulk
density of 1.96 MT/cu m
Over 20 years, 520,408 cu m are deposited
The dump is assumed to be 7.62 m deep (25 ft)
Therefore, land requirements equal 68,295 sq m or 6.8 hectares (16.8 acres)
II. Capital costs
A. Land purchase price at $12,350/hectares ($5,000/acre) = $83,980
B. Other capital costs
Note: The land area required for the dump from the underground
mine is approximately 1% of the open-pit dump. Therefore,
costs will be scaled down from the open-pit calculations pre-
sented earlier.
Bulldozer: $14,716/year4-0-/ x 0.01 = $147/year
III. Labor cost
$18,658^°-/ x 0.01 = $186/year
IV. Supervision
$4,664—/ x 0.01 = $47/year
V. Maintenance
$3,3604-°-/ x 0.01 = $34/year
VI. Insurance and taxes
Assumed to equal 270 of nominal capital cost
VII. Energy and power
$7,800^Py x 0.01 = $78/year
235
-------
TABLE 95
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE FROM UNDERGROUND
URANIUM MINING - SURFACE WASTE DUMP, LEVEL I TECHNOLOGY
(EXPRESSED IN ANNUAL EQUIVALENT 1973 DOLLARS)
Capital (land and equipment)
Labor
Supervision
Maintenance
Materials
Insurance and taxes
Energy and power
Case "A"*
$ 9,973
186
47
34
None
1,705
78
Case "B"t
$ 8,361
186
47
34
None
1,705
78
Case "C"t
$ 4,364
186
47
34
None
1,705
78
Case "D"§
$ 147
186
47
34
None
1,705
78
Total cost $12,023 $10,411 $ 6,414 $ 2,197
Metric tons of potentially 51,000 51,000 51,000 51,000
hazardous waste to surface
dump per year
Cost per metric ton of $ 0.25 $ 0.20 $ 0.13 $ 0.04
potentially hazardous waste
in surface dump per year
Total cost if entire industry $ 0.05 - $ 0.008
adopted ($106/yr)
Total cost as a percent of 0.03% - - 0%
value added
* Assumes all land purchased in year 1 and has 0 value after 20 years.
t Assumes all land is purchased in year 1 and is resold at the same
value after 20 years.
$ Assumes land is purchased yearly (247 acres) and has 0 value after
20 years. , .
§ Assumes land would be purchased even if no waste disposal was
required, and therefore, land costs do not enter cost calculation.
236
-------
Levels II and III Technology. Disposal of potentially hazardous wastes
using Level II technology is the same as Level III technology. The surface
dump is contoured, covered with top soil, seeded, and fertilized, if necessary.
In New Mexico, where most underground uranium mines are located, rock covering
may be less costly than top soil and seeding. Costs for both Levels II and III
technology are presented in Table 96. Costs are presented for four alternative
types and depth of dump covering (i.e., 4 and 12 in. of top soil and seeding,
6 and 12 in. of rock covering). Costs range from $0.31 to $0.06/MT of
potentially hazardous waste, depending on how land costs are handled. The cost
to the industry as a whole would be $0.038 to $0.088 million per year, assuming
all underground uranium mines used these technology levels. This cost represents
less than 0.5 percent of the value added in uranium mining.
Waste Treatment Costs - Wastes from Acid and Alkaline Leaching Uranium
Concentrators. Disposal of potentially hazardous wastes occurring in the
concentration of uranium can be accomplished in tailings ponds. Level I
technology requires the construction of an earthdike tailings pond which
allows zero discharge of wastewater. The tailings pond is assumed to be
within 914 m (3,000 ft) of the model plant/Natural vegetation is allowed
on the dam and tailings beaches.
The model plant used in the following analysis is taken from a recent
study by M. B. Sears.—' Much of the plant description and capital cost data
used below are drawn from the Sears study. The rationale used in selecting the
model plants is discussed on page 213. The mass balance data for.the model
plants are given in Figure 24 (p. 181) and Figure 25 (p. 187).
The tailings pond of the model plant is sited within 914 m (3,000 ft)
of the mill. It is near the upper reaches of a gently sloping natural
drainage area and at least 61 m (200 ft) from any surface stream or
permeable formation. A starter dam is constructed of native soil across
the lower end of the site. The remainder of the dam is built from the
tailings. The tailings are hydraulically classified by hydroclones or
settling. The maximum height of the dam is assumed to be 30 m (100 ft),
including 1.5 m (5 ft) of freeboard.— A minimum 0.8 hectares (2 acres)
beach area of dry tailings sand is maintained along the dam to protect
against wave action and dam instability due to liquefaction. The face
of the dam contains exposed tailings. The tailings pond is designed to
keep the maximum area of tailings wet. A 20 percent contingency of land
is included for abnormal weather conditions. Under Level I technology,
the beaches and tailings dam are covered with natural vegetation.
The capital costs of the tailings pond (Level I technology) are illustrated
in Table 97. Both acid and alkaline leach processes are presented for Wyoming
and New Mexico locations. Land area requirements vary because of the different
leach processes and different average evaporation rates in Wyoming and
New Mexico. The costs of the tailings pump and pipeline, earth dam, and
cyclone are also presented. Table 98 presents similar information on
operational costs of the pond.
237
-------
TABLE 96
DISPOSAL COSTS, UNDERGROUND URANIUM MINING
LEVELS II AND III TECHNOLOGY, SURFACE WASTE DUMP
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Capital
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Revegetation
10-cm (4-in.) topsoil and seed-
ing at $450/acrel?-/
30.5-cm (12-in. top soil and
seeding at -$l,250/acreii'
15-cm (6-in.) rock covering at
$l,000/acre— ''
30.5-cm (12-in.) rock covering
at $2,000/acreii/
Total
10-cm topsoil . .
30.5-cm topsoil
15-cm rock
30.5-cm rock
Metric tons of potentially
hazardous waste to dump per
year
Cost per metric ton -of
potentially hazardous waste in
surface dump per year
10-cm topsoil
30.5-cm topsoil
15-cm rock . -
30. 5 -cm rock
Total cost if industry adopted
($106/yr)
Total cost as a percent of value
added . .
Case "A"*
$ 9,973
186
47
34
1,705
. 78
884
2,457
1,966
3,931
$12,907
14,480
13,989
15,954
51,000
$ 0.25
0.28
0.27
0.31
0.088
0.05%
Case "n"t
$ 8,361
186
47
34
1,705
78
884
2,457
1,966
3,931
$11,295
12,868
12,377
14,342
51,000
$ 0.22
0.25
0.24
0.28
-
-
Case "C"t
$ 4,364
186
47
34
1,705
78
884
2,457
1,966
3,931
$ 7,298
8,871
8,380
10,345
51,000
$ 0.14
0.17
0.16
0.20
-
-
Case "D"6
$ 147
186
47
34
1,705
78
884
2,457
1,966
3,931
$ 3,081
4,654
4,163
6,128
51,000
$ 0.06
0.09
0.08
0.12
0.038
0.02%
* Case "A" assumes all land purchased in year 1 ?rid has 0 value after 20 years.
•f Case "B" assumes all land is purchased in year'l and is resold at the same
value after 20 years.
$ Case "C" assumes land is purchased yearly (2-^7 acres) and has 0 value after
20 years. ,
8 Case "D" assumes land would be purchased even if no waste disposal was
required, and therefore, land costs do not enter cost calculation.
238
-------
NJ
OJ
VO
TABLE 97
t
CAPITAL COST ASSUMPTIONS FOR URANIUM TAILINGS PONDS
(LEVEL I TECHNOLOGY) IN WYOMING AND NEW MEXICO
(EXPRESSED IN 1973 DOLLARS)
Wyoming
New Mexico
197
Land area in hectares (acres) —
Land cost at $12,350/hectare
($5,000/acre) '
Tailings pump and pipeliner^-
Earth dami^/
197
Cyclone installation —
Total capital cost (Case "A")*
Levelized annual cost (A)t
Total capital cost (Case "B")$
Levelized annual cost (B)t
Acid leach
($)
70 (173)
870,000
250,000
63,000
58,000
1,241,000
145,000
1,098,000
128,000
Alkaline leach
($)
44 (109)
540,000
156,000
40,000
58,000
794,000
93,000
705,000
83,000
Acid leach
($)
39 (96)
480,000
138,000
35,000
58,000
711,000
83,000
632,000
74,000
Alkaline leach
($)
24 (60)
300,000
134,000
33,000
58,000
525,000
61,000
476,000
56,000
* Case "A" assumes the land has 0 value after 20 years of; analysis.
t Assuming a 10% discount rate and 20 years of analysis. The capital recovery factor is 0.117,
$ Case "B" assumes the land is resold after 20 years at ttie same value as the purchase price.
-------
TABLE 98
OPERATING COST ASSUMPTIONS - TAILINGS POND (LEVEL I TECHNOLOGY)
40/
I. Labor cost:—
One operator required on cyclone separators for 5 months, 8 hr/day at
$8.97/hr .
Labor = $7,176/year
II. Supervision
40/
Assumed to be 257. of labor cost—
Supervision = $l,794/year
III. Maintenance
40/
Assumed to be equal to 67. of nominal cost of equipment per year—
Wyoming New Mexico
Cost of tailings pump,
pipeline, and cyclone
Total maintenance cost
per year
IV. Insurance and taxes
40/
Assumed to equal 27»/year of total nominal capital cost (land and equipment—
Insurance and taxes = 0.02/year by ($1.2 x 106) = $24,000 (Wyoming, acid-
leach, Case "A")
V. Energy and Power
A. Tailings pump assumed to consume electricity with a value of
approximately $500/year~'
B. Cyclone separator mounted on truck—only fuel required is for
truck as it travels around the tailings pond; assumed to be less
than $100/year
240
Acid
JS1
308,000
18,000
Alkaline
($)
214,000
13,000 '
Acid
JS1
196,000
12,000 •
Alkaline
($)
192,000
12,000
-------
The inputs from Tables 97 and 98 were combined to obtain the results shown
in Tables 99 and 100. Table 99 presents the total cost of building and
operating the model tailings pond in Wyoming. Costs are shown for both the
acid and alkaline leach processes. The table shows that the tailings pond
requirements of the alkaline leaching process are approximately $0.10 cheaper
per metric ton of handled waste than the tailings pond associated with the
acid leaching process in Wyoming.
Table 100 shows the cost of disposing of potentially hazardous wastes in
New Mexico. The costs per metric ton of potentially hazardous wastes are
significantly lower (using Level I technology) in New Mexico than in Wyoming.
This cost reduction occurs because less land area is required in New Mexico.
Tables 99 and 100 also present the cost to the industry as a whole, and their
relative magnitude as a percent of value added.
Disposal Costs--Uranium Tailings Pond, Levels II and III Technology.
Uranium tailings pond disposal costs are similar for all three levels of
technology. The only difference is that revegetation of the tailings dam
surface and dried exposed tailings (tailings beaches) is carried out in
Levels II and III technology. The surface area of the dam and the amount of
exposed dry tailings depends on the location of the pond and the type of
concentrator used. Table 101 presents the assumed areas used in the analysis.
The analysis further assumes that the area on the face of the dam is covered
with soil and seeded in the 1st year of the analysis. The dry beach area is
assumed, to be covered in year 16. (Therefore, it is discounted by 0.239.)
Using the land areas in Table 101, the major cost components were derived
(see Table 97, p. 239:
The results for Wyoming and New Mexico are presented in Tables 99 and 100,
respectively. In Wyoming estimates are presented for 4 and 12 in. of topsoil
cover. In New Mexico, rock covering of 6 and 12 in. is also estimated.
241
-------
TABLE 99
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE FROM OPEN-PIT URANIUM
MINING--ACID LEACH AND ALKALINE LEACH CONCENTRATORS--
LEVEL I TECHNOLOGY IN WYOMING
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Capital (levelized cost)
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Acid
Case "A11*
$145,000
7,176
1,794
18,000
24,000
600
leach
Case "B"t
$128,000
7,176
1,794
18,000
22,000
600
Alkaline
Case "A11* .
$ 93,000
7,176
1,794
13,000
16,000
600
leach
Case "B't
$83,000
7,176
1,794
13,000
14,570
600
Total cost
$196,570 $177,570 $131,570 $119,570
Metric tons of potentially hazardous 662,200 662,200 662,200 662,200
tailings deposited in pond annually
Cost per metric ton of potentially
hazardous waste
Total cost if entire industry
adopted ($106/yr)
Total cost as a percent of value
added
$ 0.30 $ 0.27 $ 0.20 $ 0.18
0.84 - - $ 0.50
0.5% • - - 0.3%
* Case "A" assumes the land is all purchased in the initial year and has no
value after 20 years. .
t Case "B" assumes the land is all purchased in the initial year but has the
same dollar value after 20 years of use.
242
-------
TABLE 100
DISPOSAL COSTS, POTENTIALLY HAZARDOUS WASTE FROM OPEN-PIT URANIUM
MINING—ACID LEACH AND ALKALINE LEACH CONCENTRATORS,
LEVEL I TECHNOLOGY IN NEW MEXICO
.(IN ANNUAL 1973 DOLLARS)
Capital (levelized cost)
Labor
Supervision
Maintenance
Insurance and taxes
Energy and power
Acid
Case "A"*
$ 83,000
7,176
1,794
12,000
14,000
600
leach
Case "B"t
$ 74,000
7,176
1,794
12 , 000
13,000
600
Alkaline
Case "A"*
$ 61,000
7,176
1,794
12,000
11,000
600
leach
Case "B"t
$ 56,000
7,176
1,794
12,000
10,000
600
Total cost
$118,570 $108,570 $ 93,570 $ 87,570
Metric tons of potentially hazardous 662,200 662,200 662,200 662,200
tailings deposited in pond annually
Cost per metric ton of potentially $ 0.18 $ 0.16 $ 0.14 $ 0.13
hazardous waste
Total cost if entire industry adopted $ 0.51
($106/yr)
Total cost as a percent of value
added
0.32%
$ 0.36
0.27=
* Case "A" assumes the land is all purchased in the initial year and has no
value after 20 years.
t Case "B" assumes the land is all purchased in the initial year but has the
same dollar value after 20 years of use.
243
-------
TABLE 101
LAND AREA OF URANIUM TAILINGS POND DAM AND EXPOSED BEACHES
(EXPRESSED IN HECTARES (ACRES))
Average area of dry
beach near end of Area of exposed tailings
20-year life of mill on face of dam
New Mexico
Acid leach 14.6 (36) 4.9 (12)
Alkaline leach 26.7 (66) 4.9 (12)
Wyoming
Acid leach 0.8 (2) 4.0 (10)
Alkaline leach 10.1 (25). 4.9 (12)
Source: Reference 19.
244
-------
REFERENCES
1. Brobst, D. A., and W. P. Pratt. United States Mineral Resources, U.S.
Department of the Interior, Geological Survey Professional Paper 820,
U.S. Government Printing Office, Washington, B.C., 1973.
2. Decarlo, J. A., and C. E. Shortt. Uranium. Mineral Facts & Problems,
Bulletin 650, p.. 221, 1970 Edition.
3. U.S. Bureau of Mines, Preprint (Minor Metals) from the 1974 Minerals
Yearbook. 7-8 p.
4. U.S. Bureau of Mines, Statistical Summary, Preprint from the 1973
Minerals Yearbook. 3-28 p.
5. U.S. Bureau of Mines. Metals, minerals, and fuels. 1973 Minerals
Yearbook. 1:1263-1286.
6. Gordon, E. Uranium - prices firm up in '74; new foreign policies affect
investment. Engineering & Mining Journal, 176(3):213-215, Mar. 1975.
7. U.S. Department of the Interior, Bureau of Mines, Commodity Data
Summaries, 1974, Appendix I to Mining and Minerals Policy, 180-181 p.
8. Gordon E. Uranium - plant expansion should start now to meet projected
demand in 1975. Engineering & Mining Journal, 174(3):125-127, Mar.
1973.
9. Gordon, E. Uranium - new development is targeted at the future nuclear
generating market. Engineering & Mining Journal. 175(3):156-158, Mar.
1974.
10. U.S. Bureau of Mines. Metals, minerals, and fuels. 1972 Minerals
Yearbook. 1:1263-1265.
11. Gordon, E. Uranium - new projects anticipate coming demand. Engineering
& Mining Journal. 177(3):190-193, Mar. 1976.
12. U.S. Bureau of Mines. Metals, minerals, and fuels. 1969 Minerals
Yearbook. 1:1110-1112.
13. U.S. Bureau of Mines. Metals, minerals, and fuels. 1970 Minerals
Yearbook. 1:1140-1141.
14. U.S. Bureau of Mines. Metals, minerals, and fuels. 1971 Minerals
Yearbook. 1:1199-1201.
15 • Engineering & Mining Journal, 1973-1974. International Directory of
Mining and Mineral Processing Operations, Published by Engineering
& Mining Journal, New York, New York.
245
-------
16. U.S. Department of the Interior, Bureau of Mines. Mineral Facts ,& Problems,
Bulletin 650, 219-242 p., 1970 Edition.
17. Clark, D. A. State-of-the-art: uranium mining, milling, and refining
industry. U.S. Environmental Protection Agency, National Environmental
Research Center, Office of Research & Development. Corvallis, 'Oregon,
EPA-660/2-74-038, June 1974.
18. U.S. Environmental Protection Agency. Development document for interim
final and proposed effluent limitation's guidelines and new source
performance standards for the ore mining and dressing industry—point
source category, 1:174-176 & 314-341, Oct. 1975.
19, Sears, M. B.s et al. Correlation of radioactive waste treatment costs
and the environmental impact of waste effluents in the nuclear fuel
cycle for use in establishing as low as practicable guides—milling
of uranium ores. Oak Ridge National Laboratory, Oak Ridge, Tennessee,
ORNL-TM-4903, 1, May 1975.
20. Merritt, R. C. The extraction metallurgy of uranium. Colorado School of
Mines Research Institute, 1971.
21. Pearson, J. S, A sociological analysis of the reduction of hazardous
radiation in uranium mines, U,S« Department of Health, Education, and
Welfare, Public Health Service Center for Disease Control, National
Institute for Occupational Safety and Health, Salt Lake City, Utah,
! April 1975.
22. Kent, J. A. Riegels handbook of industrial chemistry. 7th ed. 743-748 p.
1974.
23. U.S. Department of the Interior, Bureau of Mines. Minerals Yearbook,
v.I. Minerals, metals and fuels. 1973.
24. U.S. Bureau of Mines, Commodity Data Summaries, 1974, Appendix I to
Mining and Minerals Policy, 3rd Annual Report of the Secretary of the
Interior under the Mining and Minerals Policy Act of 1970.
25. Summary report. Phase I study of inactive uranium mill sites and tailings
piles.
26. Gordon, E, Uranium - new projects anticipate coming demand. Engineering &
Mining Journal, 177(3):190-193, Mar. 1976.
27. U.S. Bureau of Mines, Metals, minerals, and fuels. 1973 Minerals Yearbook,
1:1263-1286.
246
-------
28. Sears, Mt B., et al. Correlation of radioactive waste treatment costs
and the environmental impact of waste effluents in the nuclear fuel
cycle for use in establishing 'as low as practicable1 guides—milling
of uranium ores. Contract No. W-7405:eng-26 for ERDA, ORNL-TM-4903,
v.I, UC-11-Environmental and Earth Sciences, Oak Ridge National
Laboratory, Oak Ridge, Tennessee, May 1975.
29. Clark, D. A. State-of-the-art: uranium mining, milling and refining
industry. U.S. Environmental Protection Agency, National Environmental
Research Center, Office of Research & Development. Corvallis,
Oregon. EPA-660/2-74-038, June 1974.
30. U.S. Environmental Protection Agency. 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, v.2, EPA 440/1-75/061, October 1975.
31. U.S. Bureau of Mines. Metals, minerals, and fuels. 1973 Minerals Yearbook.
v.l, 1263-1286 p.
32. U.S. Bureau of Mines. Commodity Data Summaries, 1974, Appendix I to Mining
and Minerals Policy, 3rd Annual Report of the Secretary of the Interior
under the Mining and Minerals Policy Act of 1970.
33. MRI communication from Exxon Corporation. Highlands Uranium Operations,
Wyoming, Nov. 12, 1974.
34. Sears, M. B., et al. Correlation of radioactive waste treatment costs
and the environmental impact of waste effluents in the nuclear fuel
cycle for use in establishing 'as low as practicable1 guides--milling
of uranium ores. Oak Ridge National Laboratory, Oak Ridge, Tennessee,
ORNL-TM-4903, v.l. May 1975.
35. Sears, M. B., et al. Correlation of radioactive waste treatment costs
and the environmental impact of waste effluents in the nuclear fuel
cycle for use in establishing 'as low as practicable1 guides—milling
of uranium ores. Oak Ridge National Laboratory, Oak Ridge, Tennessee,
ORNL-TM-4903, v.l. May 1975, p. 45.
36. MRI communication with uranium and vanadium mining and concentration
companies.
37. American Mining Congress questionnaires.
38. Unpublished company data.
39. Williams, R. Waste production and disposal in mining, milling, and
metallurgical industries. Miller Freeman Publications, San Francisco,
California. 1975. p. 431.
247
-------
40. MRI estimates of time. Labor cost per hour from Assessment of Industrial
Hazardous Wastes Practices, Inorganic Chemical Industry, U.S. Environmental
Protection Agency (Contract No. 68-01-2346).
248
-------
SECTION VI
MISCELLANEOUS METAL ORES (SIC 1099)
Various miscellaneous metals are produced in the United States in small
quantities by a few mines. The miscellaneous metals covered in this chapter
include antimony, beryllium, platinum-group metals (platinum, palladium,
rhodium, iridium, ruthenium, and osmium), rare earth, tin, titanium, and
zirconium*
In 1974 there were nine mines producing these metals (Figure 32). Total
production was estimated at 503,666 MT (554,000 tons). The greatest
concentration of tonnage of primary metal products was in the production of
zirconium and titanium ores from mines in Florida and New Jersey (Table 102).
Employment, according to the Engineering & Mining Journal, Mining Enforcement
and Safety Administration, unpublished company data and Midwest Research
Institute analyses, was about 1,328 workers. Most of these mines are small,
employing fewer than 100 employees (Tables 103 and 104). Mine production data,
where available in published commodity reports, for each of the metals are
presented in Tables 105 through 107. Production data for titanium, zirconium,
and beryllium were obtained for 1974 via questionnaires and visits. These
data are reported in later sections which treat these metals individually.
Ant imony
Industry characterization.
History of the Industry. Antimony is one of the oldest metals continuously
used by man. The natural sulfide of antimony was known and used by women in
Biblical times as medicine and as a cosmetic for eyebrow painting. By the 15th
century, antimony was used in alloys for printer's type, mirrors, and bells.
Precipitation of metal from the sulfide by iron is described by Ercker in
the 17th century, and in the 18th century roast-reduction procedures came into
use. The early 1830's marked the introduction of the reverberatory furnace
for smelting antimony; 1844, the French volatilization process, and 1896, the
first appearance of electrolytic antimony on the market.
249
-------
Platinum,
NS
<-n
o
Figure 32. Location of miscellaneous ores (SIC 1099),
-------
TABLE 102
U.S. PRODUCTION OF MISCELLANEOUS METALS - 1974
(RECOVERED METALS AND CONCENTRATES)*
State Metric tons EPA regions Metric tons
Alaska
California
Florida
Idaho
Montana
New Jersey
Utah
U.S. total
NA
NA
388,000
907
150
113,000
839
502,896
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
113,000
—
388,000
--
—
—
989
NA
907
Source: References 1 and 2.
NA = Not available.
* Includes antimony, beryllium, platinum, titanium, rare earth ores, tin,
and zirconium.
251
-------
TABLE 103
EMPLOYMENT IN ACTIVE MISCELLANEOUS MINING OPERATIONS (1974)
State
Employment
Source: References 1, 2, and 3.
EPA regions Employment
Alaska
California
Florida
Idaho
Montana
New Jersey
Utah
Virginia
U.S. total
37
93
300
700
14
80
63
41
1,328
Region I
Region II
Region III
Region IV
Region V
Region VI
Region VII
Region VIII
Region IX
Region X
__
80
41
300
—
—
—
77
93
737
252
-------
TABLE 104
NUMBER, LOCATION, AND SIZE OF MISCELLANEOUS MINING, AND MILLING OPERATIONS (SIC 1099)
(1974)
Employees
State 1-19 20-49 50-99 100-249 250-499 500-999 l,000-2.,499 2,500+
Alaska 1
California 1
Florida 2
Idaho 1
Montana 1
New Jersey 1
Utah 1
N> Virginia 1
£ U.S. total 2222 1
EPA regions
Region I
Region II 1
Region III 1
Region IV 2
Region V
Region VI
Region VII
Region VIII 2
Region IX 1
Region X 1 1
Total
operations
1
1
2
1
1
1
1
1
9
1
1
2
2
1
2
Total No.
of mines
1
1
2
1
1
1
1
1
9
1
1
2
2
1
2
No. of mills
1
1
2
1
1
1
1
1
9
1
1
2
2
1
2
Source: References 1, 2, and 3.
-------
TABLE 105
WORLD ANTIMONY PRODUCTION, 1969-1973
(METRIC TONS)*
Year
1969
1970
1971
1972
1973
United States
850
1,025
929
443
589
Rest of world
65,682
65,337
65,516
67,628
70,761
Total
66,532
66,362
66,445
68,071
71,350
Source: References 4, and 6.
* Data published in tons and calculated to metric tons,
TABLE 106
WORLD PRODUCTION OF PLATINUM-GROUP METALS, 1968-1973
(KILOGRAMS)*
Year
1968
1969
1970
1971
1972
1973
United States
467
684
529
560 '
529
622
Rest of world
105,099
106,032
131,319
126,467
142,952
160,618
Total
105,566
106,716
131,848
127,027
143,481
161,147
Source: References 5 and 6.
* Data published in thousand troy ounces and calculated to kilograms.
254
-------
TABLE 107
WORLD PRODUCTION OF PLATINUM, 1968-1972
(KILOGRAMS)*
Year United States Rest of world Total
1968
1969
1970
1971
1972
218
342
249
249
156
44,851
45,784
61,149
56,360
68,148
45,069
46,126
61,398
56,609
68,304
Source: Reference 5.
* Calculated from troy ounces to kilograms.
255
-------
Prior to the 20th century, chemical and metallurgical qualities of antimony
found relatively small use. After the turn of the century, consumption
increased rapidly as the world industrial pace quickened and the low-cost,
abundant supply of Chinese ore became available.
Expansion of the automobile industry and the use of storage batteries
stimulated consumption in the 1930fs and in turn escalated secondary recovery
of antimony. The escalated demand of the war years again reactivated idle
mines in the United States during World War II. Antimony oxide used in
flameproofing compounds for military use became the leading end product used
and necessitated rapid expansion of domestic oxide-smelter capacity. Demand
for consumer products kept antimonial lead consumption relatively high until
the post-Korean War adjustment in 1953 caused a decline in foreign production,
while domestic production was essentially terminated. Beginning in 1962,
increased consumption in the United States and other industrialized countries
resulted in a world supply deficit. Today the United States provides about
10 percent of its domestic requirements while importing the balance largely
from the Republic of South Africa, Mexico, and Bolivia.
Domestic Production and Capacities. Lead-silver ores of the Coeur d'Alene
district of Idaho contributed the majority of the 589 MT of primary antimony
mined in the United States in 1973. Sunshine Mining Company and U.S, Antimony
Corporation were the only domestic producers of antimony in the United States.
Tetrahedrite concentrates from Sunshine Mining Company, Hecla Mining Company,
and Silver Dollar Mining Company were processed into cathode metal, 96 percent
antimony, at the Sunshine Mining Company's electrolytic plant. Table 108 gives
the U.S. mine production of antimony by year.
The U.S, production capacity is much higher than the current production.
Seven hundred workers were engaged in domestic mine production in 1974. World
antimony production for the period 1969 to 1973 is shown in Table 105. Demand
for antimony is expected to increase at an annual rate of about 4 percent
through 1983. Since domestic output already is far below demand, antimony oxide
producers will have to increase their imports of finished oxide, crude oxide,
and ore.
By-Products and Coproducts. Most of the antimony from domestic sources is a
coproduct of mining, smelting, and refining of other metals and ores that
contain relatively small quantities of antimqny. Because of the close mineralogical
association of antimony with lead, antimony is extracted at primary lead
refineries.
256
-------
TABLE 108
PRIMARY MINE PRODUCTION OF ANTIMONY AND ANTIMONY CONCENTRATES IN THE
UNITED STATES BY YEAR
Antimony concentrates .,'.
Year Antimony (MT) (MT) .."••'•••'.,''•
1969
1970
1971
. 1972 '
1973
851
1,025
930
444
494
5,177
6,060
4,282
1,879
2,238
Source: References 4 and 6.
257
-------
Waste Generation and Characterization,
Waste Generation. Table 109 contains the national production data for antimony
mining and concentration operations in 1974 (i.e., the statistics shown for
Idaho and Montana apply for antimony only). In 1974, the total ore mined was
198,000 MT (218,000 tons), and the total antimony products amounted to 2,666 MT
(2,930 tons). The total mining waste, consisting of waste rock only, was
90,000 MT (99,208 tons), and the total concentrator waste was 189,000 MT
(208,000 tons).
Table 110 shows the ratios of waste rock and of tailings to ore mined for
antimony operations (all data shown for Idaho and Montana). In Idaho, the
ratio of waste rock to ore mined in 1974 was 0.47, and the ratio of
concentrator tailings to ore was 0.98. For Montana, the waste rock-to-ore
ratio was 0.26 and the tailings-to-ore ratio was 0.63.
Tables 111, 112, and 113 show the 1974 and projected 1977 and 1983 potentially
hazardous waste resulting from antimony mining and concentrating operations
in Idaho and Montana.
Underground Mining Process.
Description of Typical Process. A flow sheet showing an antimony, operation
is presented in Figure 33. The two operating antimony mines employ underground
mining methods similar to those used in the mining of lead-zinc. The scale of
operation is quite different, however, between the two mines; 86 percent (907
MT) of the antimony is produced at one mine.
Conventional underground mining methods are used, with one mine employing
stope and pillar methods and the other single bench, room, and pillar. One
mine has two primary shafts and about 12 operating levels ranging from 945
to 1,646 m (3,100-5,400 ft) underground. The ore is broken by drilling and
blasting operations. One mining method is called breast stoping with the stopes
being about 30 m (98 ft) wide on each side of a pillar. The mine stopes are
backfilled with sand and a cement cap. The second mine was opened in 1971 and
is still under development.
At the mine visited, ore is hauled underground by rail to a hoist that
lifts the ore to the surface, where it is sent to a crushing plant. Waste
rock, which is removed from the mine in the same manner as the ore, is hauled
by truck to a waste dump about 0.8 km (0.5 miles) from the mine.
i
Mass Balance of Materials. The mass balance data are shown in the table
on Figure 33. ' .
258
-------
TABLE 109
PRODUCTION STATISTICS BY STATE AND EPA REGION—MISCELLANEOUS ORES, SIC 1099, FOR 1974 (METRIC)
Wastes (10-VTPY)
State
New Jersey*
Florida*
Montana
Utah
California
Idaho
National
EPA Ore mined
Region (103 TPY)
II 7,000
IV 15,422
17
91
VIII 108
IX NA
X 181
22,711
Waste
rock
0
0
5
907
912
NA
85
997
Overburden
0
0
0
0
0
NA
0
0
Concentrator Total
Concentrator wet waste dry
dry wastes ( 103 TPY) waste
-
11
90
101
NA
178
279
-
. -
125
1.021
1,146
NA
2,178
3.324
-
• -
16
997
1,013
NA
263
1,276
Products (TPY)
Primary
concentrate
113.000
388,000
150
839
989
NA .
907
502,896
By-product
concentrate
0
125,000
0
0
0
NA
1.360
125,036
Miscel-
laneous
0
42,000
0
0
0
NA
249
42,976
Total
products
113,000
555,000
150
839
989
NA
2,516
671,505
Ratio of dry t.
Waste Waste to Total
to ore product waste
.
-
0.94 106
VI 1 . 188
9..4 1,024 79.0
NA NA NA
1.45 105 21.0
0.06 1.9
Metal
ore
Tl-Zr
Tl-Zr
Sb
Be
REt
Sb
Source: References 1 and 2.
N5 Note: - = No Information. • .
Jg NA = Not available.
* Approximately 961 of the material defined as ore Is returned directly to the dredge site and Is not counted as waste rock or concentrator wastes.
t RE - Rare earth group metals.
-------
S3
ON
O
TABLE 110
RATIO OF TOTAL WASTE ROCK-OVERBURDEN-CONCENTRATOR HASTE TO ORE MINED, SIC 1099,
MISCELLANEOUS ORES FOR 1974 (METRIC)*
State
New Jersey
Florida
Montana
Utah
California
Idaho
National
gpA Ore mined
Region (103 TPY)
II 7,000
IV 15,422
17
91
VIII 108
NI
X 181
22,711
Total
waste rock
(103 TPY)
0
0
5
907
912
NI
85
997
Ratio of
total waste
rock to ore
0
• o .
0.26
10.0
8.4
NI
0.47
0.04
Overburden
(103 TPY)
0
0
0
0
0
NI
0
0
Ratio of
total overburden
to ore
n
0
0
0
0
NI
0
0
Tailings
J103 TPYJ
'
. '-
11
90
101
NI
178
279
Ratio of
tailings
to ore
-
- .'.
0.63
0.99
0.93
NI
0.98
0.01
Wet waste
(103 TPY)
. . . -
'
125
1.021
1,146
NI
2,178
3,324
Ratio of
i. . waste
to ore
-
'
7.3
11.2
10.6
'
12.0
0.15
Metal
ore
Tl-Zr
Ti-Zr
Sb
Be
RE
Sb
* Compiled from data in Table 108.
Note: - = No information.
NI = Not included.
RE = Rare earths. •
TABLE 111
TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING MISCELLANEOUS ORES, SIC 1099, FOR 1974
(METRIC)
Dry waste weight ( 103 TPY)
State
Montana
Utah
Idaho
Total
Total
EPA process
Region waste
VIII 16
VIII 997
X 263
1,276
Potentially Waste rock
hazardous
waste
11
90
178
279
Total
5
907
85
997
Potentially
hazardous
0
0
0
0
Overburden
Total
0
0
0
0
Potentially
hazardous
0
0
0
0
Concentrator tailings (10
Dry
Total
11
90
178
279
weight
Potentially
hazardous
11
90
178
279
Wet
Total
125
1,021
2,178
3,324
3 TPY)
weight
Potentially
hazardous
125
1,021
2,178
3 , 324
Metal
ore
Sb
Be
Sb
Sh-Bc
Source: References 2 and 8.
-------
TABLE 112
PROJECTED TOTAL AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING MISCELLANEOUS ORES FOR SIC 1099, FOR 1977
(METRIC)
EPA
State Region
Montana VIII
Utah . VIII
Idaho X
Total
Total
process
waste
17
1,056
294
1,367
Source: References 2, 6, and 8.
£
PROJECTED TOTAL
EPA
State Region
Montana VIII
Utah VIII
Idaho X
Total
Total
process
waste
19
1,176
357
1,552
Dry waste weight ( 103 TPY)
Potentially Waste rock Overburden
hazardous Potentially Potentially
waste Total hazardous Total hazardous
12 5 0 0 .0
95 961 00 0
199 95 00 0
306 1,061 6 00
Concentrator tailings ( 103 TPY)
Dry weight
Potentially
Total hazardous
12 12
95 95
199 199
306 306
TABLE 113
AND POTENTIALLY HAZARDOUS WASTE FROM MINING AND CONCENTRATING MISCELLANEOUS ORES, SIC
(METRIC)
Dry waste weight (103 TPY)
Potentially Waste rock Overburden
hazardous Potentially Potentially
waste Total hazardous Total hazardous
13 6 000
106 1,070 000
241 116 000
360 1,192 000
Wet weight
Potentially
Total hazardous
133 133
1,082 1,082
2.440 2,440
3,655 3.655
1099, FOR 1983
Metal
ore
Sb
Be
Sb
Concentrator tailings ( 103 TPY)
Dry weight
Potentially
Total hazardous
13 13
106 106
241 241
360 360
Wet weight
Potentially
Total hazardous
148 148
1,205 1,205
2.960 2.960
4,315 4,315
Meta 1
ore
Sb
Be
Sb
Source: References 2, 6, and 8.
-------
PRILL, BLAST, LOAD I
HAUL ro MILL
-IS //V
OK£
i
1°
WASTE
ROCK
&GUSH/HG- i
SCKEEHING
*> //v.
OKE
WET G-PINDIN&-
6, CLASSIFICATION
(TO leOVo -200 MESH)
T"
—
CONC.
co/vc
(J/VOSRFLOW
RETPEATMENT CIRCUIT
UNDERFLOW
KEG-RIND CigCuiT
(6R/ND TO Qb*. -325 MESH)
TWO FI.OTATIOM CIBCUITS
•f
F//VAL PWITE
TOSMELTEK.
^ASKOFLOAJ
fLOrAT/ON CIRCUIT +2
CONG-
SOLL,
>
•Jl
A.tJTIMOM LEACH PLANT
L&CHED W. MaiSQ3 SCL'fr
W4SH£t> i FILTERED
-|
l-li 6-1-1 -GPAPE a* £P£/V
TICK, S/LY£B-COPf>Eff £^crmc)L\
ro SMELTEQ
™
J© ' |®
"f
TAIL INS S POHO
( !
FLOTATION c/ecu/T *3
\~
®^ C0PPEK
TAILM+S
MATERIALS BALANCE AND COMPOSITIONS
\
I
i DATA ITEMS
i
1 Quantity, TPY
I Metric TPY
WASTE
ROCK
94,000
85.000
•\i.j
CRUDE
ORE
200,000 .
181,000
~\3J~ —
SILVER
CONC..
274
249
^
PYRITE
CONC.
1.500
1,360
^
FINAL
TAILINGS
196,000
178,000
\OJ
ANTIMONY
1,000
907
^U
WATER FROM
FLOTATION
2,200,000
2,000.000
^
ANTIMONY
TAILINGS
DRY
530
480
^
WATER FROM
ANTIMONY
CONCENTRATOR
1,400
1.260
Source:
FLOTATION ADDITIVES
NAME
Aerofloot 242
Methyl Amyl Alcohol
Aerofloof 31
S-3477
Creosote Oil
Zinc SulHde
Sodium Sulfite
Xanfhote Z-1I
LBAON
OF ORE
0.015
0.07
0.01
0.02
0.01
0.30
0.40
0.005
G/MT
OF ORE
7.50
35.00
5.00
10.00
5.00
150.02
199.96
2.50
References 2 and 8.
Figure 33. Mining and concentrating of antimony.
262
-------
Description of Individual Waste Streams. There is only one waste stream
associated with the underground mining of antimony. This stream consists of
waste rock that is disposed of in waste dumps. The waste rock consists mainly
of siderite and quartz with trace amounts of the valuable minerals of the
region.
Identification of Potentially Hazardous Materials. The waste rock from
mining operations does not contain any materials that are considered potentially
hazardous. The amount of total waste rock generated from antimony mining is
about 30 percent of the total amount of materials handled.
Concentrator Processes.
Flotation. Flotation is used to recover antimony and other metals and
prepare concentrate.
The number of circuits used in the flotation process varies depending
upon the number of products to be recovered in conjunction with the antimony.
A typical flow diagram for concentrating antimony using a three circuit-flotation
process is shown in Figure 33.
Crushing Plant. Ore of 38 cm (15 in.) maximum size is hoisted from the
mine to a coarse ore bin. From the bin it is fed to the crushing plant by an
apron feeder. The ore is crushed to a size of less than 8 cm (3 in.) in the
primary crusher and is passed to a secondary crusher which reduces the ore
to 2.5 cm (1 in.).
Wet Grinding and Classification. The ore is wet-ground in three ball
mills which are arranged in a closed-circuit with a classification unit. The
ore is ground to 60 percent -200 mesh to free most of the valuable minerals
from the waste rock. In normal operations, the three ball mills can grind
between 1,000 and 1,100 MT (1,102 and 1,213 tons) of ore per day depending
upon the hardness of the ore. The ore slurry is fed to the No. 1 flotation
circuit.
No. 1 Flotation Circuit. This flotation circuit consists of Fagegren
flotation cells grouped as two cleaners, four roughers, and four scavengers.
Two reagents, Aerofloat 242 and methylisobutylcarbinol, are added in this
circuit. These reagents promote the collection and froth flotation of the
silver-bearing tetrahedrite and also the chalcopyrite. Approximately 96 percent
of the silver-bearing tetrahedrite is recovered in this circuit. The concentrate
from this circuit goes to the retreatment circuit. Underflow from the No. 1
flotation circuit is sent to the No. 2 flotation unit.
263
-------
Retreatment Circuit. The retreatment circuit consists of eight flotation
machines arranged as two cleaners, four roughers, and two conditioners. Chemical
reagents (ZnSO^ and Na2SO;j) are added to the treatment circuit heads coming
from the No. 1 flotation circuit. These chemicals depress most of the galena
and pyrite present, which leaves the roughers as a tailing and is combined
with the concentrate from the No. 2 flotation circuit. The pulp fed to this
circuit contains about 20.5 to 27.4 kg Ag per metric ton (600-800 oz/ton)
of concentrate. The final concentrator product has a grade of approximately
44 kg Ag per metric ton (1,294 pz/ton). This finished product from the two
cleaners is high grade tetrahedrite concentrate. This concentrate is dewatered
and is sent to the antimony plant for antimony removal.
No. 2 Flotation Circuit. The No. 2 flotation circuit is the same as the
No. 1 flotation circuit. In this flotation circuit, Xanthate and Aerofloat
31 are added to the underflow from the No. 1 flotation circuit. This floats
the remaining tetrahedrite, galena, and some of the pyrite. The concentrate
from this circuit is combined with the retreatment circuit tailings and sent
to the regrind circuit for upgrading.
No. 3 Flotation Circuit. The No. 3 flotation circuit is the same as
the No. 1 and No. 2 flotation circuits. The purpose of this circuit is to
recover the valuable minerals that are present in the No. 2 flotation circuit
tailings. Copper concentrate is recovered, the final tailings produced are
pumped to a sand plant, and the fines are sent to a tailings pond for final
disposal. The coarse tailings are pumped to the mine, mixed with cement, and
used for backfill. Analysis of the final tailings leaving the concentrator
is shown in Table 114, along with background concentrations. The concentration
of most elements in the tailings is considerably higher than background values,
which were determined for this mining area.
Regrind Circuit. In this circuit, the low grade silver concentrate from
the No. 2 flotation unit is ground to 95 percent -325 mesh in order to free
the tetrahedrite that has been locked in middling particles with the quartz
and the pyrite. After grinding, the material is concentrated and upgraded in
two small flotation units into an end product that contains 2.7 to 3.4 kg of
silver per metric ton (80-100 oz/ton).
Antimony Concentrator. The high grade silver tetrahedrite concentrate
from the retreatment circuit is hauled by truck to the antimony concentrator
for removal of antimony.
264
-------
The tetrahedrite concentrate is leached at 100 C for about 14 hr in
concentrated sodium sulfite solution in order to dissolve the antimony. This
is a batch leaching operation. After leaching, the pulp is transferred to
thickeners, where the solids settle to the bottom, and the clear liquor is
decanted. This pregnant solution consists of antimony present as thioantimonate
and is sent to an electrolysis operation. The settled solids are washed,
filtered twice, and shipped to the smelter as a high grade silver-copper
product.
Electrolysis. The clear pregnant solution from leaching is processed
through a very complex eletrolysis operation in order to produce metallic
antimony. The electrolysis process is a batch operation in which antimony is
deposited on cathode plates as a brittle nonadherent layer. Every 4 days the
cathodes are removed for stripping. The product produced, containing about
95 percent antimony, and 5 percent arsenic, is shipped to smelters.
In the electrolysis process, barren electrolyte is recycled to the leach
plant; and anolyte solution, which has lost 50 percent of its alkalinity, is
discarded to a tailings pond. This material, known as "fouled anolyte,"
contains 43 percent of the caustic soda entering the plant.
Mass Balance of Material. The mass balance for a typical operation is
included in Figure 33. All of the wastes produced in antimony concentrating
operations are pumped to tailings ponds. These wastes are the tailings from
the No. 3 circuit and the fouled anolyte from the electrolysis plant. The waste
from the flotation circuit is the total tailings from the flotation operations
and consists of undissolved matter in the underflow (tailings) from the
flotation operation.
The tailings waste consists of the minerals present in the ore. These
elements are potentially hazardous if they should enter the environment.
Sodium hydroxide contained in the fouled anolyte gives the tailings water a
higher than normal pH and helps to prevent the metallic elements from
becoming solubilized.
Identification of Potentially Hazardous Materials. The tailings from
concentrating antimony ores are considered potentially hazardous. Table 111,
p. 260, shows the total and potentially hazardous wastes from the mining and
concentrating of antimony ores. In 1974, there were 189,000 MT dry weight
(208,000 tons), and 2,303,000 MT wet weight (2,530,000 tons) of potentially
hazardous wastes resulting from the concentrating antimony ores.
Tables 112, and 113, respectively, show the 1977 and 1983 projections for
total and potentially hazardous wastes resulting from the mining and
concentrating of antimony ores.
265
-------
Waste Treatment and Disposal.
Mining Wastes and Treatment Disposal. Based on the returned questionnaires,
company data, and our engineering judgment, there is no potentially hazardous
waste resulting from the mining of antimony ores in the United States.
Concentrating Waste Disposal Operation. A flow diagram describing the handling
of the potentially hazardous waste generated in antimony concentrating is shown
in Figure 34.
Flotation. The waste from the flotation circuit consists of tailings that
are made up of fine sand-like particles containing materials listed in Table 114.
Note that the concentration of the mill tailings is above background levels,
and therefore, considered potentially hazardous. These tailings are pumped
to a sand plant and separated by cyclones into two fractions according to size.
The coarse sands, which are 60 percent of the tailings, are pumped back to
the mine, mixed with cement, and used to fill stopes. The fine materials are
pumped to the tailings pond for final disposal.
Antimony Leach/Electrolysis. The waste from the antimony leach/electrolysis
process consists of fouled anolyte containing some antimony and 43 percent of
the caustic soda entering the plant. This spent anolyte solution is pumped
to a tailings pond for disposal. The caustic solution is added in such
quantities to raise and maintain the pH of the liquid in the tailings pond
to over 8; this is beneficial in that materials considered potentially
hazardous are not soluble in basic solution and settle out of the water and
remain in the pond.
Waste from the flotation operation goes either to the tailings pond or
back to the mine. All the waste from the leach/electrolysis process goes to
the tailings pond. The tailings pond covers about 6 hectares (15 acres) and
is about 3 km (2 miles) from the mill. The pond was started about 7 years ago;
the dam was constructed out of mine waste and gravel from a creek bottom. There
was no surface preparation for the tailings pond, except for the gravel bottom.
The pond consists of dams or dikes on all four sides and is a complete pond
in itself. Vegetation has been started on the dam. It consists of grasses,
seedlings and elderberry bushes. The bushes, grasses, and seedlings are growing
on three sides of the dam, and as the dam has just been raised, they reach
about halfway up the sides of the dam.
Levels I, II, and III Technology. Level I technology for both antimony
concentrating processes (flotation and leach/electrolysis) involves discharging
the wastes to a tailings pond. Figure 34 shows an example of Level I technology.
266
-------
M
ON
V/V0£RGROUND MIME
WASTEROCK
8S,OOO MT
ORE
I8I,OOOMT
FLOTATION
COA/CENTRA TORS
TAILINGS
/78.000MT
jar BY PRODUCT
1.360 MT
y PRODUCT
24
-------
TABLE 114
ANALYSIS OF TAILINGS FRCM ANTIMONY MILL
Concentration (ppm)
Element
Calcium
Cadmium
Copper
Iron
Potassium
Magnesium
Manganese
Sodium
Lead
Antimony
Zinc
Mill tailings
1,887
1.3
226
252,550
285
5,683
20,147
164
657
437
188
Background
1,500
-
21
11,800.
' 1,800
3,700
490
151
51
- •
150
Source: Reference 7.
268
-------
Level II technology consists of vegetating and stabilization of dams and
dikes, surface preparation of the area before establishment of tailings ponds,
and cycloning the tailings and using the coarse fraction as mine backfill.
Level II technology is equivalent to Level III technology for antimony
concentrating waste. Figure 34 shows an example of Levels II and III
technology.
One of the two operating antimony concentrators has a Level II tailings
pond and mine backfilling operation and produces 70 to 75 percent of the antimony.
The other concentrating operation is new, and there is no information available
to classify its disposal operations.
Based on the information supplied from industry, returned questionnaires,
and our engineering judgment, there were in 1974 approximately 189,000 MT
dry weight (218,000 tons) of potentially hazardous wastes resulting from the
concentrating of antimony ores disposed of on land. One hundred percent of
the waste is considered potentially hazardous (Table 111, p. 260).
Future Adequacy of the Technology Identified. There are no predicted
changes in composition of waste as air and water pollution control facilities
are installed or efficiencies improved. There may be slight changes in the
volume of waste resulting from the concentrating of antimony ores.
Energy and Cost Requirements. The energy amounts and costs associated with
the proposed treatment and control technologies have been estimated as
a portion of the total cost necessary to implement the recommended
technologies,
Waste Treatment and Disposal Costs--Antimony.
Disposal Costs - Technology Levels I, II, and III. Potentially hazardous
waste disposal practices in antimony concentration utilize a tailings pond.
All three levels of technology are identical for this analysis.
Similar to lead-zinc disposal practices, the antimony tailings pond is
initiated with a starter dam. A sand plant is utilized to build up the dikes
as the tailings accumulate. In the representative plant, the tailings pond covers
6 hectares (15 acres) and is located 3 km (2 miles) from the mill. The longer
distance from pond to mill results in higher tailings pump and pipeline costs.
Vegetation has also been started on the sides of the tailings dams.
Antimony waste disposal costs are presented in Table 115. The same cost
assumptions as the copper cost assumptions are used. Costs per metric ton
and total industry costs are presented. The total cost to the antimony
mining and concentrating industry would be from $0.31 to $0.32 million per
year. This would be 0.5 percent of the value added in mining and concentrating
antimony ores.
269
-------
TABLE 115
DISPOSAL COSTS', POTENTIALLY HAZARDOUS WASTE FROM ANTIMONY, TECHNOLOGY
LEVELS I', II, AND III', TAILINGS POND
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Case "A" Case "B1
Capital
Land (15 acres at $5,000/acre)
Starter dam*
Tailings pump and pipelinet
Sand plant*
Labor*
Supervision*
Maintenance*
Insurance and taxes*
Energy and power*
Vegetation^
Total
Metric tons of potentially
hazardous waste/year
Cost/metric ton of potentially
hazardous waste disposed
Total cost if entire industry
adopted ($106/yr)
Total cost as percent ,of value
added
$ 8,775
4,095
45,000
6,786
7,176
1,794
27,061
11,220
1,500
800
$114,207
69,398
$1.65
$0.32
0.5%
$ 7,336
4,095
45,000
6,786
7,176
1,794
27,061
11,220
1,500
800
$112,768
69,398
$1.63
$0.31
0.57.
* See major cost assumptions for copper mining.
t Estimate obtained using cost assumptions for copper mining and MRI
estimate. Tailings are pumped 2 miles on representative operations.
* Vegetation includes dams on all four sides of ponds. Cost is MRI estimate.
270
-------
Beryllium
Industry Characterization.
History of the Industry. Beryllium, discovered in 1797, was first produced
in 1828 in France and Germany. Beryllium has a high melting point and is
exceptionally strong and rigid, despite being nearly as light as magnesium.
The hardening effect of beryllium on copper was discovered in 1926, and its
industrial use for this purpose began soon afterward.
Since World War II, the unusual properties of beryllium have led to intensive
research efforts to expand the supply and utilize beryllium for specialized
structural applications and other purposes. Production of hand-sorted beryl
in the United States is negligible, and requirements are met almost entirely
from imports. Major free world producers of beryllium include Brazil, Republic
of South Africa, Argentina, Australia, and Zambia. :.,;v
Domestic Industry Statistics. The largest domestic source of beryllium-.ore
is the Brush Wellman's Spor Mountain bertrandite mine near Delta,.Utah.'-This
mine contributed almost all of the U.S. mine output in 1974.
Besides the bertrandite ores in Utah, small resources of beryl are known
in New England, South Dakota, and Colorado. Much larger deposits, but with
beryl crystals too small for cobbing, are found in North Carolina. Deposits
of bertrandite in Colorado, New Mexico, and Nevada are amenable to drilling.
Of the two beryllium companies in the United States, Brush. WeiIraan, Inc.,
is fully integrated. Bertrandite ore, mined in Utah, is converted to impure
beryllium hydroxide at its mill near Delta, Utah, about 76 km (47 miles) from
the mine. Further refining to metal, and fabrication, occurs at Elmore, Ohio,
where beryl is also processed into metal.
Use of beryllium as an alloy is increasing, but overall demand is expected
to increase at an 'annual rate of only 2 percent through 1983.
The salient beryllium mineral statistics for the United States were withheld
by the U.S. Bureau of Mines to avoid disclosing individual company confidential
data.
By-Product and Coproducts. Occasionally pegmatites are mined for beryl
alone, but more often beryl is recovered as a coproduct or by-product while
recovering other minerals such as feldspar, mica, lithium minerals, columbite,
tantalite, and cassiterite.
271
-------
Waste Generation and Characterization.
Waste Generation. The land-disposed wastes generated by the beryllium
industry in the United States consist of mining wastes (waste rock) and
concentrator wastes (tailings). Table 109, p. 259, contains the national
production data for beryllium (i.e., all statistics shown for Utah). In 1974,
the total ore mined was 91,000 MI (100,310 tons), and the total beryllium
products amounted to 839 MT (925 tons). The total waste rock generated was
907,000 MT (999,796 tons), and the concentrator tailings waste was 90,000 MT
(99,208 tons).
Table 110 shows the ratio of waste rock and of concentrator waste to ore mined.
In 1974, the ratio of waste rock to ore was 10, and the ratio of tailings to
ore was 0.99.
Tables 111, 112, and 113 show the 1974 and projected 1977 and 1983 potentially
hazardous waste resulting from beryllium mining and concentrating operations.
Open-Pit Mining.
Description of a. Typical Process. At present there is only one operating
beryllium mine in the United States. This mine employs open-pit mining methods.
The mine operation is typical of open-pit mining using conventional earth-moving
equipment in conjunction with drilling and blasting. The ore is loaded by shovel
and hauled 76 km (47 miles) by truck to the mill. A flow diagram for the open-
pit mining and concentrating of beryllium is shown in Figure 35.
The mine waste from beryllium mining consists only of waste rock. The
overburden (topsoil) and waste rock above the ore are removed by a contractor
and piled along with the waste rock from mining operations in surface waste
dumps. •
The ore is mined from a number of small pits in a generalized area which
is considered a single mine operation.
Mass Balance of Materials. The scale of operation at the beryllium mine
ranges from about 70,000 to 95,000 MT (77,162-104,720 tons) of ore mined per
year. The ratio of waste rock to ore is about 10:1, accounting for about
910,000 MT (1,003,103 tons) of waste generated annually. This is the only
waste generated in the mining operations of beryllium ores.
Description of Individual Waste Streams. The only waste stream associated
with the open-pit mining of beryllium is waste rock. This waste rock consists
of rhyolite, gravel, alluvial tuff, feldspar, mica, quartz, sanidine dolomite,
rhyodacite, and limestone.
Identification of Potentially Hazardous Waste Materials. The waste rock
from mining operations does not contain any materials that are considered
potentially hazardous.
272
-------
OPEN f/r MINE.
. BLAST LOAD l>
MAUL TO MILL,
WASTE
ROCK
s&eeet/wo-
STEAM
SULFVSlC \
AC/0
-30 MESH
SLUSKY
TH/CKEHING-
I
M4XB.-UP
SOLVENT EXTRACTION
i, SOLVENT STTflPPIHG
Recrci-eo)
STEAM
EDTA
TWO-STA&e
OVEJ3PLOYV
J, DRUMMING-
FILTRATE
WASTES®
_J
TAILINGS POND
'T^«
I ©
©
BERYLLIUM
coNC&vrwre
MATERIALS BALANCE AND COMPOSITIONS
/T\ S3\ /S\ /7\ ^?N
DATA ITEMS
Quantity, MTPY
Assay, Wt. %
Be
U3^8
F
w
WASTE
ROCK*
907,000
Present
^
CRUDE
ORE
91,000
0.5-0.7
Trace
Present
\&
LEACH
SLUDGE
SOLIDS (DRY)
90,000
BERYLLIUM
CONCEN-
TRATE
839
WATER TO
TAILINGS
POND
931.000
* Some waste rock is back Filled in open pit mine.
Source: Reference 8.
Figure 35. Mining and concentrating of beryllium
273
-------
Concentrator Process.
Description of Typical Process. The beryllium concentrating plant receives
its ore by truck from the mining operations described earlier. The concentrating
operations can be described as a tank leaching process followed by solvent
extraction, hydrolysis, and precipitation.
A typical flow diagram for the concentrating of beryllium is shown in
Figure 35.
Crushing and Wet Grinding. The crude ore containing about 0.5 to 0.7
percent BeO is hauled to the mill by truck where it is crushed and processed
through a conventional wet grinder. The ore is wet-screened, and a -20 mesh
slurry is pumped to feed tanks for the leaching operation.
Acid Leaching. The ore slurry from the grinding operations is treated
with sulfuric acid, water, and heat in large leach tanks. The ore is leached
continuously in the leaching circuit for about 6 hr. The pH of the leach
solution is kept at about 1, and the solution is continually agitated. The
solution temperature is maintained at about 93 C, and at atmospheric pressure.
This continuous leach step dissolves the majority of the beryllium contained
in the ore along with small amounts of other elements.
Thickening. The leachate from the leach tanks, containing some suspended
solids, is removed, and flocculant is added to the solution as it enters the
thickener circuit. The flocculant is added to the circuit to promote the
settling of solids. The thickener circuit consists of eight thickeners employing
countercurrent decantation (CCD) operations. This CCD system is necessary to
separate the suspended solids from the pregnant liquor. These suspended solids
would be a major problem if they were present in the solvent extraction step
of the milling operation. The underflow from the thickeners is the major waste
stream from the milling operation. This waste, which consists of undissolved
matter (tailings) and waste liquor, is piped to a tailings pond for final
disposal. The underflow from Thickeners No. 1 and No. 2 enter Thickeners No.
2 through No. 8, which are classified as wash thickeners. The wash water from
Thickener No. 3 is reclaimed and used in the wet-grinding and leaching
operations. The overflows from Thickeners No. 1 and No. 2 are pumped to the
solvent extraction vessel. The sludge and wash water from Thickener No. 8 are
pumped to the tailings pond.
274
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Solvent Extraction. The pregnant liquor from the leaching operations
is subjected to a solvent extraction (SX) process. The solvent used is a
mixture of a hydrocarbon and dialkyl phosphate with the overall process being
similar to uranium solvent extraction processes (alkylamines and organophosphorus
compounds as solvents). The process is based on the solvent being immiscible
with water and being able to form complexes with beryllium salts. These
complexes are soluble in excess solvent and can be separated from the unwanted
materials. In this stage, the beryllium is distributed between the aqueous
and organic phases; and when conditions are controlled properly, the beryllium
can be extracted quantitatively from the pregnant liquor by extraction into
the organic phase. The organic phase is selective to beryllium only, with most
other constituents of the leachate staying in the aqueous phase. The organic
phase is separated from the aqueous phase and removed to a settling tank where
the beryllium-rich organic layer rises to the surface and is separated by
decantation. The barren aqueous raffinate is pumped to a tailings pond.
Solvent Stripping. The beryllium-rich organic solvent is stripped with
sodium hydroxide in order to transfer the beryllium from the organic phase
back to the aqueous phase. The stripped organic is processed through an acid
conversion process where it is acidified with sulfuric acid, and reprocessed
to the solvent extraction tank.
Iron Hydrolysis and Sulfide Treating. The aqueous solution containing
sodium beryllate (Na2Be02) is treated with steam in an iron hydrolysis step
to produce beryllium hydroxide and sodium hydroxide (P-form). The sodium
hydroxide solution is recycled to the stripping operation. The beryllium
hydroxide produced is in the p-form, which is gelatinous and is difficult to
filter and wash. In order to convert the Be(OH>2 from the (3- to the a-form,
it is sulfided to beryllium sulfide and processed through a two-stage hydrolysis
operation. Sludges, which are produced in both the iron hydrolysis and sulfide
treating steps, are pumped to a tailings pond for disposal.
Two-Stage Hydrolysis. The beryllium sulfide from sulfide treatment is
taken through a two-stage hydrolysis process in which the beryllium sulfide
is oxidized by steam at about 88 to 93 C to beryllium sulfate. The sulfate
is reacted with an organic chelating agent, such as a sodium salt of
ethylenediaminetetraacetic acid, with addition of sodium hydroxide. The sodium
beryllate which is produced is treated with steam at 160 C in the second stage
to form sodium hydroxide and beryllium hydroxide (a;) precipitate. This
precipitate is the desired product of the concentrating operation.
Product Filtration and Drumming. The beryllium hydroxide (a) precipitate
produced in the last stage of the two-stage hydrolysis is granular and is
readily separated in vacuum filters. The beryllium concentrate is packed in
drums and shipped by rail to Ohio.
275
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Mass Balance of Materials. The scale of operations for the concentrating
of beryllium in the United States ranges from 68,000 to 91,000 MX (74,957-100,310
tons) of ore processed each year. This amounts to about 113 to 136 MT (125-150
tons) of beryllium produced per year. The beryllium is shipped as beryllium
hydroxide concentrate to Ohio. The quantity shipped each year is about 839 MT
(925 tons).
Identification of Potentially Hazardous Waste Materials. The tailings
from concentrating beryllium ores are considered potentially hazardous. Table
111, p.260 shows the total and potentially hazardous wastes from the mining
and concentrating of beryllium ores. In 1974, there were 90,000 MT dry weight
(99,000 tons), and 1,021,000 MT wet weight (1,125,000 tons) of potentially
\azardous wastes resulting from concentrating beryllium ores.
Tables 112 and 113, respectively, show the 1977 and 1983 projections for
total and potentially hazardous wastes resulting from the mining and
concentrating of beryllium ores.
Waste Treatment and Disposal.
Mining Waste and Treatment Disposal. Based on the returned questionnaires,
company data, and our engineering judgment, there were no potentially
hazardous wastes resulting from the only beryllium mine in the United States.
Concentrator Waste Disposal Operations. There is only one method being used
for concentrating beryllium. This method consists of an acid leaching process
followed by solvent extraction, hydrolysis, and precipitation.
Acid Leaching. A flow diagram describing the handling of the potentially
hazardous waste generated in beryllium concentrating is shown in Figure 36.
All concentrator waste is disposed of in the tailings pond. The tailings from
beryllium concentrating amount to essentially the same quantity as the ore
processed per year in the mill. The waste generated in the concentrating of
beryllium consists of sludges produced in the thickener circuit, the iron
hydrolysis steps, and the sulfide-treating step, along with the undissolved
matter (tailings) from the underflow of the thickener circuit. All wastes
are combined continuously in a single tank and pumped to a common tailings
pond. Of the ores processed, only 5 to 10 percent are dissolved. Thus, the
leached and washed solids form the vast bulk of the solids discarded in slurry
form. Based on makeup water and tonnage processed, the final solids content in
slurry approximates 7 to 8 percent by weight.
The waste from concentrating averages about 125 ppm beryllium going to
the tailings pond. The beryllium is measured daily, and ranges from 70 to 200
parts per million. Beryllium is the only element analyzed for in the concentrating
operation. Waste from the concentrating operation would also consist of the
minerals present in the beryllium ore.
276
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ORE
<9/,OOOMT
TAILINGS
CONCENTRATOR
SLUDGE
039 Mr
RAFFINATE
TAAJfC
FOK ALL
Source: References 2 and 8.
DAM BUILT W/TH -SOIL
CLAV BASE £
CLAV
LAMD
CLAY PLACED ON SO/L
TA/L/A/G-3
Figure 36. Beryllium wastes—technology Levels I, II, and III.
-------
Tailings in the tailings pond are the only source of potentially hazardous
waste from beryllium concentrating. This waste could possibly contain uranium,
as uranium is present in quantities very slightly higher than background in
beryllium-containing ore. The uranium, as well as small amounts of beryllium
that may be present in the tailings, could be hazardous if these particles
should become airborne. This could occur due to wind erosion and would become
a special problem during dry periods. The tailings pond in present use for
disposal of tailings from the beryllium concentrating operation covers about
81 hectares (200 acres), and is divided into two sections. The pond is
located about 1,500 m (4,921 ft) from the concentrator. The dike was constructed
with earth and soil and has a clay base. The surface preparation for the pond
consisted of clearing the vegetation and placing clay on the topsoil to seal
the bottom. The pond is being stabilized by natural vegetation. The tailings
pond is located in an arid region, and natural evaporation removes water from
the pond, and water is recycled back to the concentrator. No water is discharged
from this pond.
The use of a tailings pond is an acceptable practice to reduce potentially
hazardous waste from beryllium concentrating as long as the pond is covered
by water to prevent dry areas from forming. Dry areas could become a source
of wind erosion which would allow tailings containing beryllium and uranium
to become airborne.
Levels I, II, and III Technology. Figure 36 shows the handling of potentially
hazardous beryllium concentrating wastes and shows an example of Levels I, II,
and III technology. The levels of technology are identical since there is only
one beryllium concentrating operation in the United States, and it meets the
criteria for all three levels, namely: (1) technology currently employed by
typical facilities, (2) best technology currently employed, (3) provides adequate
health and environmental protection.
Level I technology for disposal of beryllium concentrating wastes is the
use of tailings ponds, vegetation and stabilization of dams and dikes along
with surface preparation of the area before establishment of a tailings pond.
The present beryllium concentrator waste disposal system is considered Levels
II, and III technology.
Based on the returned questionnaires, company data, and our engineering
judgment, there were in 1974 approximately 90,000 MT dry weight (99,208 tons)
of potentially hazardous waste resulting from the concentrating of beryllium
ores. This represents approximately 91 percent of the concentrator waste (wet
basis) going to tailings ponds.
278
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Future Adequacy of the Technology Identified. There are no predicted
changes in volume and composition of waste due to the future imposition of
air and water pollution controls. The technology currently practiced at the
single beryllium concentrator now operating is the total impoundment of tailings
wastes. This is the best technology currently available, and it does not present
any air or water pollution control problems.
Energy and Cost Requirements. The energy amounts and cost associated with
the proposed treatment and control technologies have been estimated as a portion
of the total cost necessary to implement the recommended technologies.
Waste Treatment and Disposal Costs--Beryllium Mines. Only one level of
technology exists for the disposal of potentially hazardous wastes in the
beryllium mining industry. Furthermore, the technology is only practiced in
one beryllium mine in the United States. A tailings pond (covering 200 acres)
is utilized. Major cost assumptions are presented in Table 116. Using the cost
assumptions in Table 116, the cost of waste treatment from beryllium was
calculated (see Table 117). The costs range from $0.58 to $0.54/MT of potentially
hazardous waste disposed. The total cost to the industry for utilization of this
technology level would be $50,000/year. This would amount to 0.5 percent of the
value added.
Platinum Group Metals
Industry Characterization
History of the Industry. The first reported use of platinum was by the
pre-Colombian Indians of Ecuador who made many articles of platinum and a
crude platinum-gold alloy. The Spaniards gave the white infusible metal the
name "platina" or little silver. Little interest was taken in the metal,
chiefly because its high melting point made working it difficult. When the
metal was recovered during alluvial gold mining in Colombia, it was usually
discarded.
Early research was directed toward establishing methods for making the native
alloy malleable. The major difficulty in the purification of platinum was due
to the persistent retention of the other metals by the platinum. In 1803,
Sraithson Tennarit recognized the presence of a new element in the insoluble
residue from the aqua regia treatment of platina. Tennant named the new metal
iridium. He also discovered the element osmium. Also in 1803, William H. Wollaston
separated palladium and rhodium from platinum residues. And finally in 1844,
the last element of the group, ruthenium, was isolated by Karl K. Klaus, a
Russian chemist.
Platinum was discovered during the 1920's in Alaska, and large-scale mining
of the placer began in 1934.
279
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TABLE 116
MAJOR COST ASSUMPTIONS - BERYLLIUM MINING - TECHNOLOGY
LEVELS T, II', AND III', TAILINGS POND
A. Land and construction of tailings pond = $180,000
Equivalent annual cost (no resale value) = $21,060/year
B. Other capital equipment costs = $44,000
Useful life = 10 years*
Net capital equipment cost = $60,984
Equivalent annual cost = $7,135/year
C. Annual operating costs = $19,000/year
Source: References 1 and 2.
* MRI estimate.
280
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TABLE 117
DISPOSAL COST, POTENTIALLY HAZARDOUS WASTES, BERYLLIUM MINING,
TECHNOLOGY LEVELS I, II, AND III
(EXPRESSED IN EQUIVALENT ANNUAL 1973 DOLLARS)
Case "A"*
Case "B"t
Capital
Land$
Other equipment
Operating cost
Total
Metric tons of potentially hazardous
waste/year
Cost/metric ton of potentially hazardous
waste
Total cost for industry ($10*>/yr)
Total cost as percent of value added
$21,060
7,135
19,000
$47,195
81,647
$0.58
$0.05
0.5%
$17,606
7,135
19,000
$43,741
81,647
$0.54
$0.05
0.5%
* Case "A" assumes all land required is purbhased in year 1 and has no
resale value after 20 years.
t Case "B" assumes all land required is purchased in year 1 and resold at
same price after 20 years.
$ Land costs were obtained from individual company estimates and include
land selling prices substantially less than $12,350/hectare ($5,000/acre).
281
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Because of the useful properties and relative scarcity of these elements,
they have high intrinsic value. Platinum and palladium are by far the most
abundant and the most important elements of the group. In recent years, the
industrial applications for platinum and palladium have greatly exceeded their
use in jewelry and the decorative arts. The industrial applications for
platinum and palladium are diverse, and the metals are used in the production
of high-octane fuels, vitamins and drugs, electrical components, synthetic
fibers, and fertilizers.
The bulk of the world supply of platinum group metals currently comes from
the Soviet Union, the Republic of South Africa, and Canada.
Domestic Industry Statistics. United States mine production of platinum
group metals accounted for less than 1 percent of the world output. The
platinum group metals are produced as the principal product in the United States
only by the Goodnews Bay Mining Company's placer operation in Alaska.
Virtually all of the remaining domestic primary metals were obtained as
by-products of copper refining in Maryland, New Jersey, Texas, Utah, and
Washington.
Domestic primary production of platinum in 1972 was 156 kg (543 Ib).
Production of platinum group metals (platinum, palladium, and rhodium) was
529 kg (1,164 Ib) in 1972 and 1973 (Tables 107, p. 255, and Table 118). The
U.S. demand for primary platinum in 1972 was 14,525 kg (31,960 Ib). The
demand for platinum group metals in the United States for existing uses is
expected to increase at less than 2 percent per year through 1983. However,
beginning in 1974, use as automobile emission control catalysts is expected
to increase U.S. demand significantly for several years in the mid-19701s.
Table 118 gives the United States platinum group metal production statistics.
By-Product and Coproduct Relationships. Gold is recovered as a coproduct or
by-product with platinum metals from placer deposits. The bulk of the platinum
group metals output is associated with nickel-copper minerals occurring in the
uLtrabasic rocks, dunite, and norite and is separated from the base metals by
ore milling, smelting, and refining processes.
Waste Generation and Characterization. Platinum is being mined in Alaska by
dredging methods. The ore is concentrated by gravity separation. We did not
visit the Alaska facility, and no questionnaire was returned; consequently, we
are unable to adequately describe the mining and concentrating process wastes.
Waste Treatment and Disposal. MRI did not assess the platinum mining and
concentrating operations with regard to potentially hazardous wastes since
we were unable to obtain information describing the mining and concentrating
of platinum ores.
282
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TABLE 118
U.S. PLATINUM GROUP METAL STATISTICS (KILOGRAMS)*
1969
1970
1971
1972
1973
Mine productiont 684
Refinery production
New metal 560
Secondary metal 11,570
529
622
10,890
560
653
8,650
529
466
7,960
622
622
8,270
Source: Reference 6.
* Data reported in troy ounces and converted to kilograms.
t From crude platinum placers and by-product platinum-group metals recovered
largely from domestic copper ores.
283
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Rare Earth Metals
Industry Characterization.
History of the Industry. The rare earth elements are those 15 chemically
similar elements with atomic numbers 58 to 71. The first seven members--
lanthanum, cerium, praseodymium, neodymium, promethium, samarium, and
europium--are commonly known as the light subgroup. The remaining eight
members—gadolinum, terbium, dysprosium, holmium, erbium, thulium, ytterbium,
and lutetium—comprise the heavy subgroup.
The group of elements was called the rare earths because the elements were
originally considered to be scarce and because of the earth-like appearance
of the oxide form. It is now realized that these elements as a whole are actually
more abundant than many better known and commonly available elements.
The production of the rare earth elements in the United States, U.S.S.R.,
India, and Australia constitutes close to 90 percent of the world total production,
U.S. demand for rare earths set a record in 1973 as consumption increased in
all major uses, especially in the iron and steel industry. The growth rate was
expected to decrease in 1974 because of slower economic growth and energy
deficiencies, but overall consumption is expected to rise owing to increasing
demand for pipeline steel, nodular iron, and petroleum catalysts. The U.S.
demand for rare earths is expected to increase at an annual rate of 3 to 4
percent through 1980.
Domestic Industry Statistics. The Molybdenum Corporation of America (Molycorp)
mining operation is the only rare earth producer in the United States. The
mine, mill, and solvent extraction plant capacities of the Molycorp complex
at Mountain Pass, California, represent about 62 percent of the total world
capacity. This operation has doubled current production. Employment at this
mine is estimated at nearly 100 workers. Domestic production statistics are
not available because they are confidential company data.
By-Product and Coproduct Relationships. Monazite produced from beach sands
and river gravel is a by-product in mining for ilmenite, rutile, and zircon.
Other minerals often found associated include gold, staurolite, sillimanite,
tourmaline, kyanite, andalusite, spinel, and corundum. Domestically, the
rougher concentrate from the heavy sands usually contains less than 5 percent
monazitej ilmenite and rutile are usually the major constituents, and zircon
is next in quantity. Production from these deposits, therefore, depends primarily
on the demand for titanium minerals. Low-grade monazite concentrate is recovered
in the beneficiation of tungsten minerals in Colorado.
Bastnaesite ores contain no accessory minerals of great value, although
barite probably could be recovered from the Mountain Pass, California, deposit.
284
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Waste Generation and Characterization. The rare earth metal ore is being
mined in California. MRI personnel were refused permission to visit the mine,
and the questionnaire describing the mining and concentrating operations was
not returned by the company. Information is not available in the literature
to adequately describe the process and process wastes.
Waste Treatment and Disposal. MRI personnel were not permitted to visit
the only existing rare earth mining and concentrating operation in the United
States. Therefore, we are unable to classify the waste treatment and disposal
practices resulting from mining and concentrating of rare earth ores.
Tin
Industry Characterization
History of the Industry. Pure tin was used in Egypt as early as 600 B.C.
and as an alloy in bronze implements many centuries earlier. While tin touches
many facets of our daily lives, it is used to improve the properties and
characteristics of other materials, and its functional applications are usually
hidden. It is used in coating cans, joining pipes or electrical conductors,
and in bearings of all types. Tin in metal form is used in collapsible tubes,
pewter ware, and pipes in the food processing industry.
Cassiterite, tin oxide, is the only commercial mineral of tin. Although
cassiterite deposits are known throughout the world, those of economic importance
are limited to a few areas. The principal areas are the placer deposits in Asia,
and Africa, and the lode deposits in South America.
Demand for tin is expected to increase at an annual rate of about 2.2 percent
through 1983.
Domestic Industry Statistics. There are no known tin deposits of economic
grade or size in the United States. The United States has such small tin
production capabilities as to be totally dependent upon other sources. Tin
is mined by marginal operators on an interruptible basis or as a coproduct of
the mining of some other materials. Small quantities of tin concentrates are
produced as a by-product of molybdenum mining in Colorado, and a coproduct
of placer gold mining operation in Alaska. Domestic production of tin supplied
less than 0.2 percent of the total U.S. demand for primary tin in 1972. The
tin mine production statistics were withheld to avoid disclosing confidential
company data. Table 119 gives the U.S. production of tin, almost all of which
is derived from imported ores.
By-Product and Coproduct Relationships. Tin is recovered as a by-product
of molybdenum at the Climax Mine in Colorado, of lead-zinc at the Sullivan
Mine in British Columbia, Canada, and as an impurity in lead refining, but
this total output is an insignificant part of the U.S. production.
285
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TABLE 119
U.S. PR§6i?£?I
-------
Waste Generation and Characterization. Waste from tin mining and concentrating
is not described because it is not presently being mined in the United States.
Waste Treatment and Disposal. There is essentially no mining or concentrating
of tin in the United States, and consequently no potentially hazardous waste.
Titanium and Zirconium
Industry Characterization.
History of the Industry.
Titanium. Although the element titanium was discovered by William Gregor
in 1790, over 100 years passed before it was put to commercial use. The first
commercial application of titanium was an alloy additive to iron and steel,
and the first ferroalloys were produced domestically in 1906. The metal
titanium has been of commercial importance only since 1948. Titanium dioxide
pigment has been produced in the United States since 1918. The use of
titaniferrous materials in welding rod coating first gained acceptance in the
mid-1930's.
Present technology dictates distinct end-use patterns for the two principal
mineral sources of titanium, ilmenite and rutile. Ilmenite is used mostly for
making titanium pigment, but rutile is used for making pigment and metal as
well as a number of other important applications.
The United States is a major source of ilmenite but not of rutile. In fact,
adequacy of the long-term world supply of titanium will depend to a large extent
on successful development of'an increased capability to use the more abundant
ilmenite instead of rutile, the known reserves of which are small and of relatively
high cost.
On the basis of annual titanium pigment capacity, the United States accounts
for an estimated 35 percent of the world total, followed by the U.S.S.R.,and
other communist countries with an estimated 25 percent> followed by the United
Kingdom, West Germany, Japan, and France. The United States accounts for about
one-half of the world productive capacity for titanium metal, followed by U.S.S.R.,
Japan, and the United Kingdom.
287
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Zirconium. Zirconium was prepared in elemental form in 1824, produced
in relatively pure form in 1914, and made into high-purity grade in 1925 by
the deBoer-van Arkel process. This process later was used for the first
commercial zirconium production in the United States. A method of producing
high-purity zirconium developed by the U.S. Bureau of Mines was adopted by
private industry in 1953. Soon afterwards the zirconium industry expanded and
was able to produce large quantities of high-purity and commercial-grade
zirconium. Today there is considerable interest in the use of zirconium metal
as structural material for nuclear reactors and chemical processing equipment.
However, the consumption of the element zirconium historically has been in the
form of mineral zircon, and as the oxide. Zircon is recovered only as a
coproduct or by-product in connection with mining for titanium minerals and
has been used extensively as facing for foundry molds, particularly in the
iron and steel industry. Zirconium oxide has use in refractory applications.
Domestic Industry Statistics
Titanium. Four mines and mills were producing titanium ores in the United
States during 1974. These were located in New Jersey, Virginia, and Florida.
There were no data reported by the mine in Virginia, and therefore statistics
for titanium in Table 109, p. 259, do not include the Virginia mine.
Demand for titanium sponge is expected to increase at an annual rate of
about 8 percent through 1983. Industrial uses, as contrasted with aerospace
uses, are increasing their share of titanium consumption, although at a slow
rate.
Table 120 shows the production and mine shipments of titanium concentrates
from domestic ores in the United States.
Zirconium. E. I. du Pont de Nemours and Company currently is the sole
domestic producer of the mineral zircon, with mines and mills at Starke and
Lawtey, Florida. Employment at these mines was estimated to be about 300
workers in 1974.
Demand for zirconium nonmetal and metal is expected to increase at annual
rates between 3 and 5 percent through 1983. Domestic zircon resources are large,
and increases in demand will probably be met by increased domestic production.
The domestic statistics concerning the production of zirconium were withheld
by the U.S. Bureau of Mines to avoid disclosing confidential company data.
288
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TABLE 120
PRODUCTION AND MINE SHIPMENTS OF TITANIUM CONCENTRATES FRCM
DOMESTIC ORES IN THE UNITED STATES*
Shipment st
Year
1969
1970
1971
1972
1973
Production
(Ml)
844,813
787,396
619,675
631,153
712,383
Quantity
(MT)
810,147
835,484
647,376
674,410
737,904
Ti02 Content
(MT)
436,281
442,069
352,715
381,822
423,738
Source: Reference 9.
* As of June 30 each year.
f Data given in tons and converted to metric tons.
289
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By-Product and Coproduct Relationships.
Titanium. Ilmenite and rutile often occur in the same deposit. Both
minerals have been recovered in significant quantities from operations in
Florida, and Georgia, but the two minerals were not always separated before
use. Rutile and ilmenite in sand deposits may be associated with zircon,
monazite, garnet, staurolite, and other heavy minerals. Most deposits are
mined chiefly for rutile and ilmenite, and one or more of the associated
minerals is recovered as a by-product; however, in some instances the titanium
values are of secondary importance.
Zirconium. By-product mineral zircon is recovered commercially in the
dredging of heavy mineral-bearing sands for ilmenite and rutile in Florida,
and Georgia. Monazite, staurolite, and xenotime also were by-products in the
mining operation in Georgia. Zircon, recovered as a coproduct from mining
operations for titanium minerals, is imported from Australia.
Waste Generation and Characterization.
Waste Generation. The national production data for titanium-zirconium are
shown in Table 109, p. 259. In 1974, the total ore mined was 22,422,000 MT
(24,716,000 tons), and the total production of products amounted to 668,000
MT (736,000 tons). There was no generation of waste rock or overburden, since
materials classified as waste rock in other mining operations are returned
directly to the mine site. Statistics for the New Jersey operation were based
on the questionnaire returned from the operating company. The State of Georgia
Bureau of Mineral Statistics reported that the Georgia mines were not operating
in 1974. The companies reported no waste disposal as all wastes are returned to
the dredge pit, covered, and revegetated as part of the mining operation.
Mining Process. The common method used for mining of titanium and zirconium
ore in this country is dredging. The ore is contained in beach sands at a
concentration of about 4 percent heavy metals and is mined by dredging.
Description of Typical Process. After the surface over the ore body is
cleared of trees, it is dredged. The dredge is a suction type equipped with an
18-rpm cutterhead. It has two spuds located at the stern which are used as
pivots in conjunction with swinglines to advance the dredge. A 51-cm (20-in.)
dredge pump powered by a 900-hp motor digs at rates up to 1,200 MT/hr (1,323
tons/hr). The presence of roots in the ored body frequently causes plug-up.
To overcome this, a "root hog" was designed to chop up the roots. This is
divided into two compartments by parallel bars with about 13-cm (5-in.)
spacings. The dredge pump receives only the material that passes through the
13-cm (5-in.) openings. The bars start at the bottom of the box and slope upward
15° to the back. At the back, there is a combination hammermill and wood chipper
consisting of a heavy rotor with hard-faced teeth that mesh between fixed dies.
The rotor turns at 300 rpm, crushing the hardpan and roots until they pass
through parallel bars.
290
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The dredge pump discharges into a 61-cm (24-in.) pipeline of 12.2-m
sections supported on pontoons. These connect the dredge with the floating
wet mill. Two 90° ells are used to make an S-curve in the pipeline for greater
dredge flexibility. Figure 37 is a flow sheet describing mining and concentrating
of the ore. There are no wastes in the mining process, since all material mined
is fed to the wet mill.
Mass Balance of Materials. The dredges mine at a rate of 815 to 1,090
MT/hr (898 to 1,202 tons/hr), and all material dredged is pumped directly to
the wet mill where the ore is upgraded to a primary concentrate. This is shown
in the table on Figure 37.
Description of Individual Waste Streams. There are no waste streams from
the dredging operation.
Concentrator Process.
Wet Milling. The wet mill is constructed on three separate barges tied
together as one unit. The flow diagram, Figure 37, describes the concentrating
of the ore. The slurry from the dredge discharges onto two vibrating head-feed
screens with 6-cm (2-in.) openings. The oversize falls by gravity into a
hammermill where lumps of hardpan are disintegrated. The discharge from the
hammermill flows by gravity to a screen with 2.5-cm (1-in.) openings. The
screen oversize consists primarily of roots which are discharged into the pond
as waste. The undersize is pumped back to the head-feed screens.
The undersize, < 0.6 cm (0.24 in.), from the head-feed screens flows by
gravity to two large rougher, feed sumps. The low density slurry from the dredge
was originally dewatered by rake classifiers. These were later eliminated when
the feed sumps were modified to serve as overflow dewatering sumps.
The wet mill treats raw ore containing about 4 percent heavy mineral at
rates up to 1,090 MT/hr (1,202 tons/hr). From this raw ore it produces a
concentrate averaging 85 percent heavy minerals with a recovery of 80 percent.
Many of the heavy minerals lost are coarser sized silicates with lower specific
gravity. Much of the Ti02 l°st *s altered porous leucoxene with low apparent
density.
291
-------
DRED&E MINING-
©
O-RINOING-
SEPARATION
(GRAVITY)
CLASSIFIERS
HI TEMSlON
SEPARATORS
MAGNETIC
SEPARATION
0/f/fS/Zf
TO DREDGE
POND
SCREENING
REFINERY
CONC.
WASTE
PRODUCT
KEFINERY
NON MAGNETIC
WASTE,®
UMDER
SIZE
SEPARATOR
SO*
PRODUCT PRODUCT PRODUCT PRODUCT
MAG-WETIC
SePAKATOK
TAIUNG-S POND
OVERFLOW
LIME TREATMENT
SOLIDS
\WATER
TO STREAM
MATERIALS BALANCE AND COMPOSITIONS
DATA ITEM
Quonttty. TPY
Melrlc TPY
Aiiov. °o TiO?
ZrOl
H«ov> Metoll
Lig^in
ft
Al
\L>
ORE
17.000.000
15. 422. US
0.8 - 1.5
2.5 - 3.2
1
w
FEED TO
REFINERIES
679.448
616.566
W
TOTAL WASTES
FROM
REFINERIES
36.700
33.294
.100
^V
OVERSIZE
TO DREDGE
POND
16.311.6
14.798
W
FUTURE
OPERATION
0
vy
TOTAL
TI02
PRODUCT
428. 153
388.414
63
0
Pr«wnt
Preient
ZIRCON
PRODUCT
122.328
110.974
Present
0
Present
Present
STAUROLITE
PRODUCT
45.876
41.618
Iron-
Aluminum
Complex
0
ZIRCORE'
PRODUCT
15.290
13.871
Kyanit«
Sillimanitc
Zirconium
Mixture
Preient
Prttant
Source: Reference 8.
Figure 37. Titanium and zirconium operations --Ti-Zr
-13.
292
-------
The wet mill uses three stages of concentration. The rougher stage contains
704 spirals which receive a concentrate averaging 10 to 15 percent heavy
minerals; a middling containing 4 to 5 percent heavy minerals; and a tailing
containing < 1 percent heavy minerals. The rougher spirals originally had five
turns, but low recovery prompted the addition of two turns to each to make
seven-turn roughers, with one cutter at the end of each turn. The top four
turns process concentrate, and the bottom three turns process middlings. The
middling is collected in four small sumps and is pumped back to the head feeder
for retreatment. The.tailing discharges into a large sump, from which it is
pumped to the pond as backfill. Since 96 percent of the dredge feed is returned
to the pond at the same rate as the dredge is mined, the dredge and the pond
actually move slowly forward in the direction of mining. About every 2 weeks
it is necessary to move the floating wet mill to keep up with the advance of
the dredge.
The rougher concentrate is collected in four small sumps and pumped to
a larger cleaner feed sump. From here it is pumped to 265 five-turn cleaner
spirals. These spirals have cutters placed at the end of each turn. The upper
four cutters receive a concentrate containing 40 to 50 percent heavy mineral.
The bottom cutter receives a middling containing 10 to 15 percent heavy
mineral. The middling flows by gravity to a small sump and is pumped back to
join the rougher middlings for reprocessing in the rougher spirals. The grade
of this stream is about equal to that of new feed.
The cleaner spiral concentrates flow by gravity to feed 132 three-turn
finisher spirals. For control of the finisher concentrate grade, clear water
is used in this stage to assist operators in making cutter adjustments. In
contrast, discolored water is used in earlier stages where no cutter adjustments
are made. The finisher tailings flow by gravity to the cleaner feed sump for
retreatment. The concentrates are collected in two sumps and pumped by 10-cm
(4-in.) pipeline to the land-based dry mill area. At times, pumping distance
has been 5 km (3 miles). Booster stations are located at 460-ra (1,509-ft)
intervals along the pipeline. Mass balance data for this operation are contained
in the flow diagram, Figure 37. Only 4 percent of the material mined is sent
to the dry milling step; 96 percent of the material is put back into the dredge
pond.
Dry Milling. Wet mill concentrate is a mixture of titanium minerals, heavy
mineral silicates, and quartz. The function of the dry mill is to recover the
titanium minerals and further separate them into an ilmenite product and a
leucoxene-rutile product. The process takes advantage of the conductivity of
the titanium minerals in high tension treatment, and the higher magnetic
susceptibility of the ilmenite in separating the two titanium fractions.
293
-------
Grain coatings on the mineral particles contribute to poor separation in
the dry mill. The coatings are removed chemically by using hot sodium hydroxide.
The scrubbed ore is loaded into feed bins by a front-end loader and
inclined conveyor. It is then dried by two countercurrent rotary dryers. The
dryers we-'e originally coal-fired, but were converted to fuel oil for economy.
A shaking conveyor transfers the ore at 149 C from the dryers to the elevators,
which distribute it to the separators. This shaker is equipped with a screen
section (28 mesh) near the discharge which removes the large quartz to waste.
The dry mill feed is first treated by 28 rougher high tension rolls operating
in parallel. Feed is distributed by offset screw conveyors. The roughers
recover about 90 percent of the titanium minerals in a concentrate containing
approximately 70 percent of these minerals. A middling fraction is taken and
returned to the circuit for reprocessing. The tailings are conveyed to 16
scavenger high tension rolls which recover 80 percent of the titanium mineral
lost by the roughers. A middling is recirculated, and the tailings are transferred
to magnets for recovery of staurolite. All products are conveyed by belt conveyors
running beneath the separators. Sand temperatures up to J.49 C make it necessary
for the belts to be constructed of special heat-resistant rubber.
The rougher concentrates are further cleaned by 12 high tension rolls
operating in parallel. The tailings from this circuit are returned to the
rougher separators for retreatment. A middling is recirculated. The concentrate
which contains 90 percent titanium minerals is passed over four magnets, each
with two banks of five rotors operating in series. The top roll is operated
at a lower flux density and scalps out the ferromagnetic tramp iron and ilmenite.
The nonmagnetic fraction from each roll is fed to the next roll for separation.
The magnetic fraction from all the rolls is collected and shipped as ilmenite.
This product contains 98 percent titanium minerals and averages 64.5 percent
Ti02« The nonmagnetic fraction from the last roll on each magnet is collected
and conveyed to a final cleaning circuit to recover the leucoxene and rutile
for shipment. After the ilmenite has been removed from the cleaner concentrates,
the nonmagnetic fraction contains 10 to 15 percent +45 mesh silicates. These
are removed by dry screening. The undersize is then treated in four recleaner
high tension separators to remove zircon and other silicates. In the original
design, the tailings from this circuit were returned to the cleaner high
tension circuit. It was found that this stream was low grade by comparison
to rougher concentrates, which make up the bulk of the cleaner feed. These
tailings often were responsible for lowering the ilmenite magnet feed grade
which caused the production of low-grade material. These tailings are now
returned to rougher high tension circuit for further treatment. The recleaner
concentrates are finished on two high tension separators. The tailings are
returned to the recleaner separators,and the high-grade concentrate shipped
as a finished product. This product contains 98.5 percent titanium minerals
and analyzes 80 percent TiC^. The overall Ti02 recovery in this dry milling
operation exceeds 97 percent.
294
-------
The finished products are elevated to storage bins. They are loaded from
these bins into covered hopper cars for transportation to titanium dioxide
pigment manufacturing plants. Cars are loaded using a timing device which
opens and closes an orifice discharge device in the storage .hoppers.
The tailings from the scavenger high tension circuit are fed to 10 high
intensity magnets, each with three rotors operating in parallel. The magnets
are equipped with a timing device which de-energizes the magnets periodically
to remove tramp iron from the poles. The magnetic product consists of a mixture
of staurolite with a small percentage of tourmaline, spinel, and silicates
with magnetic inclusions. This product contains 45 to 50 percent Al2(>3 and
13 to 15 percent Fe2Q3. It is marketed for use in portland cement manufacture
where it replaces the use of kaolinite as a source of A1203, and reduces the
amount of slag needed to obtain the 3 percent F6203 required in the cement.
The nonmagnetic fraction from the staurolite magnets, containing 30 to 35
percent zircon, is slurried and pumped to the nearby zircon spiral plant.
Staurolite magnet tailings contain 30 to 35 percent zircon, 15 to 20 percent
aluminum silicate minerals, and about 50 percent quartz. The relatively high
specific gravity of zircon makes gravity separation from the lighter heavy
minerals and quartz possible using spirals. The zircon spiral concentrates
are dried and recleaned by a process patterned after the titanium dry mill.
The wet plant originally used three stages of spirals: 32 five-turn
roughers; 32 three-turn cleaners; and 32 three-turn finishers. The tailings
from the last stages were retreated in the preceding stage, the rougher spirals
took a middling for recirculation. The tailings from the roughers were scavenged
by 24 five-turn spirals. The scavenger concentrates were returned to the roughers,
and the tailings were stockpiled for future use.
A large tonnage of zircon is used for molding sands. Many consumers require
low A1203 zircon for their process, and a maximum quality specification of
1 percent Al2<>3 was set up. This was sometimes impossible to meet with a
three-stage spiral plant utilizing a scavenger circuit. As the zircon market
expanded and demand increased, it became necessary to enlarge the plant. A
second plant was erected using four stages of concentration. There are 32.
two-turn rougher spirals. The tailings from these are remixed and fed to 32
three-turn spirals directly beneath. The concentrates from all five turns are
combined to feed 32 five-turn cleaners. The tailings from these are returned
to the roughers, and the concentrates are collected in a sump and fed to.32
three-turn finishers. The tailings return by gravity to the recleaner feed
sump. The final wet mill concentrate, which contains 95 to 96 percent zircon,
is dewatered by a screw classifier prior to drying and dry milling. After
completion of the second mill, the concentrates from the original mill were
retreated in the last two stages of the new mill to reduce the A1203 content
to the 1 percent specification. In 1958, the original mill was redesigned and
modified to be similar to the expanded mill. The two,mills now operate
independently of each other, and produce'at the required quality.
295
-------
In its natural state, the zircon has a dull brownish-purple color, caused
by breakdown of crystal lattice and alteration. This color is changed to white
by heating to 52.5 C in an oil-fired kiln. The original arrangement used an
airlift to feed the kiln discharge to the .dry mill. This has been replaced
with a 12-m (39-ft) rotating cooler and elevator. A shaker conveyor equipped
with a scalping screen is used to remove any .foreign material from the cooler
discharge.
Zircon wet mill concentrate contains a small percentage of titanium minerals
and staurolite which must be removed. Two stages of high tension are used to
separate the titanium minerals .followed by magnetic separation to remove the
staurolite. Zircon dry mill recovery has been increased by adding a scavenger
magnet to treat the mill tailings. The scavenger concentrate is returned to
the zircon magnet feed for retreatment. Some of the zircon grains lost contain
blebs of -magnetite which make the grains magnetic. This material is unrecoverable.
Zircon specifications include a maximum of .0.15 percent Fe203 and 0.25 percent
Ti02» in addition .to the 1 percent A1203. Finished zircon is stored in bins
until analysis.shows that it meets these specifications. It is then bagged
and shipped in 45-kg (1,445-oz) bags or .loaded into hopper or boxcars for bulk
shipment. The shipping operation is .set up to furnish the type of loading best
suited to the customer's requirements. The finished product averages 98 .to
99 percent zircon.
The rougher concentrates are treated on 320 five-turn rubber-lined cleaner
spirals. The cleaner tailings are returned to the rougher spirals and a middling
recirculated. The cleaner concentrates flow by gravity to feed 160 three-turn
finisher spirals located on the deck below. These make a tailing which is returned
to the cleaner spirals and a concentrate that :ls pumped to the dry mill area
for further "treatment.
Organic 'and clay materials '.are removed from the w,et mill concentrates by
.conditioning. This consists of dewatce'ring, scrubbing with caustic, rinsing,
and final dewatering. Screw classifiers, are used instead of rakes. A settling
sump was incorporated .to recover fines which overflow the classifiers. The
scrubbed concentrate .is spread by use .of a high-speed slinger "on the end of
an inclined conveyor stacker. The slinger is directional and is used to make
several smalt piles .to.allow the res'ldual,water to drain prior to drying.
A double-deck .hummer screen'separates the oversize from the feed to the
leucoxene-rutile cLeaner .circuit. 'JThis serves :two purposes: first, it separates
the coarse (+35 mesh) stauroLite; 'and 'second, .it sizes the feed to .the high
tension separators, in that circult,.'.The fleucoxene-.rutile circuit-consists of
.four high tension rotors. One"treats'the -25 +BO mesh fraction from the screens.
The other three treat the -80 mesh fraction. A.single .stage of high tension
recleaning is possible because of the lower percentage leucoxene and rutile
in the ore. A middling recirculates to the hummer screen. .The tailings were
originally designed to return to the cleaner high tension, but were later
repiped to return to the rougher high tension circuit.
.296
-------
The wastes from the dry mill are organic matter (lignite) and kaolinite,
which is pumped with the minerals from the wet mill. These wastes are harmless
and are treated to settle them on the land.
Waste Treatment and Disposal. On the basis of the information collected
and the mines visited we believe that there are no potentially hazardous
wastes resulting from the mining and concentrating of titanium and zirconium
ores. .
297
-------
REFERENCES
1. Engineering; & Mining Journal, 1973-1974, International Directory of Mining
and Mineral Processing Operations, Published by Engineering &. Mining
Journal, McGraw-Hill, New Ybrkj New York..
2. Unpublished' mining company dataw
3. U.S. Department of the Interior, Bureau of Mines, Mining Enforcement and
Safety Administration, Washington, D.C.
4. U.S., Department of- the Interior,- Bureau of Mines* Unpublished Commodity
Data Report, Antimony (C« Wyche), Dec. 1973.
5. U.S. Department of the Interior, Bureau of Mines. Unpublished Commodity
Data Report, Platinum-Group^ Metals,; Jan«> 1974.,
6. U.S.. Department of the Interior,, Bureau of Mines.. Commodity Data Summaries,
1974, Appendix I to Mining, and Minerals Policy,- 1974.
7. William, R. E.. The role of mine tailings ponds in reducing the discharge
of heavy metal ions to the environment.* Idaho University,: Moscow,. Idaho,
PBV224730, Aug.. 1973.
8. MRI communications with miscellaneous group: metal ore mining companies,
and-American Mining Congress Questionnaires.
9. U.S. Bureau of Mines. Metals, minerals, and fuels,, 1973 Minerals Yearbook.
v.I., ppv 1231-1244.
298'
-------
APPENDIX A
ACTIVE MINE AND CONCENTRATOR OPERATIONS. 1974
299
-------
ACTIVE MINE OPERATIONS. 1974
1021 - Copper (56 Mines)
Name of Operation
American Smelting and Refining Co.
Sacaton Unit
Casa Grande, Arizona
American Smelting and Refining Co.
Mission Operation
Sahuarita, Arizona
American Smelting and Refining Co.
San Xavier Operation
Sauharita, Arizona
w American Smelting and Refining Co,
o Silver Bell, Arizona
American Smelting and Refining Co.
Cr.i.o-»a 'Jnit
Wallace, Idaho
Anaconda Company
Butte, Montana
Anaconda Company
Weed Heights, Nevada
Anamax Mining Company
Twin Buttes Operation
Sahuarita, Arizona
Year
Opened
t . • .! . •. i~wa,
1972
--
1973
1951
1954
--
1953
Number of Mine Type Number of
Mines and Process Concentrators
: 1 Open Pit 1
1 Open Pit 1
1 Open Pit 1
1 Open Pit 2
1 : . Underground 1
2 Underground 1
Open Pit
1 Open Pit 2
Concentration
Process
Flotation
Flotation
Leaching
Iron Precipi-
tation'
Flotation
Leaching
Iron Precipi-
tation
Flotation
Flotation
Flotation
Leaching
Iron Precipi-
tation
Estitnate
Employee
212
708
147
381
200
2,644
670
Open Pit
Flotation
1,606
-------
ACTIVE MINE OPERATIONS, 1974
1021 - Copper (continued)
Name of Operation
Cities Service Company
Copperhill Operation
Copperhill, Tennessee
Cities Service Company
Miarni Copper Operation
Miami, Arizona .
Cities Service Corporation
Pinto Valley
mi, Arizona
Cobre Mines, Inc.
Cobre Mine
Hanover, New Mexico
Continental Copper, Inc.
Oracle, Arizona
Copper Range Company •
White Pine Division
White Pine, Michigan ;
Cyprus Bagdad Copper Company
Bagdad Copper Pit
Bagdad, Arizona
Cyprus Bruce Copper and Zinc Co. 1968
Bruce Mine Division
Bagdad, Arizona
Year Number of
Opened Mines
1899 5
1954 1
1974 1
1
Mine Type Number of
and Process Concentrators
Underground 1
Open Pit 2
Open Pit 1
Solution 1
Mining
Open Pit 1
Concentration
Process
Flotation
.Flotation
Leaching
Iron Precipi-
tation
Flotation
Leaching
Iron Precipi-
tation
Leaching
Iron Precipi-
tation
Estimate
Employee-
903
650
174
22
1974
1953
Underground
Underground
Open Pit
Underground
Flotation 18
Flotation 2,444
Leaching 433
Solvent Extrac-
tion
Flotation
Flotation 562
-------
ACTIVE MINE OPERATIONS. 1974
o
IsJ
1021 - Copper (continued)
Kame of Operation
Cyprus Pima Mining Company
Pima Operation
Tucson, Arizona
Duval Corporation
Battle Mountain Operation
Battle Mountain, Nevada
Duval Corporation
Esperanzo Operation
Sahuarita, Arizona
Duval Corporation Copper Division
Mineral Park Operation
Kingman, Arizona
Duval Sierrita Corporation
Sahuarita, Arizona
Eagle^Picher Industries, Inc.
Creta Operation
Olustee, Oklahoma
Earth Resources Company
Nacinjento Copper Mine Operation
Cuba, New Mexico
Year Number of
Opened . Mines
1957 1
1967 1
1959 1
1964 1
Mine Type
and Process
Open Pit
Open Pit
Open Pit
Open Pit
Number of Concentration Estimated
Concentrators Process ErT^Iovr.ertt
1 Flotation 975
1 Flotation 245
2 Flotation 542
Leaching
Iron Precipi-
tation
1 Leaching 511
Iron Precipi-
tation
1965
1971
Open Pit
Open Pit
Open Pit
Flotation
Flotation
Flotation
1,339
75
116
-------
ACTIVE MINE OPERATIONS. 1974
1021 - Copper (continued)
Name of Operation
El Paso Natural Gas Company
Emerald-Isle Operation
Kingman, Arizona
Year
Opened
Number of
Mines
Mine Type
and Process
Open Pit
Number of
Concentrators
Concentration
. Process
Flotation
Leaching
Iron Precipi-
tation
Estimate
E;r.p loyme
39
Federal Resources Corporation
Bonney Mine Operation
Lordsburg, New Mexico
Gold Field Corporation
San Pedro Copper Mine
Albuquerque, New Mexico. •
Hecla Mining Company
I.nkeshore Mine
Caaa Grande, Arizona
Inspiration Consolidated Copper
Company
Inspiration, Arizona
Kennecott Copper Corporation
Viayden, Arizona
Vcnnecott Copper Corporation
Cr.Vno nines Division Operation
]••;•.- i.ey, :;
-------
ACTIVE MINE OPERATIONS. 1974
1021 - Copper (continued)
Name of 'Operat"ton
Kennecott Copper Corporation
Ray Mines Division
Ray, Arizona
Kennecott Copper Corporation '
Nevada Mines Division
Ruth, Nevada. .
Kennecott Copper Corporation
. Ucah "Copper Div'isi6j>
S^lt Lake City, Utah
' . . ',
Kerr American, "Inc...
Mrmmoth Mine Operation • ..
Key stone -Wai lace Resources
.Moab Mill Operation
Moab, Utah
Magma Copper Company
San Manuel Division
San Manuel, Arizona
Magma Copper Company
Superior Division
Superior, Arizona
McAl ester Fuel Company
7.o:iia Operation
Kirk land, Arizona
Year Number of
Obehe'd Mines.
1955
1906
1956
1910
1966
Mine Type
and Process
Open Pit
Open Pit
Open Pit
Underground
Open Pit
Underground
Underground
Solution
Mining
Number of
Concentra t ors
Concentration Estimated
Process Employment
Leaching 811
Iron Precipi-
tation
Electrowinning
Flotation 2,284
.Leaching
Iron Precipitation
Flotation 4,597
Leaching
Iron Precipitation
Leaching 91
Iron Precipitation
Leaching 80
Iron Precipitation
Flotation
Flotation
2,399
1,062
Leaching 24
Iron Precipitation
-------
ACTIVE MINE OPERATIONS. 1974
1021 - Copper (continued)
Name of Operation
Micro Copper Operation
Lisbon Valley Mine
Moab, Utah
Phelps Dodge Corporation 1917
New Cornelia Branch
Ajo, Arizona
Phelps Dodge Corporation 1951
Copper Queen Branch 1878
Blsbee, Arizona
g Phelps Dodge Corporation 1942
01 Morencl Operation, and Metcalf Mine
Morencl,.Arizona
Phelps Dodge Corporation 1969
Tyrone Branch Operation
Tyrone, New Mexico
Ranchers Exploration and Development Co. 1971
Old Reliable Mine
Mammoth, Arizona
Ranchers Exploration and Development Co.
Blue Bird Mine
Miami, Arizona
Number of Mine Type Number of
Mines and Process Concentrators
Pit
Open Pit.
Open Pit*
Underground
Open Pit
Open Pit
Solution
Mining
Open Pit
Concentration Estimated
Process Employment
Leaching 15
Iron Precipitation
Flotation 633
Flotation 792
Flotation 1,079
Leaching
Iron Precipitation
Flotation 620
Leaching
Iron Precipitation
Leaching 15
Iron Precipitation
Leaching 126
Solvent Extraction
Electrowlnning
-------
ACTIVE MINE OPERATIONS. 1974
1021 ~ Copper (continued)
Name of Operation
USNR Mining and Minerals, Inc.
Copper Mountain Operation
Silver City, New Mexico
U.V. Industries, Inc.
Continental Mines Operation
Hanover, New Mexico
Year
Opened
1970
1967
Number of
Mines
Mine Type
and Process
Open Pit
Solution
Mining
Open Pit
Underground
Number of
Concentrators
Concentration
Process
Estimated
Employment
Leaching 42
Iron Precipitation
Flotation 336
* Closed in 1974.
-------
ACTIVE MINE OPERATIONS. 1974
1031 - Lead and Zinc (47 mines)
Year
Opened
1969
Name of Operation
Amax Lead Company of Missouri
Buick Lead Mine Operation
Boss, Missouri
American Smelting and Refining Co.
LeadviHe Unit Operation
Leadville, Colorado
American Smelting and Refining Co.
Ground Hog Unit
Vanadium, New Mexico
American Smelting and Refining Company
Coy Mine Operations
Jefferson City, Tennessee
American Smelting and Refining Company
Mascott Mine
Jefferson City, Tennessee
American Smelting and Refining Company
Young Mine Operation
Jefferson City, Tennessee
American Smelting and Refining Comapny 1963
New Market Mine
New Market, Tennessee
Number of
Mines
Mine Type
and Process
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Number of
Concentrators
Concentration
Process
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Estimated
Employment
289
121
124
50
700
102
150
-------
ACTIVE MINE OPERATIONS. 1974
1031 - T.ead and Zinc (continued)
Karae of. Operation
Eur.V.cr Hill Corapany
Bunker Hlllf and Crescent Mines*
Kcllogg, Idaho
Ccrro Spar Corporation
Salem, Kentucky
Clayton Silver Mine
Clayton Mine
Wallace, Idaho
Cominco American Incorporated
w Magmong Mine Operation
g Bixby, Missouri
Day Mines, Incorporated
Day Rock
Wallace, Idaho
Eagle-Picher Industries, Inc.
Shullsburg Mine Operation
Shullsburg, Wisconsin
t
Hecla Mining Company
Star Unit Area
Burke, Idaho
Hecla Mining Company
Lucky Friday Operation
Mullan, Idaho
•f Opened in 1885.
$ Opened in 1952.
Year
Opened
1885
1952
1974
1934
1968
Number of
Mines
1940
1950
Mine Type
and Process
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Number of
Concentrators
Concentration
Process
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Estimated
Employment
1,017
20
25
192
85
87
350
192
-------
ACTIVE MINE OPERATIONS. 1974
1031 - Lead and Zinc (continued)
Name of Operation
Hydro-Nuclear Corp.
Linchburg Mine
Magdalen, New Mexico
Idarado Mining Company
Ouray, Colorado
Kennecott Copper Corporation
Tintic Mine and Burgin Mine
Eureka, Utah
Leadville Lead Corporation
Sherman Tunnel Operation
Leadville, Colorado
Minerva Oil Company
Minerva #1 Mine .
Cave In Rock, Illinois
New Jersey Zinc Company
Eagle Mine Operation
Oilman, Colorado
New Jersey Zinc Company .
Sterling Hill Operation
Ogdensburg, New Jersey
New Jersey Zinc Company
Friedensville Mine Operation
Center Valley, Pennsylvania
Year
Opened
1966
1944
Number of Mine Type
Mines and Process
1 Underground
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Number of
•Concentrators
Concentration Estimat
Process Einploym'
Flotation 45
Flotation 355
Flotation 286
No Processing 75
Flotation 115
Flotation 262
Flotation 156
Flotation 207
-------
ACTIVE MINE OPERATIONS. 1974
10.31 - Lead and Zinc (continued)
" Name of Operation
New Jersey Zinc Company,
Jefferson City Mine Operation
Jefferson City, Tennessee
New Jersey Zinc Company,
Austinville Operation
Austinville, Virginia
pzark Lead Company
Sv:.eetwgter, Missouri
^ GzarkrMahoning Company
p Johnson Work§
RQseelare, Ill|t\o.is
Year
Opened
1968
1937
PenJ Orsille Mines end Metals Company '-•=•
Met a. line Falls, Washington
Resurrection Mining Company
Resurrection Mine
Leadville, Colorado
St. Joseph Minerals Corporation 1973
Brushy Creek
Bonne Terre, Missouri
St. Joseph Minerals Corporation 1967
Fletcher Operation
Bonne Torre, Missouri
Number of
•Mines
1
1
Mine Type
and Process
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Number of
Concentrators
Concentration Estimate^
Process Eraplovme'
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
126
467
264
160
92
50
120
144
-------
ACTIVE MINE OPERATIONS. 1974
1031 - Lead and Zinc (continued)
Name of Operation
St. Joseph Minerals Corporation
Indian Creek #23
Bonne Terre, Missouri
St. Joseph Minerals Corporation
Viburnum Operations
Bonne Terre, Missouri
St. Joseph Minerals Corporation
Balmat-Edwards Mine Operations
Edwards, New York
E. G. Somrnerlath Enterprises
Marion, Kentucky
Standard Metals Corporation
Sunnyside Operation
Silverton, Colorado
U.S. Steel Corporation
Zinc Mine
Jefferson City, Tennessee
U.V. Industries
Princess Mine and Hanover Mine
Hanover, New Mexico
Year
Opened
1954
1960
--
--
Number of
Mines
1
2
2
1
Mine Type
and Process
Underground
Underground
Underground
Open Pit
Number of
Concentrators
1
1
1
1
Concentration
Process
Flotation
Flotation
Flotation
Extraction
Heavy Media
Estimated
Employment
114
266
501
20
Underground
Underground
Underground
Flotation
Flotation
Flotation
126
200
400
-------
ACTIVE
OfrERAT-IGNS; 1974
1092 - Mercury (2 mines)
N.ime of Operation
Mine Management Company
Khoxvillie Mine
Napa County, California
One Shot Mining Company
Mantiattan Mine
Napa County, California
Year
Opened
1970
i'968
. lumber b£
• - Mines •
1
i
Mine type
-and 'Prbce'ss
Open Pit
Open Pit
Number of Concentration
Concentrators Process
1 Roasting
Condensing
1 Roasting
Condensing
Estimated
Employment
5
15
-------
ACTIVE MINE OPERATIONS. 1974
1094 - Vanadium and Uranium (158 mines and 17 concentrators)
Kame of Operation
Anaconda Company
Jackpile Faguate Unit Corporation
Grants, New Mexico
Atlas Corporation
Big Indian Mine
Moab, Utah
Continental Oil Company - Pioneer Nuclear
Conquista Project
Falls City, Texas
Cotter Corporation
Cotter Mill Operation
Canon City, Colorado
Cotter Corporation
Schwartzwalder Operation
Colorado
Year Number of Mine Type
Opened Mines and Process
2 Underground
Open Pit
Underground
1973
1960
No Mine
Open Pit
Underground
Number of Concentration Estimated
Concentrators Process Employment
1 Alkaline 597
Leaching
Ion Exchange
1 Dual Process 89
Acid Leaching
Plus Alkaline
Leaching
1 Acid Leaching 357
Dual Process 77
Acid Leaching
Plus Alkaline
Leaching
No Processing 83
Dawn Mining Company
Well Pinit Mine
Ford, Washington
Exxon Company USA
Highland Uranium Operation
Casper, Wyoming
1972
Open Pit
Open Pit
Acid Leaching 250
Ion Exchange
Acid Leaching 83
Solvent Extraction
-------
ACTIVE MINE OPERATIONS. 1974
- Vanadium and Uranium (continued)
Name of Operation
Exxon Corporation
Exxon Company, USA Division
Fall City, Texas
Federal American Partners
Riverton, Wyoming
Four Corners Exploration Company
D;QJ5 Mine
Grants, Newf Mexico
Homestake Mining Company
E33 Mine
Grants, New Mexico
Kerr-McG.ee Corporation
Grants, New Mexico
Kerr-McGee Corporation*
Shirley Basin Operation
Casper, Wyoming
Mountain West Mine, Inc.
Betty Mine
Blanding, Utah
Petrotomics Company & K.G.S.—J.V.
Joint Venture Mine
Casper, Wyoming
.Year Number of Mine Type
Opened Mines and Process
1970
1969
10
1960
Open Pit
Open Pit
Underground
Underground
Underground
Underground
Open Pit
Open Pit
Open Pit
Number of Concentration Estimated
Concentrators Process Employment
0 No Processing 55
Acid Leaching 75
Ion Exchange
Solvent Extraction
No Processing 10
No Processing 15
Acid Leaching 1,200
Solvent Extraction
No Processing 124
No Processing 19
Acid Leaching 85
Solvent Extraction
-------
ACTIVE MINE OPERATIONS. 1974
1094 - Vanadium and Uranium (continued)
Name of Operation
Ranchers Jvrploration and Development Corp.
(jfuaCS, New Mexico ;
Rio Algom Corporation
Lisbon Mine Operation
Moab, Utah
Susquehanna-Western, Inc.
Falls City, Texas
Union Carbide Corporation
Wilson Springs Operation
H- Hot Springs National Park, Arkansas
Union Carbide Corporation
Uravan Operation
Uravan, Colorado
Union Carbide Corporation
Riverton, Wyoming
United Nuclear Corporation
Grants, New Mexico
United Nuclear-Homestake Partners
Grants, New Mexico -
Utah International, Inc.
Shirley Basin Operation
Casper, Wyoming
Year
Opened
1972
1970
1969
1972
1958
Number of Mine Type Number of Concentration
Mines and Process Concentrators Process
1
2
1
3
Open Pit
Underground
Open Pit
Open Pit
0
1 Alkaline
; Leaching
0 No Processing
1 Salt Leaching
Estimated
Employment
15
172
35
177
25
Underground
Open Pit
Underground
Underground
Open Pit
Solvent Extraction
for Vanadium
Concentrate
Acid Leaching 293
Ion Exchange
Vanadium Recovery
118
405
Alkaline Leaching 450
Solvent Extraction
Acid Leaching
Ion Exchange
474
-------
ACTIVE MINE OPERATIONS. 1974
1094 - Vanadium and Uranium (continued)
Name of Operation
Utah International, Inc.
Lucky McMine Operation
River.ton, Wyoming
Western Nuclear
Rox Operation
Casper, Wyoming
Western Nuclear, Inc.
Golden Goose Operation .
Jeffrey City, Wyoming
Other unidentified uranium mines
(including 37 in Colorado)
Year
Opened-
1957
Number of
• Mines
2
Mine Type
•. and Process
Open Pit:
Underground
Number of
Concentrators
1 .;
Concentration
Process
Acid Leaching
Ion Exchange
Estimated
Employment
351
83
Open Pit
Open Pit
Open Pit
Underground
*" Operations at this site were suspended indefinitely in November, 1974.
Acid Leaching
Ion Exchange
Solvent Extraction
150
151
267
-------
ACTIVE MINE OPERATIONS. 1974
1099 - Miscellaneous Mtning (9 nines)
Name of Operation :
.American Smelting and Refining Co.
Manchester Mine
Lake Hurst, New Jersey
Brush Wellman, Inc.
Delta Mill Operation
Delta, Utah
Brush Wellman, Inc.
Spar Mountain Operation
Delta, Utah
E. I. duPont
Highland Mine
Lawtey, Florida
E. I". duPont
TraiLri'Jge Mine
Starke, Florida
Felspar Corporation
Kontpeller Mine
Montpelier, Virginia
Goodnews Bay Mining, Company
Platinum, Alaska
Molybdenum Corporation of America
Mountain Pass Operation
Nipton, California
Year
Opened
1968
1968
1952
1952
Number of
Mines
Mine Type
and-Process
Dredge
Open Pit
Dredge
Dredge
Open Pit
Dredge
Open Pit
Number of'
Concentrators
Concentration
Process
Gravity
Estiaia
Enploy-
80
Leaching
Solvent Extraction
Precipitation
Electrostatic
Magnetic and
Gravity Separation
Electrostatic
Magnetic and
Gravity Separation
Electrostatic
Magnetic and
Gravity Separation
Gravity
Flotation
52
11
150
150
37
93
-------
ACTIVE MINE OPERATIONS. 1974
1099 - Miscellaneous Mining (continued)
Kame of Operation'
Sunshine Mining Company
Big .Creek Mine
Kellogg, Idaho
U.S. Antimony Corporation
Babbitt Operation
Thompson Fall, Montana
Year
Opened
I960
1971
Number of
Mines
1
1
Mine Type
and Process
Underground
Underground
Number of
Concentrators
2
1
Concentration
Process -.
Flotation
Hot Leaching
Flotation
Heavy Media
Separation
Estirsated
Employment
700
14
-------
APPENDIX B
MINES AND.CONCENTRATORS VISITED
QUESTIONNAIRES RECEIVED
SAMPLE QUESTIONNAIRE
319
-------
Company Mame
OX
N>
O
1.
2.
3.
4.
5.
6.
7.
8.
9.
10.
11.
12.
13.
14.
15.
16.
17.
Mine Management Cuopany
One' Shot Mlntn Company
Aaox Lead Comp ny of Missouri
St. Joe Minera s Company
St. Joe Minera s Company
St. Joe Minera s Company
St. Joe Minera s Conpany
Idarado Mining Company
United States Steel Company
Kennecott Copper Corporation
Anaaax Mining Company
Cyprus Pioa Mining Company
McAl ester Fuel Company
Phelps Dodge Corporation
Magau Copper Company
Kennccutt Cupper Corporation
Ranchers Exploration and Development Corp.
Kennecott Copper Corpoiatlci
Brush Ufellman
E. 1. duPont
IB. Bunker Hill Conpany
19. Sunshine Mining Company
20. Union Carbide Corporation
21. Cities Service Company
22. 'Hie Anaconda Company
23. United Nuclear - Hone Stake Partners
24. Cotter Corporation
2%. Union Carbide Corporation
26. Exxon Company
27. Utah International, Inc.
HIKES VISITED
Location ,_ _
Him- Name
Knoxvllle
Manhattan
Bulck
Brushy Creek
Fletcher
Indian -Creek
• Viburnum
Ida ratio
L>a via- Bible
Bingham
Twin Buttes
Pima
Zonla
Copper Queen -
Lavender Pit
Magma
Ray
Bluebird
Burg 1 Ln
Trlxlc
Topaz.
Trail Ridge .
Highland
Bunker Hill
Crescent
Sun ah 1 ne
North Wilson
East Wilson
Span Id Ing
Cal luway
Boyd
Cherokee
Tennessee
Eureka
Jack Pile-Pagvate
Ambroslalake
Operations
Schwartzwalder
—
Highlands
--
City
Hapa County
Hapa County
Buick
Bonne Terre
Bonne Terre
Bonne Terre
Bonne Terre
Ouray
Jefferson City
Salt Lake City
Suhuarlta
Suhuarlta
Kirk land
Bis bee
Superior
Hayden
Miami
Eureka
Delta
_ Starke
l.awtey
Kellogg
Ke 1 Logg
Hot Springs
Copper Hill
Grants
Grants
Canon City
Uravan
Casper
Shirley Basin
State
California
California
Missouri
Missouri
Missouri
Missouri
Missouri
Colorado '
Tennessee
Utah
Arizona
Arizona
Arizona
•Arizona
Arizona
Arizona
Arizona
Utah
Utah
K Lor Ida
Florida
Idaho
Idaho
Arkansas
Tennessee
Hew Mexico
New Mexico
Colorado
Colorado
Wyoming
Wyoming
SIC Ho.
1092
1092
1031
1031.102)
1031,1021
1031,1021
1031,1021
1031,1021
1031
102 1 , 1099
1021
1021
1021
1021
1021
1021
1021
1031
1099
1099
1099
1031
1021
1021,1031
1099
1094
1031
1094
1094
10V4
1094
1094
1094
Mir
Ho.
1
1
1
1
2
1
3
1
I
1
I
1
1
1
1
1
1
'l
2
1.
1
1
1
I '
1
3
5
1
4
1
3
1
1
ies Concentrator
Type Mo.. Typ«
Open pit ' I Roast and Condense
Open Pit • 1 Roaat and Condense .
Underground 1 Flotation
Underground
Underground
Underground
Underground
Underground
Underground
Open Pit
Open Pic
Open pit
In Situ
Underground
Open pit
Underground
Open Pit
Open Pit
Underground
Flotation :
Flotation
Flotation
Flotation
Flotation
Flotation
2 Flotation, 1. Laach Precipitation
1 Flotation, 1 Leach Precipitation
Flotation
Usch - precipitation
Flotation
Leach - precipitation
Flotation
Leach - precipitation
Leach - Electrowinalng
Flotation
Leach - Solvent Exchange - Elactrovinnlng
Flotation
Open Pit 1 Acid Leach - Solvent Extraction - Hydrolysis
Dredge 2 Electrostatic - Magnetic Separation -
Uet
Dredge 2 Dry - Electrostatic - Magnetic Separation
Underground 1 Flotation
Underground
Underground 2 Flotation (Antimony Plant)
Open Pit 1 Salt roast - Water Leach Precipitation
Underground 1 Flotation
Open Pit Alkaline Laach - Precipitation
Underground Roast - Leach - Filtration - Precipitation
Underground Leach - Autoclave - Filtration * precipitation
Underground Leach - Ion Exchange - Precipitation
Open Pit Leach - Solvent Extraction
Open Pit Leach - Ion Exchange
-------
TABLE B-2
QUESTIONNAIRES RECEIVED
I.
2.
3.
4.
5.
6.
7.
8.
9.
10.
11.
12.
13.
14.
15.
16.
17.
18.
19.
20.
21.
22.
23.
24.
25.
26.
27.
28.
29.
30.
31.
32.
33.
34.
35.
36.
37.
38.
39.
40.
41.
42.
43.
44.
45.
Company Name
Idarado Mining Company
Union Carbide
Conqulsta Project
Phelps Dodge
Phelps Dodge
Phelps Dodge
Phelps Dodge
Bunker Hill Company
Cotter Corporation
Continental Copper
Comlnco America, Inc.
Utah International, Inc. •
Cities Service Company
,
Amax Lead Company
Petrotomlcs Company
Day Mines, Inc.
Exxon Company
Kennecott Copper
Cyprus Pima Mining Company
The Anaconda Company
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
American Smelting and Refining
Copper Range Company
The New Jersey Zinc Company
The New Jersey Zinc Company
The New Jersey Zinc Company
The New Jersey Zinc Company
Company
Company
Company
Company
Company
Company
Company
Company
Company
Gulf Resources and chemical Corporation
Kennecott Copper
Magma Copper Company
Magma Copper Company
Union Carbide Corporation
Union Carbide Corporation
Hecla Mining Company
Hecla Mining Company
Hecla Mining Company
Cyprus Mines Corporation
Brush Wellman
Location
Telluride, Colorado
Hot Springs, Arkansas
Falls City, Texas
Morencl, Arizona
Blsbee, Arizona
Ajo, Arizona
Tyrone, New Mexico
Kellogg, Idaho
Canon City, Colorado
Oracle, Arizona
Blxby, Missouri
Rlverton, Wyoming
Copperhlll, Tennessee
Boss, Missouri
Casper, Wyoming
Wallace, Idaho
Casper, Wyoming
Blngham Canyon, Utah
Tucson, Arizona
Grants, New Mexico
Imrael, Tennessee
Jefferson County, Tennessee
Wallace, Idaho
Sahuarita, Arizona
Sahuarita, Arizona
Lake Hurst, New Jersey
Dem ing, New Mexico
Casa Grande, Arizona
Silver Bell, Arizona
White Pine, Michigan
Frledensvllle, Pennsylvania
Ivanhoe, Virginia
Jefferson City, Tennessee
Gllman, Colorado
Metallne Falls, Washington
Ray, Arizona
San Manuel, Arizona
Superior, Arizona
Gas Hills, Wyoming
Uravan, Colorado
Burke , Idaho
Mullan, Idaho
Casa Grande, Arizona
Bagdad , Arizona
Delta, Utah
SIC No.
1031,1021
1094
1094
1021
1021
1021
1021
1031
1094
1021
1031
1094
1021,1031
1031
1094
1031
1094
1021
1021
1094
1031
1031
1021
1021
1021
1099
1031
1021
1021
1021
1031
. 1031
1031
1031
1031
1021
1021
1021
1094
1094
1031
1031
1021
1021,1031
1099
Mine
Name
Idarado
North Wilson, East
. Wilson, Spauldlng
Conqulsta Project
Morencl -Me ten If
Copper Queen -Lavender Pit
New Cornelia
Tyrone
Bunker Hill - Crescent
Schwartzwalder
Oracle Ridge Project
Magmont
Lucky Me
Calloway-Boyd, Cherokee
Tennessee-Eureka
Bulck
Shirley Basin
Dayrock
Highland Operations
Blngham
Ptma
Jackplle-Pagvate
Young-Coy- Immel
New Market
Galena
San Xavler
Mission Unit
Manchester
Ground Hog -Dem ing
Saca ton Unit
Sliver Bell
White Pine
Frledensvllle
Austlnville-Ivanhoe
Jefferson City
Eagle
Pend Orel lie
Ray
San Manuel •
Magma " -
Gas Hills District
Uravan
Star Unit Area
Lucky Friday
Lakeshore Project
Cyprus
Topaz
Concentrator
No.
1
3
4
2
2
I
1
2
1
1
1
1
5
1
1
1
1
1
1
2
3
I
I
1
1
1
2
I
1
1
1
2
1
1
1
1
1
1
2
3
1.
I
I
I
1
Type No. Type
Underground
Open Pit
Open Pit
Open Pit
l-Open Pit
I -Underground
Open Pit
Open Pit
Underground
Underground.
Underground
Underground
Open Pit
Underground
Underground
Open Pit
Underground
Open Pit
Open Pit
Open Pit
Open Pit
Underground
Underground
Underground
Underground
Open Pit
Open Pit
Dredge
Underground
Open Pit
Open Pit
Underground
Underground
Underground
Underground
Underground
Underground
Open Pit
Underground
Underground
' Open Pit
Underground
Underground
Underground
Underground
Underground
Open Pit
1
I
I
2
I .
1
2
1
1
1
1
1
I
1
I
1
1
3
I
1
I
1
1
1
1
I
I
1
2
1
1
1
1
I
1
3
1
I
I
1.
1
1
1
1 •
1
Flotation
Saltroast - Solvent Extraction
Hydrome ta 1 lurgy
Flotation - Leach Precipitation
Flotation
Flotation
Flotation - Leach Precipitation
Flotation
Leach - Auto Clave
Flotation
Flotation
Acid Leach Ion Exchange
Flotation
Flotation
Leach - Solvent Extraction
Flotation
Acid Leach Solvent Extraction
Flotatlon(2) - Leach
Precipitation
Flotation
Leach - Resin in Pulp
Flotation
Flotation
Flotation
Acid Leach Precipitation
Flotation
Gravity
Flotation
Flotation
Flotation - Leach
Precipitation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation - Leach Precipitation
Leach Electrowlnnlng
Flotation
Flotation
Leach ion Exchange Precipitation
Leach Ion Exchange Precipitation
Flotation
Flotation
Flotation
Flotation
Hot Leach - Solvent Extraction
Precipitation
-------
Revised 10/15/74
(SAMPLE QUESTIONNAIRE)
WASTE DISPOSAL BACKGROUND INFORMATION
(for MRI Project No. 3952-D)
1. Corporation Name:
Mailing Address:
Mine Names and Locations: 1.
2.
3.
4.
2. Person to contact regarding information supplied herein:
Mr./Ms. . - '''... '. .
Address , .,.'. „ '...-.;.
Phone
322
-------
3. Final Concentrator Product(s):
(primary) ' TPY* ; Grade,7._
(byproduct) • TPY ; Grade, 7._
(byproduct) TPY __. ; Grade,%_
4. Mine Operations:
(a) Type: Open Pit ; Underground ; Age of Mine_
(b) Water (from mine operation only) discharged to tailing ponds--describe
source(s) '
Gallons per day:_ ; pH_
(c) Solids handled:
Ore- TPY
Solid Waste (waste rock and/or overburden) TPY
Percent waste rock • • ; Percent overburden
5. Leaching Operations (if applicable):
(a) Type: Dump ; Heap ; Vat ; Waste
Other (specify)
(b) Solvent: Acid ;• Alkali ; Cyanide
t
Other (specify) .
6. Concentrating Operations:
(a) Grind: Percent above 100 mesh ; Percent below 100 mesh
(b) Process used (check); Froth Flotation ; Gravity ;
Other (specify)
(c) Volume of water in circuit: gallons per minute
(d) Water discharged to tailings ponds--describe source(s)
Gallons per day:
*TPY"tons per year 323
-------
7. Solid Waste Disposal:
Source #1
TPY (dry short tons)
' Tailing impoundment ; Area (Acres)
Mine backfill ; Other (specify)
Source #2
TPY (dry short tons)
Tailing impoundment.... . ; Area (Acres)_
Mine backfill ; Other (specify)
Source #3
TPY (dry short tons) .......
Tailing impoundment ..; Area (Acres)
Mine backfill ; Other (specify)
Source #4
TPY (dry short tons)
Tailing impoundment ; Area (Acres) . .
Mine backfill ..._ ; Other (specify)
(Indicate additional sources on back of sheet)
8. What waste disposal techniques do you use for all of your wastes? Please
check techniques used, and percentage of the total waste treated by the
method. ,
Check Percent of
Method Total Waste
Burial (deep) . ... _ui_«^
i
Burial (surface) . . ..
Detoxification (chemical and biological) ._ . . .
Recovery and Reuse , ._ . . .
Deep Well Injection _ . . . ...
Mine Disposal . . .. ._
Open Burning .. . . :
Incineration _^_ ._ .
Open Dumping .
Tailing Pond __^_ _^_ ^_
Other
-------
9. Do you handle your own waste disposal? Yes No
10. Percent of solid waste sold ; used for
11. Liquid discharged to tailing ponds or other land disposal (characteristics
in milligrams per liter).
.. Total
Suspended Dissolved IDS, Major
Solids Solids Constitutent
Discharge #1 '
•
Discharge #2 '•
Discharge #3 •. .
Discharge #4 '
(list additional discharges on back of sheet)
12. Solid Waste Characteristics: (see next page)
13. Approximate cost for treatment and land disposal of wastes $ / tons of waste
325
-------
11. Solid Waste Characteristics: (attach additional sheets for-more than four sources)
Indicate, by a check mark, which of the following are analyzed for: In the land-disposed solid wastes; show the sensitivity level (In parts per
million) of the analytical technique used, and the teat frequency (D = Dally, W " Weekly, M • Monthly, Y » Yearly).
Radioactivity
Asbestos Arsenic Beryl limn Cadnium CI.. .iniium Coppc* Cyanldus. Mercury Pesticides Selenium Zinc Lead . ( PC 1) **
Source tfl (Analyzed for, check) . _
Concentration, ppra*
Average ______ . _______ ______
Range •' .
Test Sensitivity, ppm __
Teat Frequency . • __. . .
Source 02 (Analyzed for, check)
Concentration, ppm
Average
B«ng«
Teat Sensitivity, ppm
Test Frequency
Source 03 (Analyzed1 for, check)
Concentration, ppm
Average _____
Range
Teat Sensitivity, ppra
Test Frequency.
Source 04 (Analyzed for, check) ___
Concentration, ppm
Average
Range
Test Sensitivity, ppm
Test Frequency
* Parts per million
** plco curie per liter
-------
APPENDIX C
GEOLOGY AND MINERALOGY
327
-------
GEOLOGY AND MINERALOGY
Copper Ores
Copper ores occur in rocks of nearly all kinds and ages and in
ore deposits of many types. All copper ores are secondary deposits and can
be classified geologically into a multitude of types, but in this discussion
they will be grouped into two classes—disseminated "porphyry," and all
others. The disseminated deposits, usually containing less than 1 percent
copper, are large masses of more or less altered and decomposed igneous
rock, in which copper is uniformly, though sparsely, distributed in the form
of small particles and veinlets; the rock usually is of a granitic type and
porphyritic texture, hence the term "porphyry copper." Other types of
deposits include bedded deposits, vein deposits, massive replacement
deposits, etc. The practical importance of the porphyry type is that its
size and physical character encourage application of large-scale, low-cost
mining, and metallurgical methods.
The principal copper minerals as they occur in each state where
copper is mined are listed in Table C-l. Of the sulfide ores listed,
bornite, chalcopyrite, and enargite .are considered the "primary" minerals
which are redeposited by igneous processes deep in the earth's crust.
Minerals such as covellite and chalcocite were largely formed as "secondary"
deposits of copper leached from the sulfides close to the surface and
precipitated near the water level. The oxide minerals such as chrysocolla,
malachite, and azurite were formed through oxidation of the surface sulfides.
Native copper is chiefly located in the oxidized zones of an ore deposit;
however, this is not the case in the major native copper deposits in
Michigan, which are generally considered to be "primary" in source.1.' The
copper at White Fine, Michigan, from which a little more than 5 percent of
the United States primary copper currently is produced, is a large stratiform
ore body 1.2 to 7.6 meters thick and several kilometers across. The present
ore column, containing about 1.2 percent as chalcocite and native copper, is
confined to the basal beds of the Nonsuch Shale. This upper Keweenawan (upper
Precambrian) formation is a distinctive gray siltstone unit about 183 meters
thick, overlying and overlain by thick red-bed sequences. All these rocks
contain abundant volcanic debris, but the latest known igneous activity within
80 kilometers antedated the Nonsuch Shale. The rocks are unmetamorphosed and
only moderately deformed, mainly by tilting and faulting.-
328
-------
SUMMARY OF MINERALOGICAL SYSTEMS
State
Arizona
and adjacent
New Mexico
Tennessee
Utah
Mineral
Ankerlte
Barite
Bornite (bn)
Chalcocite (cc)
Chalcopyrite (cp)
Tennantite
Enargite (en)
Galena (gn)
Tetrahedrite
Hematite (hm)
Magnetite (mg)
Pyrite (py)
Quartz (qz)
Sphalerite. (sl)
Azurite
Chrysocolla
Cove 1 lite
Serpentine
Cuprite
Ralloyslte
Hemimorphite
Llmonite
Malachite
Manganite
Olivealte
Ps Home lane
Tremolite
Anhydrite
Pyrrhotlte
Cubanlte
Molybdenite
Gypaum
Pyrrhotlte
Pyrtte
Chalcopyrlte
Quartz
Calclte
Sphalerite
Magnetite
Sphalerite
Quartz
Chalcedony (opal)
Calclte
Montmor 1 1 lonold
COPPER ORE AT
Chemical
Compos it loo
2CaC03-MgC03-
FeC03
BaS04
CusFeSA
Cu2S
CuFeS2
5Cu2Sl2ZnS-2As2-
S3
Cu3AsS4
PbS
3Cu2S-SloS2-Fe-
Zn-Ag
Fe203
Fe304
FeS2
S102
ZnS
2CuC03-Cu(OB)2
CuS 105-^0
CuO
3Mg02(Sl02)'2H20
Cu20
Al203.Si02.H20
Zn4Sl207(OH)2'H20
Hydrated ion
oxides
CuC03-Cu(OH)2
Mn203-H20
4CuO-AS205'H20
Mn02-Ba-K2-Na20-
H20
Ca2Mg5Si8022-
CaS(0^lF)2
FeuS12
CuFe2S3
MoS2
CaS04-2H20
PenSi2
FeS2
CuFeS2
St02
Cu2S
ZnS
Fe304
ZnS
S102
S102-nH20
CaC03
(Ca.Na)Al-Mg.Fe.
MINES (1021)
Specific
Gravity
2.9-3.2
4.50
5.06-5.08
5.5-5.8
4.1-4.3
4.37-4.49
4.4-4.5
7.57-7.59
4.4-5.1
5.26
5.16
5.02
2.65
3.9-4.1
3.4-3.8
2-2.24
4.6
2.5-2.65
5.8-6.2
2.0-2.2
3.4-3.5 .
3.6-4
4.00
5.16
4.1-4.4
3.3-4.7
3.0
2.8-3.0
4.58-4.65
4.03-4.18
4.62-4.73
2.32
4.58-4.65
5.02
4.1-4.3
2.65
2.6-2.71
3.9-4.1
5.26
3.9-4.1
2.65
1.73-2.16
2.6-2.71
2-3
Solubility
Reagents
s. a.
s. a.
s. HN03
s. HN03
dec. HK03
d HN03
s. HN03
dec. HN03
dec. HN03
s. a.
s. HC1
s. HN03
a. only HF
s. HC1
s. a.
s. a.
s. HNO
dec. a.
s. HC1
1.
gel. a.
s. HC1
s. a.
3. HC1
a. HNOy
s. HC1
-
s. HC1
dec. a.
s. a.
s. a.
s. HC1
dec. a.
s. HK03
dec. HN03
s. only HF
s. HC1
3. HC1
s. a.
s. HC1
s. only HF
- .
s. HC1
s. s. a.
Sl.Al-Au.Ag.O-H20
Chlorite
Wollastonlte
Pyrite
Chalcopyrlte
(MgAi-Fe)12'(SlAl)
8020 'OH16
CaO-Sl02
FeS2
CuFeS2
-2.6-3.3
2.8-2.9
5.02
4.1-4.3
sl. a. a.
dec. HC1
s. HN03
dec. HN03
Solubility
Water
1.
Sl. 3.
sl. s.
1.
sl. a.
1.
I.
i.
1.
i.
i.
1.
1.
dec.
1.
1.
1.
1.
sl. dec.
sl. s.
1.
i.
sl. dec.
sl. s.
sl. s.
i.
i.
1.
i.
1.
1.
i.
I.
sl. s.
sl. s.
Hardness
3.5-4
3-3,3
3
2.5-3
3.5-4
3-4
2.5-2.75
3-4
5-6
5.5-6.5
6-6.5
7
3.5-4
3.5-4
2
1.5-2
2.5-3.5
3.5-4
1-2
5
5-5.5
3.5-4
5.5-6.5
3
5.7
5-6
3-3.5
3.5-4.5
3.5
1-1.5
1.5-2
3.5-4.5
6-6.5
3.5-4
7
3
3.5-4
5-6
3.5-4
7
~ 6
3
1-2
2-3
4.5-5
6-6.5
3.5-4
329
-------
TABLE C-l (Continued)
Michigan
Montana
,
Mineral
Ga lena
Bornite
Cubanite
Molybdenite
Arsenopyrlte
Stannite
Garnet
TremoHte
.
Talc
Serpentine
Epidote
Gypsum
Tourmaline
Graphite
Phodonite
Apatite
Copper
Copper
Silver
Chalcocite
Bornite
Chalcopyrite
Covellite
Digenite
Pyrite
Greenoekite
Sphalerite
Ga lena
Calclte
Quartz
Dolomite
Barite
Chlorite
Mom-iaartilonite
Muscovite
Orthociase
Epidote,
^
Hematite
Acanthi te
Aikinitc
Alabandite
Alunite
Andorite
Ankerite
Apatite
Arsenopyrite
Barite
Anhydrite
Chemical
Composition
PbS
CusFeSy
CuFe2S3
MoS2
FeA5S
Cu2FeSn4
Ca-Fe.Al-Mg(S104)-
X
Ca2MgsSl8022-
(OH1F)2
3MgO*4S102-H20
3Mg02(S102).
2H20
Ca2'FeAl2Si3Oi2-
- (OH)
CaS04-2H20
Al-B-S102-Lt-Fe.
Na-Mg
C
(MnFeCa)Si03
(CaF)Ca4(P04)3
Cu
. Cu
Ag
Cu2S
Cu5FeS7
CuFeS2
CuO
. Cu9S5
FeS2
CdS
ZnS
PbS
CaC03
S102
CaMg(C03)2
BaS04
(MgAlFe)12(SlAl)-
(Mg,Ca)0-Al203-
H2-KA13(S104)3
KA1S1308
Specific
Gravity
7.57-7.59
5.06-5.08
4.50
4.62-4.73
5.9-6.2
4.3-4.5
3.5-4.25
3.0
2.6-2.8
2.5-2.65
3.4-3.5
2.32
2.9-3.2
2.09-2.23
3.57-3.76
3.17-3.23
8.95
8.95
10.5
5.5-5.8
5.06-5.08
4.1-4.3
4.6
5,5
5.02
4.5-5
3.9-4.1
7.57-7.59
2.6-2.71
2.65
2.86
4,50
2.6-3.3
2-3
2.76-3
2.55-2.63
H2K(t%,Fe)3(Al,Fe)-3.4-3.5
(S104)3
F6203
Ag2S
PbCuBiS3
MnS
KA13(OH)6(S04)2
PbAgSb3S$
Ca(Fe.Mg) (C03)2
Cae(F-C10H)(K>4)3
FeAsS
BaSo4
CaSo4
5.26
7.2-7.3
6.1-6.3
4.1
2.6-2.75
5.33-5.37
2.9-3.2
3.17-3.23
5.9-6.2
4.50
2.3-3.0
Solubility
Reagents
dec. HN03
s. HN03
a. a.
s. a.
s. a.
s. a.
1.
i.
s. a.
dec. a.
1. a.
s. HC1
t.
i.
i.
s. HC1 and
HN03
s. HN03
s. HN03
s. a.
s. HN03
s. HN03
dec. HN03
s. a.
s. a.
s. HN03
a. HC1
s. aci
dec. HN03
s. HC1
s. HF
-
s. a.
si. s. a.
si. s. a.
i.
i.
i.
s. a.
s. HN03
dec. HN03
s. HC1
s. H2S04
si. s. a.
s. a.
s. a.
s. a.
s. a.
s. HC1
Solubility
Water
_
si. s.
1.
1.
si. s.
i.
i.
1.
1.
i.
i.
i.
i.
i.
i.
i.
1.
i.
i.
-
si. s.
si. s.
i.
i.
si. s.
i.
1.
-
1.
i.
-
i.
1.
i.
i.
i.
1.
1.
1,
i.
i.
i.
i.
si. s.
i.
si. s.
i.
si. s.
Hardness
2.5-2.75
3
3*3.5
1-1.5
5.5-6
4
6.5-7
5-6
1
2.5-3.5
6
1.5-2
7-7.5
1-2
5.5-6.5
4.5-5
2.5-3
2.5-3
2.5-3
2.5-3
3
3.5-4
1.5-2
2.5-3
6-6.5
3-3.5
3.5-4
2.5-2.75
3
7
3.5-4
3-3.5
2-3
1-2
2-2.25
6-6.5
6
5-6
2-2.5
2-2.5
3.5-4
3.5-4
3-3.5
3.5-4
4.5-5
5.5-6.0
3-3.5
3-3.5
330
-------
TABLE C-l (Concluded)
State
Montana
Mineral
Bornite
Calclte
Chabezite
Chalcocite
Chalcopyrite
Covelllte
Digenite
Dolomite
Enargite
Ferberite
Fluor ite
Galena
Greenockite
Gypsum
Helvite
Hematite
Heulandite
Huebnerlte
Magnetite
Marcasita
Molybdenite
Orthoclass
Polybaaite
Proustite
Pyrite
Quartz
Rhodochrosite
Rhodonite
Rutile
Scheelite
Siderite
Sphalerite
Stephanite
Stilbite
Strooeyerite
Tenhantite
Tetrahedrite
Uraninite
Wavellite
Wolframite
Wurtzite
Chemica 1
Composition
Cu FeS
CaC03
CaAl2Sl4Oi2'
6H20
Cu2S
CuFeS2
CuS
CujSj
CaMg(C03)2
Cu3AsS4
FeN04
CaF2
PbS
CdS
CaS04-2H20
Mn4SSi3Be30|2
Fe203
6H20
MnWD4
Fe304
FeS2
MoS2
KAlSi-jOg
(Ag,Cu)l6Sb2Su
Ag3AaS3
FeS2
S102
MnC03
(Mn,Fe,Ca)Si03
Ti02
Ca(W,Mo)04
FeC03
ZnS
AgjSbS4
CaAl2Si7018'
7H20
CuAgS
Cu12As4Si3
Cu,2Sb4Si3
UDj'Th.Zr.La.Y,-
Pb.Hc
A13(OH)3(P04),'
5H20 '
.MnWD4
ZnS
Specific
Gravity
5.06-5.08
2.6-2.71
2.1
5.5-5.8
4.1-4.3
4.6.
5.5
2.86
4.4-4.5
7.51
3.18
7.57-7.59
4.5-5
2.32
3.2-3.4
5.26
2.1-2.2
7.12
5.16
4.9
4.62-4.73
2.55-2.63
6-6.2
5.57
5.02
2.65
3.7
3.57-3.76
4.23-5.5
6.1
3.96
3.9-4.1
6.2-6.3
2-2.2
6.1-6.3
4.37-4.49
4.4-5.1
9-9.7
2.32
7-7.5
3.98
Solubility
Reagents
a. HN03
s. HC1
d. HC1
s. HN03
dec. HN03
a. HN03
s. a.
3. HNOj
d. a.
s. a.
dec. HN03
s. HC1
s. HC1
s. HC1
s. a.
d. HC1
d. HC1
s. HC1
s. con. HN03
s. a.
1.
dec. HN03
s. HN03
a. HN03
s. only HF
s. h. HC1
sl. s. HC1
1.
d. HC1
s. h. HC1
s. HC1
s. h. HN03
dec. HC1
s. HN03
d. HN03
d. HN03
s. a.
s. HC1
dec. a.
s. a.
Solubility
Water
al. s.
-
i.
- '
sl. s.
-
i.
*
1.
sl. s.
-
i.
i.
i.
i.
3l. S.
i.
-
i.
i.
i.
i.
i.
Sl. 3.
i.
i.
i.
i.
1.
i.
i.
1.
sl. s.
i.
-
-
sl. s.
s.
i.
I.
Hardness
3
3
4.5
2.5-3
3.5-4
1.5-2
2.5-3
3.5-4
3
4-4.5
4
2.5-2.75
3-3.5
1.5-2
6
5-6
3.5-4
4-4.5
5.5-6.5
6-6.5
1-1.5
6-6.5
2-3
2-2.5
6-6.5
7
3.5-4
5.5-6.5
6-6.5
4.5-5
4-4.5
3.5-4
2-2.5
3.5-4
2.5-3
3-4
3-4
5.5
3.5-4
5-5.5
3.5-4
331
-------
Pyrite is the characteristic sulfide of the Nonsuch Shale. The
principal sulfide mineral of the cupriferous zone at the base of the
formation is chalcocite. The succession of minerals outward and upward
from the center of the cupriferous zone is native copper, chalcocite,
bornite, chalcopyrite, and pyrite. The bornite and chalcopyrite are
confined to a narrow fringe at the margin of the cupriferous zone. This
pattern of mineral zones is believed to represent reaction between sygentic
pyrite and introduced copper.
Butte District. Montana
Butte, Montana, is one of the world's outstanding examples of
metal and mineral zoning. Three crudely concentric zones were delineated
by Reno Sales in 1913 (2. p. 58): (1) A Central Zone, "occupying an area
of altered granite in which the ores are characteristically free from
sphalerite and manganese minerals"; (2) An Intermediate Zone, "in which the
ores are predominantly copper, but are seldom free from sphalerite"; (3) A
Peripheral Zone, "in which copper has not been found in commercial
quantities."!/
The copper zone can be associated with both the Intermediate and
Central Zones. Chalcopyrite is an important mineral at the outer edge of
the copper zone in every part of the district. However, deeper within the
zone, the arrangement of copper minerals is not smoothly symmetrical with
respect to the margin.
The pattern of distribution of Cu-Fe-S series minerals in the
large veins of the copper zone is not symmetrical with respect to the
metal zones. There is a fringe of chalcopyrite-bornite all around the
copper front. However, within the copper zone, pyrite-chalcocite-enargite-
covellite-digenite assemblages are prevalent near its eastern edge from
the 1,158-mef.er level to the surface, whereas deep in the large Anaconda
veins in the western part of the copper zone, chalcopyrite is dominant.
Here the chalcopyrite grades upward into bornite-chalcocite ores and
eastward into the high-sulfur assemblages of the Leonard mine.—
Second, copper ores are mined in the Berkeley pit in the
southeastern part of the district. At the present time, they account for
about half of Butte1s production of copper.
332
-------
Nevada
The Yerington, Lyon County, Nevada, ore body is a disseminated
porphyry copper deposit. It is overlain with gravel varying from 0 to 61
meters in thickness. The basement rock is granodiorite intruded by a quartz
raonozite porphyry. While ore is found in both, most of the ore body occurs
within the quartz monozite. Chrysocolla is the chief copper mineral. Most
of it is well-disseminated, but the remainder occurs in seams and fractures. ,
The zone of secondary enrichment is usually small, only a minor amount of
chalcocite having been found below the oxide zone. In the sulfide zone, ',
chalcopyrite is the principal copper mineral.
The Copper Canyon porphyry copper deposit, consisting of the
east and west ore bodies is mainly in fractured and altered sedimentary
rocks of the Pennsylvanian age, and adjacent to a granodiorite laccolith.
In the east ore body, most ore occurs as hypogene sulfides that replace
former hematite and calcite bearing zones of the lower member of the Middle
Pennsylvanian Battle Formation. In the west ore body sulfide-rich raanto
deposits occur generally closer to the granodiorite as replacement of the
Pennsylvanian Pumpernickel Formation.— The hypogene minerals include
pyrite, chalcopyrite, pyrrhotite, and marcasite, with lesser concentrations
of arsehopyrite, native gold and silver, and traces of sphalerite, molybdenite,
and galena.
Arizona
The sedimentary rocks in the Pima mine area in the Tucson District
are thought to be of Permian age. The sequence is as follows: a rusty
brown quartzite; a thin-bedded, serpentined, dolomitic limestone; a highly
altered (clay-garnet) limestone, these overlain by a massive thick-bedded
arkosic quartzite which varies from rusty brown to light gray.
The high-grade copper sulfide ore body is a replacement of
chalcopyrite in the clay-garnet material. Within these areas of .
dissemination, some pyrite is present, both as small veinlets and as
disseminated material. Scattered pods of zinc sulfides have been found
within the ore zone. Some molybdenite occurs as plating along the minor
slips within the ore zone. The main ore minerals within the oxidation
zone are native copper and chalcocite. Minor amounts of chrysocolla and
azurite are present.
The host rock of the massive chalcopyrite ore bodies is an
intensely altered limestone consisting of a mixture of kaolin and small
garnets, with occasional remnants of calcium carbonate.
333
-------
Nearly all of the ore production from the Twin Buttes mines
came from a northwesterly trending mineral zone extending through the
Copper Glance, Copper Queen, and Copper King mines.
The geology of the Twin Buttes region is composed of quartz
monzonite porphyry. The porphyry includes limestone, siltstone, and
clastic rocks. The clastic rocks are a thick sequence of interbedded
arkoses, arkosic-like pyroclastics, meta-siltstones and siliceous tuffs
which uniformly overlie the limestones, siltstones, and quartzites on
the south.
On the southwest and northeast, porphyry contacts the sedimentary
rocks. These sedimentary rocks are irregularly altered and mineralized.
The quartzites are weakly mineralized, 'but the silicated, garnetized and
chloritized limestones and siltstones show the strongest copper
mineralization. The clastic rocks are bleached and mineralized in an
extensive area on the south with better copper grades occurring to the
southeast where they are in contact with the porphyry.
Unlike many of the porphyry copper deposits in the area, the
Twin Buttes quartz monzonite porphyry contains scattered, overall weak
mineralization.
Chalcopyrite is the most abundant and widespread copper sulfide
mineral, but the mineral suite includes secondary chalcocite and covellite
and minor bornite. Molybdenite occurs in variable but economically
important amounts throughout the mineralized zone. Some sphalerite and
galena occur locally. Trace amounts of gold, insignificant silver, and
widespread pyrite occur, and in places magnetite is abundant. The two
iron minerals are found mostly in the mineralized limestone and siltstone
beds.
( i
There are large tonnages of oxidized rock of good copper content
in the oxide zone, and in places secondary copper sulfides substantially
enhance sulfide ore grade.
Copper minerals in the oxide zone consist principally of
chrysocolla and tenorite or melaconite. Cuprite and native copper are
locally plentiful, and.there are some brachantite and minor malachite.
The Magma mine at Superior, Arizona, has produced over 11,790,000
metric tons of ore yielding 6.8 x 10° kilograms of copper. It is a
mesothermal deposit, and although the bulk of the ore has come from the
Magma vein, most of the recent production has come from replacement
deposits in limestone.
334
-------
The mine lies in a Precarabrian and Paleozoic sequence of rocks
that includes schist, diabase, quartzite, shale, and limestone; nearby,
these rocks are intruded by igneous bodies of Mesozoic age that were
probably the source of the ore-bearing fluids. The area is extensively
faulted. There are two principal strike faults of large displacement and
numerous smaller steep and flat faults that complicate the prediction of
ore occurrence.
The principal copper minerals in the vein are chalcopyrite,
bornite, enargite, tennantite, chalcocite, and digenite; these minerals
vary in abundance according to the district zoning pattern. The upper
part of the mine yielded considerable zinc as sphalerite, and the oxidized
zone was rich in silver. The chief gangue materials are pyrite, quartz,
and hematite. In the limestone replacement bodies, chalcopyrite and bornite
are the chief ore minerals.
Ducktown District, Tennessee
The Ducktown ore deposits have eight massive, sulfide ore bodies
that occur in highly folded and metamorphosis graywacke, graywacke conglomerate,
mica schists, chlorite-garnet schist, and staurolite schist of Precambrian age.
The ore deposits are tabular bodies that have been extensively folded,
generally conform to the enclosing rocks, but show local complex differences.
The ore deposits are composed principally of the minerals
pyrrhotite, pyrite, chalcopyrite, sphalerite, and magnetite. Gangue minerals
are quartz, calcite, actinolite, tremolite, hornblends, garnet, and masses of
schistose wall rock. Three fault systems are present that have influenced
ore control and ore body configuration. Hydrothermal effects are evident
in wall rock alteration. Evidence of retrogtade metamorphism is extensive in
the wall and ore-associated rock. Preliminary isotopic age dates have
indicated four possible metamorphic events during the Paleozoic Era.
t
Ore genesis is considered to be hydrothermal replacement of
receptive beds, predominantly highly calcareous zones, quartzitic zones,
and brecciated shear zones. Ore deposition is thought to have occurred
during the Devonian Period, with later mobilization and recrystallization
of the ore and gangue during Middle and Late Paleozoic times.—'
335
-------
Bingham District., Utah
The Bingham mining district, located in the Oquirrh Mountains
about 32 kilometers southwest of Salt Lake City, Utah, is well known as
the site of the Kennecott Copper Corporation's porphyry copper deposit
situated in and near the Utah Copper or Bingham stock.
Two main types of intrusive igneous rock are present in the
district. The Last Chance stock and the southern and eastern portions
of the Bingham stock are composed of a relatively dark equigranular-to-
seriate porphyritic rock which probably averaged quartz monzonite in
composition before alteration and has been called "dark porphyry" and
"granite." A second body called "granite porphyry" or "light porphyry"
intrudes the north side of the "dark porphyry." This intrusive has the
typical porphyry texture with an aphanitic quartz-feldspar groundmass as
found at other porphyry copper deposits, and it appears to form a center
for the mineralization patterns of the district. Extrusives, although
present, contain no ore.
The intrusives consist of granite, quartz, monzonite, granodiorite,
quartz diorite, granite porphyry, quartz monzonite porphyry, and biotite
quartz latite porphyry. The extrusives consist of andesite, andesite
porphyry, andesite breccia, agglomerates, ash: deposits, and rarely,
basalt.U
The mineralogy of the district is comprised of chalcopyrite,
bornite, and molybdenite, with low pyrite. The lead-zinc ore bodies are
likewise massive coarse-grained sulfides of galena, sphalerite, with
varying amounts of pyrite and with some copper, gold, and silver. While
the galena itself contains silver, much of the silver' is found as
argentiferous tetrahedrite-tennantite.
Gangue minerals associated with the ore are calcite, quartz,
silica, saponite, montmorillonoid clay, talc, chlorite, wollastonite,
diopside, and mica. Calcite and quartz are by far the most common.
Arizona and Adjacent New Mexico
Arizona and western New Mexico contain two-thirds of the leading
copper mines in the United States. Production of molybdenite, lead, zinc,
and by-product gold and silver is important.
336
-------
Precambrian ore deposits are largely limited to the Yavapai
Series in Central Arizona, and most of the metal production has come from
massive sulfide .deposits, dominantly pyrite but containing different amounts
of chalcopyrite, tennantite, sphalerite, and galena.
Most of the important ore deposits are of Mesozoic and early
Tertiary age, dominated by the porphyry copper deposits spatially related
to stocks, plugs, sills, or dikes of quartz-bearing porphyritic to granular
rocks. Host rocks of these deposits include the associated intrusive rocks,
Precambrian schist and granitic rocks, Paleozoic limestone, and Mesozoic
arkose and siltstone. Chalcopyrite and pyrite are the dominant hypogene
sulfide materials in the porphyry copper deposits, accompanied by minor
molybdenite. Supergene sulfide minerals are chalcocite and minor covellite.
Supergene oxide-copper minerals are important locally.
Intense hydrothermal alteration accompanied the metallization in
all of the porphyry copper deposits. Five types are recognized: (1)
propylitic; (2) argillic; (3) potassic; (4) quartz-sericite (without clay);
and (5) lime silicate.U
Breccia pipes, veins, and limestone replacements account for
some important deposits of Mesozoic and early Tertiary age, and presumably
these are genetically related to the various intrusive rocks of this age.
Lead and Zinc Ores
The common lead minerals are either sulfide ores (galena, PbS),
or oxide or carbonate ores (anglesite, PbSO,, and cerussite, PbCO_). Galena
is the most abundant lead mineral found in deposits that have been exploited
in the United States. Galena is commonly associated with zinc, silver, gold
and iron minerals. However, in a few districts the ore bodies are
characterized by very simple mineralization, with the lead mineral present
to the virtual exclusion of other ore minerals. A noteworthy example is
southwestern Missouri, where the lead ores are generally composed of galena
in an essentially nonmetallic gangue, with relatively little silver, zinc,
or copper present. Table C-2 shows the mineralogy of the major lead-zinc
districts.
337
-------
TABLE C-2
Missouri
Colorado
SUMMARY OF MINERALOGICAL SYSTEMS
Mineral
Pyrite
Sphalerite
Galena
Pyrrhotite
Chalcopyrite
Magnetite
Hematite
Bornite
Willemite
Quartz
Dlopaide
Tremollte
Serpentine
Talc
Barite
Anhydrite
Chlorite
Ilvaite
Grossularite (garnet)
Anthophyllite
Clinozoislte
Sericite
Albite
Dolomite
Galena
Sphalerite
Chalcopyrite
Bravo ite
Pyrice (rare)
Marcaslte
Bornite (sparce)
Millerite (sparce)
Dolomite
Calcite
Die kite
Quartz
Covelllte
Chalcocite
Malachite
Dolomite
Quartz
Pyrite
Sphalerite
Galena
Chalcopyrite
Cerusslte
Cerargyrite
Embolite
Ankerite
Argenclte
Barite
Rhodochrosite
Pyrargyrite
Chalcocite
Covellite
Angles ite
Unlfanl te
LEAD ZINC ORE AT
Chemical
Compos it ion
FeS2
ZnS;
PbS
FeuS12
CuFeS2
Fe304
Fe203
Cu5FeS4
Zn2S104
S102
CaMgSi203
Ca2Mg5Si8022-
(Ott,F)2
3Mg02(Si02)-
2H20
Mg3Si40io(OH)2
B2S04
CaS04
Na5Al3F14
H20'CaO-4FeO-
Fe.0--4Si.02
Ca3Al2isi04)3
CMg^e^SisO^-
(OH,F)2
Ca2Al3Si13Oi2(OH)
PbC03
NaAlSi308
CaMg(C03)2
. PbS
ZnS
CuFeS2
NoFeS2
FeS2
FeS2
CusFeS4
«lS.
CaMg(C03)2
Cu2S
A1203-2S102'2H20
S102
CuS
Cu2S
CuC03'Cu(OH)2
CaMg(C03)2
Si02
FeS2
ZnS
PbS
CuFeS
PbC03
MINES (1031)
Specific
Gravity
5.02
3.9-4.1
7.57-7.59
4.58-4.63
4.1-4.3
5.26
4.9-5.3
4.9-5.4
3.9-4.1
2.65
3.22-3.38
3rt
• U
2.5-2.65
2.58-2.83
4Crt
• Jv
2.98
3.00
4.0
3.53
2.85-3.57
3.2-3.4
6.5
2-63
2.86
7.57-7.59
2.5-2.65
4.1-4.3
4.62
5.02
4.85-4.9
4.9-5.4
5.3-5.6
2.86
2.6-2.71
2.62 '
2.65
4.6
5.5-5.8
4
2.86
2.65
5.02
2.5-2.65
7.57-7.59
4.1-4.3
6.53-6.57
AgCl 5.55
Ag(Cl-Br) 5.8
Ca(Fe-Mn-Mg)(C03)2 2.8-3.1
Ag2S 7.2-7.4
BaS04
MnC03
Ag3-SbS3
Cu2S
CuS
PbS04
PbMoOi
*.a
3.7
5.85
5.5-5.8
4.6
6.37-6.39
6.5-7.0
Solubility Solubility
Reagents Water
a. HN03 si. 3.
3. HC1 i.
dec. HN03
dec. a. si. dec.
dec. HNOj si. s.
s . a . i.
a. HC1 i.
a . a . si . s .
a a {
a 4 a . L .
s. only HF 1.
i. i.
i. i.
dec. a. i.
i. i .
1. i.
a. HN03 v. s.
si. s. i.
gel. HC1 i.
i. i.
1. i.
i. i.
3. HN03 i.
1. 1.
a.a. sl. a.
dec. HN03
dec. a. 1.
dec. HN03 sl. s.
Sa i
. 3 . L .
3. HN03 sl. 3.
9. con. HNO 3 sl. s.
s. a. sl. s.
a. HN03 1.
s. a. sl, s.
s . HC1
1. i.
s. only HF 1.
3. HN03 1.
s. HN03 i.
t
s . a . i- .
3. a. sl. s.
s. only HF I.
s. HN03 sl. s.
dec. a. i.
dec. HN03
dec. HN03 sl. s.
a. a. ' i.
s,. MH4OH i.
s. NH40H i.
s. a. sl. s.
s. HN03 I.
1{
. i. •
. s.h. HC1 1.
dec. HN03 1.
s. HN03 1.
s HNO i i •
s. a. sl. s.
s. a. sl. 9.
i .
liartltiudS
6-b.5
3.5-4
2.5-2.75
3.5-4.5
3.5-4.0
5-6
5.5-6.5
5-5
7
5.5-6.5
5-6
2.5-3.5
^
3-3.5
3.5
3.5-4
5.5-6
6.5-7
c e t
J . J-o
6c
. J
3-3.5
6ft ^
~O . J
3.5-4
2.5-2.75
2.5-3.5
3.5-4.0
5.5-6
6-6.5
6-6.5
3
3-3.5
•1 C A
J . J~*+
3"
2. 5
7'
1.5-2
2.5-3
3.5-4
3.5-4
7
6-6.5
2.5-3.5
2.5-2.75
3.5-4
3-3.5
i) t:
£, • J
1.5
3.5-4
2-2.5
3-3.5
3.5-4
9 S
£, . J
2.5-3
1.5-3
2.5-3
2.75-3
338
-------
TABLE C-2 (Concluded)
State
Colorado
Utah
Idaho
Mineral
Pyromorphi.ce
Smlthsonlte
Calami ne
Chalcanthlte
Galena
Sphalerite
Argentlte
Tenoantlte
Proustlte
Polybasite
Barite
Rhodochroalte
Calcite
Quartz
Pyrlte
Enarglte
Tetrahedrite
Altalte
Blsnuthinlte
Bornlte
Chalcopyrite
Chalcoclte
Jamesonite
Tetradymite
Cerussite
Angles ice
Pyrolusite
Brochantlte
Coplayite
Halloyaite
Epsomlte
Lanarklte
Malachite
Melanterite
Minium
Smlthsonite
Galena
Sphalerite
Tetrahedrite
Chalcopyrite
Pyrrhotlte
Pyrlte
Arsenopyrlte
Magnetite
Hematite
Stlbnite
Uranlnlte
Gersdorffice
Boulangerlte
Polybaaite
Pyrargyrite
Bournonlte
Scheelite
Barite
Quartz
Siderite
Dolomite
Ankerite
Calcite
Chlorite
Hornblende
Biotite
Garnet
Chemical
Composition
3Pb3-P208-PbCl2
ZnC03
2ZnO-Si02'H20
CuS045H20
PbS
ZnS
Ag2S
5Cu2S-2ZnS-2AsS3
Ag3AsS3
(AgCu)16Sb2Su
BaS04
MnC03
Cu2S
S102
FeS2
Cu3AsS4
(CujFe)12Sb4Si3
PbTe
Bi2S3
Cu5FeS4
CuFeS
Cu2S
Pb4FeSb6Si4
Bi2(S1Te)3
PbC03
PbS04
Mn02+2H2°
CuS04'3Cu(OH)2
(Fe1Mg)Fe4(S04)5-
(OH)220H20
A1203-2S102'H20
Ma SO. -7H20
Pb2o-so4
CuC03-Cu(OH)2
FeS04- 7H20
Pb304
ZnC03
PbS
ZnS
(CuF2)12Sb4Sl3
CuFeS
FellS12
FeS2
FeS2'FeAg2
Fe304
Fe203
Sb2S3
U02 + Th, Zr,
La, Y, Pb
NiS2tUAs2
Pb5Sb4Sn
9Ag2S-Sb2S3
3AgS-Sb2S3
Pb5CuSbS3
CaH04
B2S04
S102
FeC03
CaMg(C03)2
Ca(Fe,Mg,Mn)-
(C03)2
Cu2S
NajAl^F^
(Ca,K_,K,Al,Fe)-
S102
K(Mg,Fe)3AlSi30io-
(OH,F)2
Fe3Al2(S104)3
Specific
Gravity
6.5-7.1
4.42-4.44
3.4-3.5
2.2
7.57-7.59
2.5-2.65
7.2-7.4
4.37-4.49
5.57
6.0-6.2
4.5
3.7
2.6-2.71 .
2.65
5.02
4.4-4.5
4.6-5.1
8.15 •
6.75-6.81
4.9-5.4
4.1-4.3
5.5-5.8
5.63 .
7.2-7.6
6.53-6.57
6.37-6.39
4.73-4.86
3.8-3.9
2.08-2.17
2-2.2
1.75
6.4-6.8
4
1.9
8.9-9.2
4.42-4.44
7.57-7.59
2.5-2.65
4.6-5.1
4.1-4.3
4.58-4.65
5.02
5.9-6.2
5.26
4.61-4.65
9.7
5.6-6.2
6.0-6.2
6-6.2
5.8
5.80-5.86
6.0
4.5
2.65
3.85
2.86
2.8-3.1
2.5-2.71
3.0
3.02-3.45
2.7-3.1
4.25
Solubility
Reagents
s. HN03
s. a.
gel. *
s. a.
dec. HN03
dec. a.
s. mo3
dec. HN03
-
dec. HN03
i.
s. h. HCl
s. HCl
a. only HF
s. HN03
s. aq. reg. andHNO
dec. HN03
s. a.
s. h. HN03
s. a.
dec. HN03
s. HN03
s. a.
s. HCl
s. a.
s. a.
s. HCl
s. HN03
s. a.
i.
s. a.
si. s. dll. a.
s. a.
s. a.
s. a.
s. a.
dec. HN03
dec. a.
dec. HN03
dec. HNOj
dec. a.
s. HN03
dec. HN03
s. a.
s. a.
s. a.
dec. h. HNO,
j
a. a.
dec. HN03
dec. HN03
3. a.
'dec. HCl
i.
s. only HF
s. h. HCl
s. a.
a. a.
-
s. HCl
si. s.
i.
dec. H2S04
i.
Solubility
Water
i
si. s.
I.
s.
-
i.
i.
i.
1.
i.
i.
1.
-
i.
Sl. S.
3 1>
i.
i.
i.
si. s.
Sl. 3.
i.
I.
1.
i.
si. s.
i.
i.
s.
i.
s .
1.
I.
s .
Sl. S.
si. a.
.
i.
1. .
Sl. 3.
si. dec.
si. s.
si. s.
.1.
i.
si. s.
i.
si. a.
1.
i.
Sl. 3.
1.
i.
i.
i.
Sl. S.
Sl. S.
-
i.
i.
i.
i.
Hardness
3.5-4
4-4.5
4.5-5
J.2-2.5
2.5-2.75
2.5-3.5
2-2.5
3-4
2-2.5
2-3
3-3.5
3.5-4
3
7
6-6.5
3
3-4.5
3
2
3
3.5-4
2.5-3
2.5
1.5-2
3-3.5
2.5-3
2-2.5
3.5-4
2.5-3
1-2
2-2.5
2-2.5
3.5-4
2
2.5
4-4.5
2.5-2.75
2.5-3.5
3-4.5
3.5-4
3.5-4.5
6-6.5
5.5-6
5-6
2
5-6
5.5
2.5-3
2-3
2.5
2.5-3
4.5-5
3-3.5
7
3.5-4
3.5-4
3.5-4
3
3.5-4
5-6
2.5-3
6.5-7
339
-------
The important types of lead deposits are the following:—
1. Deposits formed in sedimentary rocks without apparent genetic
relationship with igneous rock. They occur as tabular replacements of
receptive strata, usually in limestone reefs and dolomites. The ores of
this type usually contain galena, sphalerite, and pyrite, but seldom gold,
silver, or antimony to any appreciable extent.
2. Deposits formed at shallow or intermediate depths genetically
associated with igneous rocks, characterized by complex ores and comprising
(a) vein deposits, (b) disseminated pyritic deposits of igneous rocks, and
(c) silver-lead replacement in limestone.
3. Deposits in beds formed at high temperature and pressure in,
or genetically associated with, igneous rocks. The ore minerals are zinc
blende, galena, pyrite or pyrrhotite, quartz, calcite, garnet, and rhodonite.
4. Contact-metamorphic deposits containing characteristic minerals
found near contacts of limestone with igneous rocks. The ore minerals are
galena and its oxidation products in a gangue of calcite, rhodonite, garnet,
pyroxene, hornblende, and tremolite.
The average grade of lead ores mined contain between 3.0 and 8.0
percent lead.
Zinc ores are aggregates of minerals. The most common zinc mineral
is sphalerite or zinc blende (ZnS), which with its oxidation products
smithsonite (ZnCC^) and hemimorphite (Zn^Si207(OH)2-H20), forms the chief
zinc minerals of the world. Zincite (ZnO), willemite (Z^SiO^.), and
franklinite ((Zn,Mn)0-Fe203) occur in a unique and very important zinc
deposit at Ogdensburg, New Jersey .i/
v_,
Sphalerite almost always occurs in association with galena, the
sulfide of lead. It may also be associated with copper or other base-metal
sulfides or occur alone. Next to iron, cadmium is the most common
substitutional impurity in the sphalerite lattice; typical cadmium content
of zinc concentrates runs about 0.3 percent.
Most zinc ores occur in limestones and dolomites but also occur
in sufficient amounts at many mines, including those of Butte, Montana;
the Coeur d'Alene, Idaho; and Ogdensburg, New Jersey.
340
-------
Missouri
The southeast Missouri lead district located about 113 kilometers
south of St. Louis, embraces four important subdistricts and several minor
ones. The important subdistricts are Mine La Motte, the Old Lead Belt,
Indian Creek, and the Viburnum Lead Belt.
In terms of past production, it is one of the world's largest
districts, having produced nearly 9,979,000 metric tons of lead. The vast
majority of this production was from the Old'Lead Belt centered around
Bonne Terre and Flat River. However, the new mines in the Viburnum Lead
Belt promise to be at least as productive as the older area.
The lead-producing area of Missouri is located on the flank of
a somewhat circular-shaped positive structure known as the St. Francis
Mountains, the northeastern part of the Ozark dome. The core of this
structure is formed by Precambrian igneous rocks, mainly rhyolitic
porphyries intruded by granites and diabasic rocks; they are the oldest
rocks in the mining region. Many of the erosion-formed igneous knobs
are exposed at surface, and others remain buried by Cambrian and Ordovician
sedimentary formations.—'
The primary sulfide minerals include galena, sphalerite,
chalcopyrite, siegenite, bravoite, pyrite, and marcasite.
Galena occurs as bedded or sheet deposits along disconformities
and bedded planes, in parts as replacements but locally as open space
fillings.; as disseminated crystals and crystalline aggregates in several
types of dolomite and black shales; and as fractured fillings.
Sphalerite, although a minor constituent for the district as a
whole, may be abundant locally, and .in these areas commonly is associated
with gray shaly dolomite and black shale.
Chalcopyrite is abundant in the eastern part of the district and
is rare in the western part. The copper mineral occurs as disseminations
in nearly all Bonne Terre lithologies and as thin seams along bedding
planes. Where it occurs, chalcopyrite usually is associated with galena.
Siegenite and bravoite are present
-------
Leadville District, Colorado
The Leadville District is on the west flank of the Mosquito Range
in central Colorado. The ore deposits are in a sequence of dolomites and
quartzites, Cambrian through Mississippian in age and about 152 meters
thick, which is extensively intruded by porphyry sills, dikes, and plugs
of Tertiary age. The sedimentary rocks and sills dip about 15 degrees east
and are broken by many faults, which are predominantly of a northwest trend.—
The.ore deposits are principally manto replacement deposits
confined to three dolomite units in the stratigraphic sequence. Of these,
the Leadville Dolomite is the most productive. Many replacement bodies
have vein roots or branch from veins. The veins occupy faults and are
productive mainly in the section of quartzites and dolomites and included
sills.
The primary ores consist primarily of pyrite, masmatitic
sphalerite, and galena, but locally contain chalcopyrite, silver minerals,
bismuth minerals, and gold. Principal gangue minerals are manganosiderite
and jasperoid.
Coeur d'Alene District. Idaho
The Coeur d'Alene district in the panhandle of Idaho' is one of
the major- lead-zinc-silver producing areas in the world.
Ore occurs in a series of steeply dipping replacement veins of
relatively simple mineralogy. Six periods of mineralization, ranging from
Precambrian to Tertiacy in age, are recognized; The productive galena'
sphalerite-bearing and tetrahedrite-bearing veins of the main period of
mineralization are younger than.the Late Cretaceous stocks but are also
probably Cretaceous in age.-'
The primary sulfide minerals include galena, sphalerite,
tetrahedrite, chalcopyrite, and-pyrite. Principal gangue minerals are
pyrrhotite, magnetite, and arsenopyrite.
342.
-------
The major producing mines are clustered in two groups. One group
lies north of the Osburn fault in a large triangular-shaped arc that extends
north and east of the town of Wallace and is called the Mullan-Burke-
Ninemile area. This group includes such famous old mines as the Hecla,
Hercules, Stanford-Mammoth, Tamarack, Tiger-Poorman, Frisco, Star, and
Morning. The other group lies on the south side of the Osburn fault, and
extends from near Wallace westward beyond Kellogg nearly to Pine Creek.
This group, somewhat more elongate than the MulIan-Burke-Ninemile area,
includes such famous mines as the Galena, Sunshine, and Bunker Hill.
Numerous other mines and prospects occur in this group, which also include
the productive string of mines along the East Fork of Pine Creek. The
eastern third of this group is called the Silver Belt of the Coeur d'Alene
district because of the silver content of the ore.
Tintic Mining District, Utah
The East Tintic. district in central Utah produces gold, silver,
copper, lead, and zinc. All of this ore has been produced from blind ore
bodies in Paleozoic sedimentary rocks that are concealed beneath
hydrothermally altered volcanic rocks and cut by intrusive bodies of
Eocene age. The Paleozoic rocks, which range in age from Early Cambrian
to Mississippian, form the core and limbs of a large, north-trending
asymmetric anticline that is cut by low-angle thrust faults, high-angle
transcurrent faults, mineralized fissures and faults.—
The ore bodies generally may be grouped into two classes: (1)
massive replacement bodies that are rich in silver, lead, zinc, and
manganese; and (2) fissure veins that are valuable primarily for their
content of gold, copper, and silver.
The-primary ore minerals of the.replacement deposits are
argentiferous galena and sphalerite with some of the silver probably
present in blebs of argentite and tennantite in the galena and in small
amounts of other silver minerals such as proustite, pearceite, and
polybasite. Barite, rhodochrosite, manganosiderite, calcite, and quartz
are the gangue minerals. Native gold, enargite, and tetrahetrite are the
principal ore minerals of the auriferous copper veins in Tintic Quartzite;
and various gold and silver tellurides and tetrahedrite are the chief ore
minerals in the gold-telluride veins in Tintic Quartzite. Quartz, pyrite,
barite, and clay minerals are the main gangue minerals of the vein deposits,
343
-------
Balmat-Edwards District, New York
The zinc deposits of the Balmat-Edwards Division of the St. Joseph
Lead Company in northern New York provide some 10 percent of the domestic
zinc produced annually within the United States. These complex ore bodies
are contained within marbles of the Frecambrian Granville series, in a
repetitive sequence of dolomites and silicated units.
The assemblage of ore minerals is simple. Pyrite and sphalerite
are abundant, and galena, pyrrhotite, and chalcopyrite are in minor amounts.
Gangue minerals include quartz, diopside, tremolite, serpentine,
talc, carbonate, barite, and anhydrite. In the supergene ores, chlorite is
abundant, with minor ilvaite and garnet. In most, if not all cases, the
nonmetallic minerals found within ores are either reworked fragments of
adjacent wall rock or late products of retrograde metamorphism in the
surrounding calcsilicate marbles.—
There is a wide range of mineral association and texture.
Macroscopically, most of the ores are massive to disseminated coarsely
crystalline aggregates of sphalerite and pyrite in gangue of quartz,
carbonate, and diopside. Where inclusions of wall rock are abundant, the
sulfide bodies have the aspect of ore breccias. Pyrite occurs with
sphalerite either as cubes and pyritohedra in discrete subhedral or anhedral
grains up to 10 centimeters, or in massive aggregates almost without
sphalerite or gangue. Sphalerite usually is coarse-grained but ranges from
0.01-millimeter grains to subhedral crystals over 8 centimeters in length.—'
Galena, chalcopyrite, and pyrrhotite are most conspicuous as
gash veins and disseminated blebs, often with minor sphalerite, in diopside
rock at some distance from zinc ore bodies. Occurrences of chalcopyrite and
pyrrhotite within ore bodies can be observe'd only in polished sections.
Chalcopyrite appears as fine blebs and chains within sphalerite and
pyrrhotite, and pyrrhotite has complex relations with pyrite and sphalerite.
Gangue minerals may be ne'arly absent in some of the high-grade
ores, or may predominate in disseminated or Weakly mineralized zones.
344
-------
Lehigh County. Pennsylvania
The Friedensville zinc mine of the New Jersey Zinc Company is
located about four miles south of Bethlehem, Pennsylvania, in the Saucon
Valley, an infolded and down-faulted block of Cambro-Ordovician carbonate
sediments surrounded by hills of the Precambriah gneiss and Cambrian
quartzite except where breached by Saucon Creek at its entrance into the
Lehigh River.
The minerals identified to date are dolomite, calcite, chert,
rare microcline feldspar, sphalerite, pyrite, rare chalcopyrite, calamine,
smithsonite, rare greenockite, sauconite, and carbonaceous materials and
sericite in streaks. The dolomite is present as fragments in the breccia,
medium to gray in color, and as white veinlets. Calcite occurs as white
veinlets as does the quartz. Chert occurs as ellipsoids, generally
flattened.
Sphalerite is generally dark gray in color and on a wet face
in the mine is difficult to distinguish from the dark gray host. While
in general the grain size is medium to coarse, locally it is flint-like
and is characterized by conchoidal fracture.
Pyrite is widely distributed in the ore as masses and in
subhedral to enhedral crystals; it commonly occurs banded with the
sphalerite. Table C-3 shows the distribution of the mineralogy in the
major zinc districts.
Mascott-Jefferson City Zinc District. Tennessee
The Mascott-Jefferson City district currently is the largest
zinc-producing area in the United States. The producing mines, Mascott,
Davis, Young, New Market Zinc, Coy, and the Jefferson City, are 24 to 45
kilometers northeast of Knoxville in the Valley and Ridge physiographic
province, an area between the Great Smoky Mountains and the Cumberland
escarpment.
The ore is strata bound, occurring essentially within a
stratigraphic range of 61 meters in the Lower Kingsport and Upper Longview
formations of Lower Ordovician age. In the lower strata of the mineralized
section, the ore generally is found in coarsely crystalline clastic dolomite
breccia containing silica and chert debris within a sequence of aphanitic
limestones. The upper strata of the mineralized section are fine-grained
primary dolomites that, in the ore zones, have been fragmented in mosaic
patterns by solution collapse with the interfragmental space filled by
white gangue dolomite and sphalerite.
345
-------
TABLE C-3
SUMMARY OF MINERALOGICAL SYSTEMS
Tennessee
New Jersey
Mineral
Pyrite
Chalcopyrite
Galamine
Smithsonite
Greenockite
Dolomite
Calcite
Sphalerite
Dolomite
Chalcopyrite (rare)
Calcite
Sphalerite
Pyrite
Chalcopyrite
Calamine
Smithsonite
Greenockite
Dolomite
Calcite
Sphalerite
ZINC ORE AT MINES (1031)
Chemical
Composition
FeS2
CuFeS2
2ZnO'Si02-H20
Zn C03
CdS
C2Mg(C03)2
Cu2S
ZnS
CaMg(C03)2'CO
CuFeS2
Cu2S
ZnS
FeS2
CuFeS2
2ZnO-Si02-H20
Zn C03
CdS
C2Mg(C03)2
Cu2S
ZnS
Specific
Gravity
5.02
4.1-4.3
3.5
4.42-4.44
4-9
2.86
2.6-2.71
2.5-2.65
2.86
4.1-4.3
2.6-2.71
2.5-2.65
5.02
4.1-4.3
3.5
4.42-4.44
4-9
2.86
2.6-2.71
2.5-2.65
Solubility
Reagents
s. HN03
dec. HN03
gel. a.
s. a.
s . a.
s. a.
a. HCl
dec. a.
s. a.
dec. HN03
s. HCl .
dec. a.
s. HN03
dec. HN03
gel. a.
s. a.
s. a.
s. a.
s. HCl
dec. a.
Solubility
Water
si. s.
si. s.
i.
si. s.
i.
si. s.
-
i.
si. s.
si. s.
-
i.
si. s.
si. s.
i.
si. s.
i.
si. s.
-
i.
Hardness
6-6.5
3.5-4
4.5-5
4-4.5
3-3.5
3.5-4
3
2.5-3.5
3.5-4
3.5-4
3
2.5-3.5
6-6.5
3.5-4
4.5-5
4-4.5
3-3.5
3.5-4
3
2.5-3.5
-------
The mineralogy is unusually simple, sphalerite being the only
important sulfide. Pyrite is very scarce. Dolomite gangue is abundant,
silica was deposited in quite small amounts, calcite is minor, and fluorite
and arite are very rare.
The principal ore-controlling structures are rubble breccia zones
that were formed during the post-Lower Ordovician-pre-Middle Ordovician
erosion interval by supergene solution as a phase of the development of a
regional karst system of underground drainage that extended to depths of at
least 244 meters. The major Appalachian orogenic structures are post-ore
and have been superimposed on the ore bodies.—'
Uranium and Vanadium Ores
Uranium metal never occurs in its pure form in nature, but is
always present in combinations with other elements, forming a uranium-
bearing mineral. There are over 100 uranium minerals. Some occur as
primary minerals and many others as secondary minerals. Primary minerals
are those that have not been altered since their formation by igneous
action. Secondary minerals are those that are formed as a result of
decomposition of the original primary minerals.
The four known primary uranium ore minerals are:—
Uraninite - Crystalline uranium oxide.
Pitchblende - Amorphous uranium oxide.
Davidite - Rare earth-iron-titanium-uranium-oxide.
. Coffinite - Adamantine uranium oxide.
The six known secondary uranium ore minerals are: —
Carnotite - Potassium-uranium vanadate.
Tyuyamunite - Calcium-uranium vanadate.
Torbernite and meta-torbernite - Hydrous copper-uranium
phosphates.
Autunite and meta-autunite - Hydrous calcium-uranium
phosphates.
347
-------
Uranophane - Hydrous calcium-uranium silicate.
Schroeckingerite - Complex hydrated sulfate, carbonate
and fluoride of calcium, sodium, and uranium.
Vanadium occurs in igneous rocks as well as in sedimentary
formations. Alteration of the original mineral is very common, and
alteration products are numerous. Patronite (vanadium sulfide), for
example, oxidizes to the sulfate; and, in contact with calcium minerals,
it is transformed to various hydrated calcium vanadates.
The chief ore minerals of vanadium are patronite, VS.; carnotite,
K20-2U03'V2C>5.3H20; roscoelite (vanadium mica); and vanadinite, Pb(PbCl) (V20,),.
Table C-4 shows the mineralogy of the major uranium-vanadium districts.
Grants District, New Mexico
Uranium of the Grants region occurs predominantly in the
continental sandstones of the upper part of the Jurassic Morrison Formation,
but significant lesser deposits are found in limestone of the Jurassic
Todilts Formation and in black shale of the Cretaceous Dakota Formation.—'
The deposits are disseminations that form runs ranging from a few hundred
metric tons to several million metric tons. The ore consists mainly of
uraninite, uraniferous carbonaceous material, coffinite, and such secondary
oxidized minerals as tyuyamunite, carnotite, and uranophane. In sandstone,
the ore runs were localized by mudstone, interstitial carbonaceous material,
and primary sand channel trends. In limestone, ore deposition was related
to folds and fractured zones.
The great bulk of production has come from the Ambrosia Lake
mines in the Grants district and the Jackpile and Paguate mines in the
Laguna district. The ores, which generally have been 0.20 to 0.30 percent
UNjOg are treated at mills operated by Kerr-McGee Corporation, United
Nuclear-Homestake Partners, and the Anaconda Company, all in the Grants
district.
348
-------
TABLE C-4
SUMMARY OF MINERALOGICAL SYSTEMS
State
New Mexico
Wyoming
Mineral
Coffinite
Uraninite
Tyuyamunite
Carnotite
Zippetite
Andersonite
Bayleyite
Pyrite
Barite
Calcite
Metatorbernite
Phosphuranylite
Schoepite
Uraninite
Pyrite
Hematite
Calcite
Coffinite
Molybdenum
URANIUM-VANADIUM ORE AT MINES (SIC-1094)
Chemical
Composition
U(Si04)(OH)4
U02
Ca(U02)2(V04)2 '7-10-5H20
K2(U02)2(V04)2-H20
2U03*S03-5H20
Na2Ca(U02)(C03)3-6H2)
Mg(U02)(C02)2'18H2O
FeS2
CaC03
Cu(U02)2(P04)2'4H2C>4
Ca(U02)4(P04)2-(OH4)-7H20
4U03'9H20
U02
FeS2
Fe203
CaC03
U(Si04)(OH)4
MoS2
Specific
Gravity
5.1
10.96
3.7-4.3
3.8-4.2
3.5
2.8-2.86
2.05-2.06
5.02
2.6-2.71
3.22
3.0
4.8
10.96
5.02
4.9-5.3
2.6-2.71
5.1
4.7
Solubility
Reagents
s. a.
s.HN03
cone. H2SOg
s. a.
s. a.
-
s. a.
s. a.
s. a.
s. HCl
s. HN03
s. a.
s. a.
cone . H2SC>4
s.HN03
s. a.
s. HCl
s. HCl
s. a.
dec. HN03
Solubility
Water
i.
i.
i.
i.
-
s.
s.
si. s.
i.
i.
i.
si. s.
i.
si. s.
i.
i.
i.
i.
Hardness
5-7
5-6
Soft
1-2
2-3
6-6.5
3
2-2.5
2
2-3
5-6
6-6.5
• -^
6
3
5-7
1-1.5
-------
TABLE C-4 (Concluded)
CO
tn
O
Utah
Mineral
Uraninite
Coffinite
Carnotite
Tyuyamunite
Rauvite
Pyrite
Uranophane
Vanadinite
Pyroxene
Uraninite
Coffinite -
Quartz
Barite
Galena
Pyrite
Yttrium (rare)
Carnotite
Tyuyamunite
Chemical
Composition
U02
U(Si04)(OH)4
K2(U02)2(V04)2-H20
Ca(U02)2(V04)2«7-10-5H20
CaO-2U03-6V205-20H20
FeS2
Ca(U02)2(Si03)2-(OH2)5H20
Pb5(V04)3Cl
CaMg(Si03)2
U02
-
U(Si04-)(OH)4
Si02
BaS04
PbS
FeS2
Y
K2(U02)2(V04)2-H20
Ca(U02)2(V04)2-7-10-5H 0
Specific
Gravity
10.96
5.1
3.8-4.2
3.7-4.3
3.4-3.6
5.02
3.81-3.90
6.6-7.1
3.2-3.38
10.96
5.1
2.7
4.3-4.6
7.4-7.7
5.0
4.34
3.8-4.2
3.7-4.3
Solubility
Reagents
cone. H2S04
s. HN03
s. a.
s. a.
s. a.
-
s. a.
gel. HC1
v.s. HCl
i.
cone. H2S04
s. HN03
s. a.
s. only HF
i.
dec. HN03
i.
v.s. dil. a.
s. a.
s. a.
Solubility
Water
i.
.
-------
Colorado
The Colorado Plateau region has been the principal domestic
source of uranium and vanadium. Most of the deposits occur in streamlaid
lenses of sandstone in the Chinle and Morrison Formations. The ore minerals
below the zone of oxidation are low-valent oxides and silicates of uranium
and vanadium and uraninite, coffinite, montroseite, and vanadium-bearing
mica, chlorite, and clay. All of these, except the quite stable vanadium
silicates, oxidize readily to a variety of high-valent uranium and vanadium
minerals. The common copper sulfides are widespread in sparse amounts and
abundant enough in a few places to be ore minerals. Accessory minerals
are mainly sulfides; pyrite and marcasite are common but not abundant, and
trace amounts of galena and sphalerite are widespread. Minerals containing
molybdenum, selenium, chromium, nickel,cobalt, and silver are also present,
but only in a few places are these minerals abundant enough to be recognized.
Ores containing uranium and vanadium minerals have been mined from
the Salt-Wash Member of the Morrison Formation from many localities in the
Colorado Plateau region since about 1900. The most productive deposits are .
in a relatively small area in southwestern Colorado referred to as the
Uravan mineral belt. Principal metals recovered from the Uravan mineral
belt ores are uranium and vanadium. Radium.was recovered early in the
1900's but no substantial amounts have been recovered since about 1923.
Trace amounts of other minerals have been detected in the ores, but none
of these occurs in quantities sufficient to warrant recovery. Metals
detected other than those found in common rock-forming minerals include
molybdenum, copper, silver, selenium, chromium, nickel, cobalt, rare
earths, and manganese.
Principal uranium minerals in the unoxidized ores are uraninite
and coffinite, while the main uranium mineral in the oxidized ores is
carnotite with minor amounts of tyuyamunite. Vanadium-bearing clays,
consisting of chlorite and hydromica, are the main vanadium minerals in
both the oxidized and unoxidized ores. Montroseite, a low-valeht vanadium
oxide occurs in significant amounts in the unoxidized ores. The vanadium
in the vanadiferous clays is firmly mixed and is relatively unaffected
during oxidation; however, montroseite oxidizes readily first to an
intermediate mineral, corvusite. Upon further oxidation, the vanadium
forms a series of vanadates consisting of carnotite, hewettite,
metahewettite, pascoite, rauvite, fervanite, and hummerite.
The uranium and vanadium minerals-almost always occur intimately
mixed together regardless of the environment containing the mineralization
or type of mineral association. Occurrences of one metal without the
other in large amounts are virtually unknown in the Uravan mineral belt
area. Generally, the amount of vanadium exceeds the uranium in ratios
ranging from 3:1 to 10:1 for large quantities.
351
-------
The intensity of uranium-vanadium mineralization varies
considerably throughout a typical deposit, ranging from weakly mineralized
rock to ore containing several percent U^Og and V20c. Average shipping
grades of typical mines are 0.2 to 0.3 percent U-jOg and 1 to 2 percent
V205.
Shirley Basin District. Wyoming
The Wind River Formation of Eocene age is the host rock for
large high-grade uranium deposits in the Shirley Basin. The major deposits
are in a northwest-trending belt of sandstones that were deposited in
stream channels and that were lithologically favorable for uranium
accumulation.
The^major deposits within the belt.are at, or near, the margins
of large tongues of altered sandstone formed in the most massive parts of
thick sandstone beds. At least two tongues of altered sandstone are
present in the belt, one of which is 8 kilometers long, 5 kilometers wide,
and 21 meters thick; the other tongue is somewhat smaller.
Ore bodies consist of a few hundred metric tons to several
hundred thousand tons of material containing.from 0.10 to about 2.00
percent UoOg. The ore mineral is uraninite associated with pyrite,
marcasite, hematite and calcite.
Mercury Ore
Mercury ore is found- in rocks of all geologic ages and all
classes. The common host rocks are limestone, calcareous shales,
sandstone, serpentine, chert, andestie, basalt, and rhyolite. Two
general types of cinnabar ores can be distinguished: (1) disseminated
ore, in whichithe cinnabar has impregnated a more or less fine-grained
or highly brecciated gangue; and (2) ore deposited in fissures and
cracks in the country rock. Mercury is recovered almost entirely from
the sulfide mineral cinnabar (HgS--86.2 percent mercury and 13.8 percent
sulfur). The distribution of minerals and the mineralogy of the
California mercury district is shown in Table C-5.
352
-------
TABLE C-5
SUMMARY OF MINERALOGICAL SYSTEMS
State
California
i-n
Mineral
Stibnite
Cinnabar
Pyrite
Marcasite
Dolomite (sparse)
Calcite (sparse)
Alunite
Kaolinite
Halloysite
Montmorillonite
Zeolite
Serpentine
Quartz
Aragonite
MERCURY ORE
Chemical
Composition
HgS
FeS2
CaMg(C03)2
K2A16(OH)12(S04)4
2(Al203) -Si02'H20
AL203'2Si02«H20
(MgC2) 0- A1203 • 5SiOj
Na20. A1203 . 3S i02 • 21
Mg3Si205(OH)4
Si02
AT MINES (SIC -1092)
Specific
Gravity
4.61-4.65
8.1
5.02
4.85-4.9
2.86
2.7-2.9
2.6-2.8
2.59
2.0-2.2
>-2H20 2.0-3.0
120 2.2-2.35
2.55
2~.65
2.93
Solubility
Reagents
s. a.
i.
s. a.
s. c. HN03
s. h. HCl
s. HCl
s. H2S04
sl. dec. HCl
i.
i.
gel. a.
s. a.
s. only HF
s. HCl
Solubility
Water
i.
i.
sl. s.
Sl. 8.
sl. s.
S.
sl. s.
i.
i.
i.
i.
i.
sl. s.
Hardness
2
2-2.5
6-6.5
6-6.5
3.5-4
3
3.5-4
2.5
1-2
1-2
5-5.5
2.5-3.5
7
3.5-4
-------
Sulphur Bank is unique among mercury deposits of the western
United States. A major ore body was formed in the recent past by
hydrothermal processes that are represented today by actively flowing hot
springs. Metacinnabar and stibnite have deposited recently, and it seems
probable that some deposition of these minerals still is going on.
The major sulfide minerals are pyrite, marcasite, and cinnabar.
Metacinnabar occurs in local concentrations. Stibnite is present in minor
amounts.
Miscellaneous Ores
Pure titanium does not exist in natural form because of its
great radioactivity with other elements. Minerals rich in titanium are
few, and only ilmenite (FeliC^) and rutile (TiC^) have commercial
importance. Theoretically, pure ilmenite contains 52.67 percent TiC^,
and ilmenite from rock deposits and some sand deposits commonly contains
42 to 50 percent Ti02« Some sand deposits, however, yield altered
ilmenite containing 60 percent or more Ti02- In rock formations, chiefly
anorthosite and related rocks, ilmenite is characteristically associated
with iron ores, principally magnetite, but in some places with hematite.
In sand or sedimentary deposits ilmenite occurs with rutile,
monazite, zircon, garnet, staurolite, and other minerals of high specific
gravity--the so-called "heavy metals." Titanium minerals found in beach
sands come from weathered igneous rocks. Fragments of the rocks are
carried to secondary deposits (beach, dune, and placer) by erosion or
rainfall where wave and wind action aid in concentrating the heavy
minerals.
Zirconium is associated with other metals in many minerals, but
is recovered only from the minerals zircon and baddeleyite.
!
Zircon, a zirconium orthosilicate (Zr02'Si02), grades from
colorless to pale yellow, to brownish .yellow, to grayish green or reddish
brown. It has a hardness of 7.5, a specific gravity of 4.2 to 4.86, a
conchoidal fracture, and an adamantive luster.
Baddeleyite, the zirconium dioxide, grades from colorless, to
yellow, to brown, to black. It has a hardness of 6.5, a specific gravity
of 5.5 to 6.0, a conchoidal to uneven fracture, and a greasy to vitreous
luster.
354
-------
Beryllium
Beryllium is a constituent of over 30 known minerals. The only
present commercial source of beryllium, however, is the mineral beryl, a
beryllium-aluminum silicate that theoretically can contain 14 percent
beryllium oxide. Beryl occurs principally in long, prismatic, hexagonal
crystals that usually have a vitreous luster and are emerald green or
pale green, grading into light blue, yellow and white. It has a hardness
of 7.5 to 8.0, a specific gravity of 2.63 to 2.80. Table C-6 contains
the mineralogy for the major deposit in Utah.
0
Beryl occurs in pegmatites and granites. In pegmatities beryl,
if present, is usually disseminated in small grains; occasionally, however,
large crystals 0.6 to 0.9 meters in diameter and many meters long are
found. Almost all domestic beryl is recovered as a coproduct from the
following minerals: feldspar, mica, columbite, tantalite, lithium minerals
or cassiterite.
In late 1959, a high concentration of beryllium was discovered
in tuff (associated with small amounts of fluoride) in the Thomas Range,
Utah. The beryllium mineral was identified as bertrandite (2Be2SiO/-l^O)
and occurs on the flats surrounding Spor Mountain in the western part of
the Thomas Range in western Juab County, 74 kilometers northwest of Delta,
Utah. On the east side of Spor Mountain, these deposits are in quartzanidine
crystal tuff of the older volcanic group; on the west side, they are found in
the vitric tuff, which commonly contains dolomite pebbles, of the younger
volcanic group. The tuff is believed to be favored as the host rock because
of its high porosity and permeability. In addition, small beryllium-bearing
veins have been found in the Paleozoic dolomite in a few places adjacent to
volcanic rocks.
Rare Earth
The rare earths are a closely related family of 15 metals
(lanthanum, cerium, praseodymium, neodymium, promethium, samarium,
europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium,
ytterbium, and lutetium). The principal commercial source of these metals
has been the mineral monazite, a rare earth phosphate containing thorium;
but the discovery in 1949 in California of a large deposit of bastnaesite,
a fluorocarbonate of the rare earths, added another source to the United
States supply. The rare earth silicate minerals, cerite and allanite, are
considered commercial sources.
355
-------
TABLE C-6
SUMMARY OF MINERALOGICAL SYSTEMS
-Mineral
Bertrandite
Quartz
Chlorite
Biotite
Hornblende
Augite
Hypersthene
Apatite
Zircon
Topaz
Tourmaline
Rutile
Cassiterite
Trioysiite
Montnierillonite
Calcite
Ankerite
Opal
Chalcedony
Orthoclase
Fluorite
Limonite
Dolomite
Black Manganese Oxide
i's Home lane
Pyrolusite
BERYLLIUM ORE AT MINES (SIC- 1099)
Chemical
Specific
Composition Gravity
Be4Si07(OH)2
Si02
12(Mg-Al-Fe)-
(SiAl)8-020(OH)16
K(MgFe)-3-AlSi3-
010(OHF)2
K'CaAl.Fe.Mg-
Na-Si02
CaMg(Si02)2Mg-
AlFeSiOfc
FeMgSi03
(CaF)Ca4(P04)3
ZrO2-Si02
[Al(FOH)]2Si04
AlBSi03-Li,Fe,Mg
Ti02
Sn02
Si02
Si02
(KgCa)-0-Al203-
5S102-2H20
CaC03
Ca(Fe,Mg,Mn)(C03)2
Si02-nH20
Si02
KAlSi308
CaF2
Fe203-H20
MgCA(C03)2
Mn02(B2,K2,Na )-
0,H20
MnO,'2HoO
2.6
2.65
2.6-3.3
2.7-3.3
3-3.5
3.2-3.5
3.4-3.5
3.17-3.23
4.7
4.7
2.9-3.2
4.26
7.0
2.27
2.28-2.33
2.0-3.0
2.7-2.9
2.8-3.1
1.9-2.3
2.6-2.64
2.57
3-3.3
3.6-4.0
2.86
3.3-4.7
4.73-4.86
Solubility
Reagents
i.
s. HF
i.
dec. H2S04
i.
i.
i.
a. a.
i.
i.
s. H2S04
• s. cone. H2S04
i.
s. h. NA2C03
i.
s. HC1
s. a.
s. HF
i.
i.
s. a.
s. HC1
s. h. HC1
s. HC1
s. HC1
Solubility
Water
i.
i.
i.
i.
i.
i.
i.
v. si. s.
i.
i.
i.
i.
i.
i.
i.
s.
i.
i.'
i.
i.
i.
i.
si. s.
i.
i.
Hardness
6
7
2-3
2.5-3
5-6
5.5-6
5-6
4.5-5
5-6
7.5
7-7.5
6-6.5
6-7
6-7
7
1-2
3
3.5-4
5.5-6.5
7
6
4
5-5.5
3.5-4
5-7
2-2.5
-------
A massive barite-carbonate-bastnaesite lode is located in the
Mountain Pass area of San Bernardino County, California. This deposit of
bastnaesite, a fluorocarbonate of the rare earth metals, was discovered in
April 1949. The main ore body is about 762 meters long and 152 meters
wide. A tremendous tonnage is available, with a grade of about 10 percent
total rare earth oxides.-'
Antimony
Antimony occurs in many ore minerals. The most common ores,
with their compositions, are shown in Table C-7.
TABLE C-7
MINERALS OF ANTIMONY^
Name of Mineral Formula Percent Antimony
Stibnite Sb2S3 71.4
Stibiconite H2Sb205 74.5
Senarmonite Sb203 - 83.3
Cervantite Sb203Sb205 78.9
Valentinite Sb2°3 83<3
Livingstone HgS - 2Sb2S-' ' 53.0
Jamesonite PbSbSc 29.4
Antimony deposits may be classified into two distinct general
types, but gradations exist within the two. 'The first is the simple
type, consisting primarily of antimony minerals in a siliceous or
carbonate gangue with only small or negligible quantities of other metals.
In almost all deposits, the original antimony mineral is stibnite, the
antimony sulfide, but native or metallic antimony is occasionally found.
Where the ores have been exposed to extensive oxidation, the original
stibnite has been entirely converted to antimony oxides.
357
-------
The second type of deposit is mineralogically complex and is
very commonly also structurally complex. In many places, stibnite is the
antimony-bearing mineral, but often the metal is only one component of
minerals that also contain lead, copper, silver, and mercury. In addition,
in most deposits of the complex type, other nonantimony-bearing metallic
minerals are present and are commonly more abundant than the antimony-
bearing minerals. With few exceptions, the ores are mined primarily for
lead, gold, silver, mercury, zinc, copper, or tungsten. The antimony may
be a by-product of only minor value; or, in some instances, the value of
the ore decreases because of the antimony content.
The Sunshine Mine in the panhandle of Idaho is the major antimony
producing mine in the United States. The mine lies on the south side of
the Osburn fault and extends from near Wallace, westward beyond Kellogg
to Pine Creek, and is adjacent to the Mullan-Burke-Ninemile area.
Ore occurs in a series of steeply -dipping replacement veins of
relatively simple mineralogy. Six periods of mineralization, ranging from
Precambrian to Tertiary in age, are recognized.
The major vein minerals at the Sunshine Mine are: siderite,
quartz., pyrite, tetrahedrite, chalcopyrite, galena, and sphalerite. The
valuable minerals are tetrahedrite (Cu, Fe, Zn, Ag)^2 Sb4 S^-j, which
contains silver, antimony and copper; chalcopyrite (GuFeS2), which contains
copper, and galena (PbS), which.contains lead. Tetrahedrite is the most
important ore mineral. Some of the tetrahedrite occurs as massive sulfide,
but most of it occurs as disseminated particles and veinlets in siderite
and quartz. The chalcopyrite and galena content are very erratic, varying
from 0.01 to 0.2 percent in-the mill heads.
358
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359
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APPENDIX D
REFERENCE BIBLIOGRAPHY
GLOSSARY
UNITS AND CONVERSION FACTORS
360
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379
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GLOSSARY
Adit - A horizontal passage driven from the surface for the working of
a mine.
Anode - Electropositive pole of an electrolytic cell.
Anolyte - The electrolyte adjacent to the anode in an electrolytic cell.
Bench - A ledge which in open-pit mines forms a single level of operation
above which ore or waste are excavated from a continuous face.
Blading - Digging and pushing dirt without picking it up.
Berm - A horizontal shelf or ledge built into an embankment of sloping wall
of an open pit, to break the continuity of a long slope to increase
the stability.
Burn - To pulverize first with very heavy explosive charges.
Burn Cut - A drill hole pattern widely used in fast moving tunnels. Holes
are uncharged and serve as a relief zone.
Breasting - A short leading stall worked at right angles to, and forming
the face of the main levels.
Breast Stoping - A method of stoping employed on veins where the dip is
not sufficient for the broken ore to be removed by gravity.
Borehole - A hole made with a drill, auger or other tools used in mining.
Backfill - Waste sand or rock used to support the roof of a mine after
removal of the ore.
Cathode - The. electrode where electrons enter (current leaves) an operating
system such as an electrolytic cell.
Cementation - The action of precipitating copper from a solution of copper
sulfate by addition of iron to the solution.
Coffer Dam - A temporary watertight enclosure from which the water is pumped
to expose the bottom of a body of water and permit construction
as of foundations or piers.
.380
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Concentration - Separation and accumulation of economic minerals from gangue
(sometimes called milling).
Concentrator - A plant where ore is separated into concentrates and tails.
Concentrate - The valuable material recovered from ores.
Competent - Rock formations in which no artificial support is needed to
maintain a cave free borehole.
Countercurrent Decantation - The clarification of water and the concentration
of tailings by the use of several thickeners in
series. The water flows in the opposite direc-
tion from the solids.
Cyclone Classifier - A device for classification by centrifugal means of
fine particles suspended in water, whereby the coarser
grains collect at and are discharged from the. apex of
the vessel, while the finer particles are eliminated
with the bulk of the water at the discharge orifice.
Drifter - A person skilled in the use of air-driven percussive rock drills
and other processes used in excavating tunnels or passages.
Drift - A horizontal passage underground which follows the vein of ore.
Dike - An embankment of earth or stone used to contain concentrator tailings.
Dump Leaching - Term applied to dissolving and recovering minerals from
sub-grade ore materials from a mine dump. The dump is
irrigated with water, sometimes acidified, which perco-
lates into and through the dump, and runoff from the
• bottom of the dump is collected and mineral in solution is
recovered by a chemical reaction.
i
Electrolysis - Chemical change resulting from the passage of an electric
current through an electrolyte.
Electrolyte - A nonmetallic electrical conductor in which current is carried
by the movement of ions instead of, electrons with the libera-
tion of matter at the electrodes.
v
Froth Flotation - A flotation process in which the minerals floated gather
in and on the surface of bubbles of air or gas.
Gangue - Undesired minerals associated with ore, mostly nonmetallic. It is
usually rejected as tailings in concentrating ores.
381
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Heap Leaching - A process used for the recovery of copper from weathered
ore and material from mine dumps. The material is laid
in beds alternately fine and coarse until the thickness
is roughly 20 ft. It is treated with water or the spent
liquor from a previous operation. Intervals are allowed
between watering to allow oxidation to occur. The liqiidl?
seeping through the beds is fed to tanks, where it is
treated with scrap iron to precipitate the copper from
solution.
Hummocky Surface - A surface which is lumpy or has small uneven knolls.
In-situ Leaching - Leaching conducted on an ore body that is not removed
from its original or natural position.
Keyway - A structure used in tailings pond dams to provide strength and
stability. The keyway is constructed below and above ground
surface, and the dam is built .over the keyway.
Launder - An open channel or trough in which scrap iron is contained.
Pregnant liquor is channeled through the iron to precipitate
copper.
Lift - The method used to raise the level of ore, for leaching operations,
to build the final heap.
Long Wall Retreating - A mine system of long wall working in which the de-
veloping headings are driven narrow to the boundary
or limit line, and the ore seam is extracted by long
wall faces retreating towards the shaft.
Mesothermal'Deposit - A mineral deposit formed at moderate temperature and
moderate pressures, in and along fissures or other
openings in rocks, by deposition at intermediate
depths, chiefly from hydrothermal fluids derived
from consolidating intruding rocks.
Muck - Rock or ore broken in the process of mining.
Mucker - A laborer who loads broken ore into cars for transport.
Middlings - Particles incompletely liberated into concentrate or tailings.
Generally treated in another process, to produce a secondary
. concentrate.
Overburden - Material of any nature that overlies an ore deposit.
382
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Panel Mining - A system of ore extraction in which the ground is laid off
in separate districts or panels with pillars of extra size
being left between the panels.
Plutonic Rock - Igneous rocks formed deep within the earth under the in-
fluence of heat and pressure, solidified from a molten mass.
Pregnant Solution - Solution containing a recoverable concentration of metals,
Precipitation - The act of separating a solid .form from a solution by
chemical means.
Proctor Density - A rating given to soils after compaction tests have been
made. The result can be correlated with the moisture con-
tent and the load bearing ability of the soil.
Raise - A mine opening driven upward from the back of a level to a level
above.
Raffinate - The aqueous solution remaining after metal values have been
extracted by the solvent.
Refinery - A term sometimes applied to the plant in which metal or valuable
mineral is extracted from an ore or concentrate.
Rippers - An accessory that is either mounted or towed at the rear of a .
tractor and generally used in place of blasting as a means of
loosening compacted soils and soft rocks for scraper loading.
The ripper has long, angled teeth that are forced into the
ground surface, ripping the earth loose.
Scraper - A surface vehicle mounted on large rubber-tired wheels, fitted
at the bottom with a cutting blade, and pushed by tractor.
As the vehicle moves along, the cutting blade loads a bucket
contained in the vehicle. When full the scraper is transported
to a dumping point where the material is discharged through the
bottom of the vehicle in an even layer.
Skip Hoist - A bucket or car operating up and down a defined path, receiving,
elevating, and discharging bulk materials.
Slimes - A material of extremely fine particle size encountered in ore treat-
ment. For example, a product of wet-grinding containing valuable
ore in particles so fine, as to be carried in suspension by water.
383
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Smelter - A plant where ore concentrates are melted in a furnace in order
to recover the metal values.
Spent Liquor - Liquor that has had all metallic values removed by processing.
Solvent Extraction - A method of separating one or more substances from a
mixture by treating a solution of the mixture with a
solvent that will dissolve the required substance.
Stabilization - Chemical, mechanical or vegetative treatment designed to
increase or maintain the stability of soil and improve its
engineering properties.
Stope - An excavation from which ore has been taken in a series of steps.
Stoping - Excavating ore by means of a series of horizontal, vertical, or
inclined workings in veins or large irregular bodies of ore, or
by rooms in flat deposits.
Stopping - A permanent wall built to close off the unused, or no longer
needed crosscuts in a mine.
Tailings - The portions of concentrated ore that are too poor in mineral
values to be treated further. Disposed of as a waste material.
Tail Water - Water immediately downstream from a structure.
Thickeners - A vessel or apparatus for reducing the proportion of water in
a pulp.
Tuff - A rock formed of compacted volcanic fragments generally smaller than
four millimeters in diameter.
Vat Leaching - Leaching of copper ore with sulfuric acid in an open tank'.
Waste Rock - Barren or submarginal rock which has been mined, but is not of
sufficient value to warrant treatment and is therefore removed
ahead of the concentrator.
Wet-grinding - The milling of materials in water or other liquid.
384
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UNITS AND CONVERSION FACTORS
cm
in.
ft
m
mile
km
ft2
m2
acre
hectare
gal.
gal.
cu m
Ib
kg
oz (troy)
oz (troy) /ton
ton
MT
long ton
MT
Multiply by
2.54
0.3937
0.3048
3.281
1.609
0.6214
0.0929
10.764
0.4047
2.471
0.0283
35.3356
3.785
0.003785
264.2
0.4536
2.2046
31.103
0.0342
0.907
1.1023
1.016
0.9842
in.
cm
m
ft
km
mile
m2
ft2
hectare
acre
m3
ft2
liter
cu m
gal.
kg
Ib
g
kg/Ml
MT
ton
MT
long ton
vio!454
385
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