SW-145c.2 Prepublication issue for EPA libraries
and State Solid Waste Management Agencies
ASSESSMENT OF INDUSTRIAL HAZARDOUS WASTE PRACTICES
IN THE METAL SMELTING AND REFINING INDUSTRY
Volume II
Primary and Secondary Nonferrous Smelting and Refining
This final report (SW-245c.2) describes work performed
for the Federal solid waste management program
under contract no. 68-01-2604
and is reproduced ae received from the contractor
The report is in four volumes: (1) Executive Summary, (II) Primary
and Secondary Nonferrous Smelting and Refining, (III) Ferrous Smelting
and Refining, and (IV) Appendices
Copies will be available from the
National Technical Information Service
U.S. Department of Commerce
Springfield, Virginia 22161
U.S. ENVIRONMENTAL PROTECTION AGENCY
1977
Rv
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This report has been reviewed by the U.S. Environmental Protection
Agency and approved for publication. Its publication does not signify
that the contents necessarily reflect the views and policies of the U.S.
Environmental Protection Agency, nor does mention of commercial products
constitute endorsement or recommendation for use by the U.S. Government.
An environmental protection publication (SW-145c.2) in the solid waste
management series.
n
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ABSTRACT
investigations of on-land disposal of process and pollution
control residuals from the United States metal smelting and refining
industry were conducted. This volume presents the results of studies
of the U.S. primary and secondary non-ferrous smelting and refining
industry. Characteristics of each industry sector, including plant
locations, production capacities, and smelting and refining processes,
have been identified and described.
Land-disposed or stored residuals, including slags, dusts, and
sludges have been identified and characterized for physical and chemical
properties. State, regional, and national estimates have been made of
the total quantities of land-disposed or stored residuals and potentially
hazardous constituents thereof.
Current methods employed by the primary metals industry for
the disposal or storage of process and pollution control residuals on
land are described. Principal methods include lagoon storage of sludges
and open dumping of slags. Methods of residual treatment and disposal
considered suitable for adequate health and environmental protection
have been provided.
Finally, the costs incurred by typical plants in each primary
sme]ting and refining category for current and environmentally sound
potentially hazardous residual disposal or storage on land have been
estimated.
m
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ACKNOWLEDGMENTS
The EPA Project Officers responsible for overall direction of
this program were Messrs. Allan Pearce and Timothy Fields, Jr., Office
of Solid Waste, Hazardous Waste Division. Technical program performance
was vested with Calspan Corporation, Buffalo, New York. The Calspan
Project Engineer was Mr. Richard P. Leonard. Mr. Richard Brown, Calspan
Consultant, assumed technical responsibility for investigations in Primary
Copper, Lead and Zinc Smelting. Dr. John Y. Yang of Calspan was respon-
sible for technical program performance in the Secondary Metal Smelting
;iml Refining Industry. Dr. Robert Ziegler, Calspan Consultant, was
iv.|>oir: i hi c for performance in the Primary Aluminum Sector. Mr. Hans
Ki-if
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UBLE OF CONTENTS
i jon Page
ABSTRACT i 1 i
ACKNOWLEDGEMENTS i V
LIST OF FIGURES Vli
LIST OF TABLES i X
I . CONCLUSIONS 1
I I. PRIMARY NON-FERROUS SMELTING AND REFINING 3
1. Primary Smelting and Refining of Copper (SIC 3331) 4
2. Primary Smelting and Refining of Lead (SIC 3332) 44
3. Primary Smelting and Refining of Zinc (SIC 3333) 64
4. Primary Smelting and Refining of Aluminum (SIC 3334) 103
5. Primary Smelting and Refining of Antimony (SIC 3339) 132
6. Primary Smelting and Refining of Mercury (SIC 3339) 153
7. Prim;iry Smelting and Refining of Titanium (SIC 3339) H>(>
H. I'i'im.try Sjiu-11 Inj.; and Refining of Tungsten (SIC 3339) 178
9. Primary Smelting and Refining of Tin (SIC 3339) 194
10. Primary Smelting and Refining of Magnesium (SIC 3339) 206
11. Primary Smelting and Refining of Cadmium (SIC 3339) 210
12. Primary Smelting and Refining of Arsenic (SIC 3339) 220
13. Primary Smelting and Refining of Selenium and Tellurium
(SIC 3339) 221
14. Primary Smelting and Refining of Gold and Silver (SIC 3339) 224
15. Primary Smelting and Refining of Platinum (SIC 3339) 227
16. Primary Smelting and Refining of Bismuth (SIC 3339) 229
17. Primary Smelting and Refining of Cobalt (SIC 3339) 231
18. Primary Smelting and Refining of Zirconium and Hafnium
(SIC 3339) 233
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TABLE OF CONTENTS (Cont'd)
Section Page
III. SECONDARY NON-FERROUS SMELTING AND REFINING 235
1. Secondary Smelting and Refining of Copper (SIC 33412) 239
2. Secondary Smelting and Refining of Lead (SIC 33413) 262
:!>. Soeondary Smelting and Refining of Aluminum (SIC 33417) 283
IV. I.I ST OF Rl-raUNCI-S i 309
VI
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LIST OF FIGURES
Figure No. Page
1 Primary Copper Smelting and Fire Refining 10
2 Primary Electrolytic Copper Refining 11
3 Primary Lead Smelting and Refining 49
4 Primary Electrolytic Zinc Refining 68
5 Primary Pyrometallurgical Zinc Smelting and Refining 70
6 Process Diagram for Primary Aluminum Smelting 107
7 Primary Aluminum Production 109
8 Pyrometallurgical Antimony Smelting and Refining 134
9 Electrolytic Antimony Refining 136
10 Primary Mercury Production 156
11 Primary Titanium Sponge Metal Production 168
12 Primary Tungsten Production 180
13 Primary Tin Smelting and Refining 195
14 Production of Magnesium from Sea Water 207
15 Production of Magnesium by Ferrosilicon Process 208
10 Cadmium Recovery in the Zinc Retort Process 213
17 Cadmium Recovery from Zinc Metal 215
JS Cadmium Recovery from Lead, Zinc or Copper Smelter Dusts .. 216
19 Cadmium Recovery from Purification Sludge of Zinc Electro-
winning Operations 218
20 Primary Gold and Silver Production 225
21 Primary Cobalt Smelting and Refining 232
22 Secondary Copper Pyrometallurgical Smelting and Refining .. 241
23 Secondary Copper Electrolytic Refining 243
24 Secondary Soft and Hard Lead Smelting 265
vil
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LIST OF FIGURES (Con't)
Figure No. Page
25 Secondary Anitmonial Lead Smelting 266
26 Secondary White Metal Smelting 268
27 Secondary Aluminum Smelting 287
viii
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LIST OF TABLES
vf,, Page
J Primary Copper Smelters and Refineries 5
2 fJeographical Distribution of Primary Copper Smelters and
Re finers o
.'•; Waste Residual Factors, Primary Copper Smelters and Fire
Refiners 16
4 Yearly Generation of Residuals from Typical Plant Copper
Smelting and Fire Refineries 16
5 Waste Residual Factors, Primary Copper Electrolytic
Refineries 17
6 Yearly Generation of Waste Residuals from Typical Plant,
Copper Electrolytic Refineries 17
7 Estimated State, Regional and National Solid Waste from
Primary Copper Smelters and Fire Refiners 18
8 Estimated State, Regional and National Solid Waste from
Primary Copper Electrolytic Refiners 25
9 Treatment and Disposal Technology Levels, Primary Copper
Smelters and Fire Refiners, Dusts 31
10 Treatment and Disposal Technology Levels, Primary Copper
Smelters, Fire Refiners and Electrolytic Refiners-Sludges ..33
11 Cost of Level I Treatment and Disposal Technology -
Primary Copper Smelting and Fire Refining 36
12 Cost of Level I Treatment and Disposal Technology -
Primary Copper Electrolytic Refining 39
13 Cost of Level III Treatment and Disposal Technology -
Primary Copper Smelting and Fire Refining 40
14 Cost Summary for Treatment and Disposal Technology -
Primary Copper Smelters and Fire Refiners 42
15 Cost Summary for Treatment and Disposal Technology -
Primary Copper Electrolytic Refineries 43
16 Characteristics of U. S. Lead Smelters 45
17 Geographical Distribution of Primary Lead Smelting and
Refining Capacity 46
IX
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LIST OF TABLES (Cont'd)
Table No.
Page
18 Waste Residual Factors, Primary Lead Smelting and
Refining 53
19 Yearly Generation of Waste Residuals From Typical Plant,
Primary Lead Smelting and Refining 53
20 Hstimated State, Regional and National Solid Waste from
Primary Lead Smelting and Refining 54
21 Treatment and Disposal Technology Levels, Primary Lead -
Sludges 58
22 Costs of Level I Treatment and Disposal Technology,
Primary Lead Smelting and Refining 61
23 Cost Summaries for Treatment and Disposal Technologies,
Primary Lead Smelting and Refining 63
24 Characteristics of Primary Zinc Plants in the United States.. 65
25 Geographical Distribution of Primary Zinc Smelters and
Refiners 66
26 Solids in Limed Effluent Streams Present at an Electrolytic
Zinc Plant 71
21 Solids in Limed Effluent Streams Present at a Pyrometal-
lurgical Zinc Plant 72
28 Waste Residual Factors, Primary Zinc Electrolytic Refining 75
29 Waste Residual Factors, Primary Zinc Pryometallurgical
Refining 75
30 Yearly Generation of Waste Residuals From Typical Plant,
Primary Zinc Electrolytic Refining 76
31 Yearly Generation of Waste Residuals From Typical Plant,
Primary Zinc Pryometallurgical Refining 76
32 Estimated State, Regional and National Solid Waste from
Primary Electrolytic Zinc Smelting and Refining 77
33 Estimated State, Regional and National Solid Waste from
Primary Pryometallurgical Zinc Smelting and Refining 80
34 Treatment and Disposal Technology Levels, Primary Electro-
lytic Zinc - Sludges 86
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LIST OF TABLES (Cont'd)
Table No. Page
I 'JT-II I twiil ami l)i%j)(i-uil I'm liinilo^y l.nvnli, Primary Pyromolul"
lu!'j.;ii;4l J.IIK- I fun I'rt'.'i* Ke.Mduti.s 88
3o Treatment und Disposal Technology Levels, Primary Pyrometal-
lurgical Zinc - Sludges « 90
37 Cost of Level I Treatment and Disposal Technology, Primary
Zinc Plant - Electrolytic Reduction 93
38 Cost of Level 1 Treatment and Disposal Technology, Primary
Zinc Plant - Pyrometallurgical 96
31.' Cost of Level 111 Treatment and Disposal Technology, Primary
Zinc Plant - Electrolytic Reduction Plants 98
40 Cost of Level III Treatment and Disposal Technology, Primary
Zinc Plant - Pyrometallurgical Plants 99
41 Cost Summary for Electrolytic Zinc Smelting and Refining ...100
42 Cost Summary for Pyrometallurgical Zinc Smelting and
Refining 101
43 Geographical Distribution of Primary Aluminum Plants 104
44 Raw Material and Energy Requirements for Aluminum Production ... 106
45 Waste Residual Factors, Primary Aluminum Smelting and
Refining 112
46 Yearly Generation of Waste Residuals, Typical Plant, Primary
Aluminum Smelting and Refining 113
47 Estimated State, Regional and National Solid Waste from
Primary Aluminum Production 114
48 Treatment and Disposal Technology Levels, Primary Aluminum -
Scrubber Sludges 120
49 Treatment and Disposal Technology Levels, Primary Aluminum -
Spent Pot liners and Skimmings 122
50 Treatment and Disposal Technology Levels, Primary Aluminum -
Shot Blast Dust, Cast House Dust 124
51 Cost of Level I Treatment and Disposal Primary Aluminum Plant ..127
52 Cost of Level III Treatment and Disposal Technology, Primary
Aluminum Plant 129
XI
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LIST OF TABLES (Cont'd)
Table No. Page
53 Cost Suirmary for Primary Aluminum Smelting and Refining . . . 130
54 Geographical Distribution of, Primary Antimony Metal
Smelting and Refining ..................................... 133
55 Waste Residual Factors, Primary Antimony Smelting and
Refining .................................................. 138
f>(> Yearly limiorut ion ol' Waste Residuals, Primary Antimony..... 138
57 Kstimated State, Regional and National Solid Waste from
Primary Antimony Smelting and Refining .................... 139
58 Treatment and Disposal Technology Levels, Pyrometallurgical
Antimony - Blast Furnace Slag ............................. 142
59 Treatment and Disposal Technology Levels, Electrolytic Anti-
mony - Anolyte Sludge ..................................... 144
60 Cost of Level I Treatment and Disposal Technology, Primary
Antimony Pyrometal lurgical Plant .......................... 147
61 Incremental Costs for Land Sealing and Runoff Collection
for Slag Disposal ......................................... 148
62 Cost of Level 111 Treatment and Disposal Technology, Primary
Antimony Hloctrolytic Plant .......................... ...... ISO
<>3 Cost Summary for I'yrometullurgical Antimony Smelting and
Refining [[[ 151
64 Cost Summary for Primary Antimony Electrolytic Refining ..... 152
65 Salient Domestic Mecury Statistics ......................... 153
66 Geographical Distribution of Primary Mecury Production ..... 154
67 Waste Residual Factors, Primary Mercury Smelting and Refining . . 158
68 Yearly Generation of Waste Residuals from a Typical Mercury
Smelter [[[ I58
69 Estimated State, Regional and National Solid Waste from
Primary Mercury Production
Treatment and Disposal Technology Levels, Primary Mercury -
Coiulcnsci- Wastewater
71 Costs of Level I Treatment and Disposal Technology, Primary
Mercury Production .........................................
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LIST OF TABLES (Cont'd)
'I able \'o. Page
73 Geographical Distribution of Primary Titanium Sponge Metal
Plants .................................... . ................ l
74 Waste Residual Factors , Primary Titanium Refining .. ........ 170
75 Yearly Generation of Waste Residuals from Typical Titanium
Refining Plant ............................................. 170
76 Estimated State, Regional and National Solid Waste from
Primary Titanium Production ................................
77 Treatment and Disposal Technology Levels, Primary Titanium -
(.'hi orinator Sludges ................ , ....................... 173
78 Cost of Level I Troatmont and Disposal Technology, Primary
Ti tan iuni Rel'i n i MJ.; .......................................... 175
79 Cost Summary for Primary Titanium Refining ................. 177
80 Estimated State, Regional and National
Primary Tungsten Production ................................ 179
81 Waste Residual Factors, Primary Tungsten ................... 182
82 Yearly Generation of Wastes from Typical Tungsten Plant .... 182
83 Estimated State, Regional and National Waste from Primary
Tungsten Product ion ........................................
.si I'reat snout .ind Disposal Technology Levels, Primary Tungsten -
Digestion Residue .......................................... 185
85 Treatment and Disposal Technology Levels, Primary Tungsten -
Sludge [[[ 187
86 Cost of Level I Treatment and Disposal Technology, Primary
Tungsten Plant ............................................. 190
87 Cost of Level II Treatment and Disposal Technology, Primary
Tungsten Plant ..... . ....................................... 191
88 Cost Summary for Primary Tungsten Refining ................. 193
89 Waste Residual Factors, Primary Tin Smelting and Refining .. 197
90 Yearly Generation of Waste Residuals for Typical Plant,
Primary Tin Smelting and Refining .......................... 197
91 Estimated State, Regional and National Waste from Primary Tin
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LIST OF TABLES (Cont'd)
Table No. Page
'.)2 Trcutmont and Disposal Technology Levels, Primary Tin -
Smelting Slag 200
93 Costs of Level I Treatment and Disposal Technology, Primary
Tin Smelting and Refining 203
94 Cost Summary for Primary Tin Smelting and Refining 204
95 Trace Element Composition of Limestone . 209
96 Producers of Cadmium Metal 210
97 Geographical Distribution of Primary Cadmium Production ... 211
98 Analyses of Cadmium Plant Residue ,, 219
99 United States Producers of Selenium and Tellurium Metals and
Compounds 221
100 Geographical Distribution of Secondary Copper Smelters .... 240
101 Waste Residual Factors, Secondary Copper Smelting and
Refining 245
102 Yearly Generation of Waste from Typical Plant, Secondary
Copper Smelting and Refining 246
103 Estimated State, Regional and National Solid Waste from
Secondary Copper and Brass and Bronze Smelting 247
1U4 Treatment and Disposal Technology Levels, Secondary
Copper - Blast Furnace Slag 251
105 Treatment and Disposal Technology Levels, Sc^'i.^ry
Copper - Copper-F.lectrolytic Sludge 2Sj>
106 Cost of Level I Treatment and Disposal Technology,
Secondary Copper Blast Furnace 256
107 Cost of Level I Treatment and Disposal Technology, Secondary
Copper, Electrolytic Refining 257
108 Cost of Level III Treatment and Disposal Technology,
Secondary Copper, Electrolytic Refining 259
109 Cost Summary for Pyrometallurgical Smelting and Refining
of Secondary Copper 260
110 Cost Summary for Electrolytic Smelting and Refining of
Secondary Copper 261
Xiv
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1.1ST 01-' TAULhS (Cont'dj
Tahle_Nu. Pago
111 Geographic Distribution of Secondary Lead Production
Capacity 263
112 Waste Residual Generation Factors, Secondary Lead Smelting
and Refining 270
113 Annual Generation of Waste Residuals from Typical Plants,
Secondary Lead Smelting and Refining 271
1H l-.sti mated State, Regional and National Solid Waste from
Secondary Lead and White Metal Smelting 273
III) Treatment and Disposal Technology Levels, Secondary Lead -
SO Scrubwater Sludge and Acid Leach Sludge 277
116 Cost of Level I Treatment and Disposal Technology Secondary
Antimonial and Soft Lead 280
117 Cost Summary for Secondary Antimonial and Soft Lead
Smelting and Refining 282
118 Geographical Distribution of Secondary Aluminum Smelters . 284
119 Regional and National Production Capacity for Secondary
Aluminum 286
120 Waste Generation Factors, Secondary Aluminum Smelting and
Kofi n ing 290
!,.'! Yearly Generation of Waste Residuals from Typical Secondary
Aluminum Smelting 290
122 Estimated State, Regional and National Solid Waste from
Secondary Aluminum Smelting 291
123 Treatment and Disposal Technology Levels, Secondary
Aluminum - Scrubber Sludge 299
124 Treatment and Disposal Technology Levels, Secondary
Aluminum - High Salt Slag 301
125 Cost of Level I Treatment and Disposal Technology, Secondary
Aluminum Plant - Reverberatory Smelting 304
l~'(> Cost Summary for Secondary Aluminum Reverbatory Smelting ... 306
127 Cost Summary for Secondary Aluminum Dross Smelting 307
XV
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SECTION I
CONCLUSIONS
The primary and secondary non-ferrous smelting and refining
industry disposes or stores large quantities of process and pollution
control residues on land. These residuals are predominantly inorganic
slags and sludges containing silicates, oxides and sulfates, and chlorides
in some industries. Sludges are often residues of SO- scrubwater
treatment or process wastewater treatment with lime. Consequently,
they contain calcium sulfates, calcium sulfites, calcium hydroxides,
and calcium carbonates. It was found that recycling of dusts from
emissions controls was a relatively common practice, although some
dusts are disposed on land. The high metallic content of dusts often
allows their recycle to the production process, an economically attractive
and viable alternative to land disposal.
The prJncipal potentially hazardous constituents found in
primary and secondary smelting residuals are heavy metals, including
arsenic, cadmium, lead, zinc, copper, chromium, antimony, and nickel.
Because of trace amounts in the concentrates and ores from which the
metals are recovered, the primary base metal smelting and refining
industries (i.e., lead, copper, zinc, antimony, mercury, tungsten, and
tin) produce a wider variety of heavy metals in residues, including
arsenic, cadmium, lead, zinc, copper, antimony, nickel and mercury.
Bauxite, the ore from which aluminum is recovered, is essentially
devoid of toxic heavy metals. Fluorides and very small amounts of cyanide
appear in sludge and potlincr residues because of the fluoride contained
in input cryolite (Na,AlF ), and cyanide produced in potliner consumption.
Magnesium produced from electrolysis of seawater or from dolomitic
limestone also uses heavy metal deficient raw materials and will have
negligible concentrations of heavy metals in residuals.
The predominant practices used in the primary and secondary
non-ferrous smelting and refining industries for disposal of residuals
arc lagooning and open dumping. Slags and other solid residues are
generally open dumped on land. Scrubwater from wet emissions control
and process wastewater with or withoul lime treatment is generally
rout nl to tin lined settling pits or to unJinod lagoons. Settled sludge is
oft i-n dredged from pits or lagoons and stored or disposed of on land.
Industries which produce relatively small quant ities of sludge will often
leave sludges permanently in lagoons. The use of unlincd settling pits
and lagoons is the predominant practice.
The use of lined lagoons for sludge settling, chemical fixation
of sludge before land disposal, and sealing of ground areas before
hazardous waste disposition arc techniques for adequate environmental
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protection. In a number of industries including primary copper, zinc
and lead, immediate recycle or shipment of residuals to be further
processed for metallics recovery (i.e., dusts, sludges) will preclude
leaching of potentially hazardous heavy metal constituents.
Although the presence of potentially hazardous constituents
in slags, sludges and dusts has been shown, it is generally not known
if these materials are leached in disposal environments. In the event
that significant leaching is demonstrated, a number of suggestions are
made with respect to practices for adequate health and environmental
protection. It is suspected that sludges and dusts may leach heavy
metals and other potentially hazardous constituents to a greater extent
than most slags because of fine particle size and consequent suscepti-
bility to weathering processes. Solubility tests conducted on collected
waste samples and described in Appendix B generally supported this
hypothesis.
Recommended practices in the event of demonstrated signifi-
cant leaching of potentially hazardous constituents include the use of
lined lagoons for storage or permanent disposal of sludges. Leachable
sludges which are dredged or pumped from lagoons or settling pits and
dumped on land can often be chemically "fixed" so that leaching of
heavy metals can be prevented. Alternatively, sealing of soil in
disposal areas with bentonite or other sealants should prevent leachate
percolation.
In the event that slags, calcine residues, retort residues,
or other land-disposed or stored solid residues are shown to leach signi-
ficantly, then soil sealing at disposal or storage areas would be needed.
(Ail lection of runoff from disposal dumps containing slags, sludges or
dusts with leachablo heavy metals or other potentially hazardous consti-
tuents may be needed. Collected runoff would require treatment before
discharge or retention and evaporation ir> lagoons.
In a number of industry sectors, it was fouiid that some sludges,
dusts, or other residues are stored on open ground for periods ranging
from months to years before processing for further metallic recovery.
In such cases, immediate recycle- or storage in concrete pits before
reprocessing will preclude leaching of potentially hazardous constituents.
In some industries the use of dry air pollution control systems
can greatly reduce or eliminate the quantity of land-disposed waste.
l:\ampJes are the use of dry absorption beds in the primary aluminum
industry and silica-impregnated baghouses in the secondary aluminum
smelting sector.
Future air and water pollution controls are expected to
increase the quantities of land-disposed sludges, particularly sulfite
and sulfate sludge residues from control of S02 emissions from primary
copper and secondary lead smelters.
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SI-CHUN li
I'UIMARV NON-FliKKOllS SMHLlINCi AND REFINING
This section presents the results ot investigations and
analyses ot on-land disposal or storage of process and pollution
control residuals from the United States primary non-ferrous smelting
and refining industry. Characteristics of each industry sector includ-
ing plant locations, production capacities and smelting and refining
processes have been identified and described.
Land disposed or stored residuals including slags, dusts
and sludges have been identified and characterized physically and
chemically. State, regional and national estimates have been made of
the total quantities of land disposed or stored residuals and poten-
tially hazardous constituents thereof.
Current methods employed by the primary metals industry for
the disposal or storage of process and pollution control residuals on
land are described. Principal methods include lagoon storage ot
sludges and open dumping of slags. Methods of residual treatment
and disposal considered suitable for adequate health and environmental
protection have been provided. Finally, the costs incurred by typical
plants in each primary smelting and refining category for current and
environmentally sound residual disposal or storage on land have been
estimated.
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PRIMARY SMELTING AND REFINING OF COPPER (SIC 3331)
INDUSTRY CHARACTERIZATION
The major U.S. copper smelting and refining companies are
vertically integrated and have mining, smelting, refining, fabricating
facilities, and marketing organizations. Other large producers mine
and process through the smelting or refining stage and other companies
mine and concentrate their ore and ship the product to custom plants for
smelting and refining.
Copper smelting capacity in the United States in 1974 totaled
8,<)
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ho.s>t_rn£. The object of roasting copper sulfide ores and concentrates
is to regulate the amount of sulfur so that the material can he
effjcicutly smelted and to remove certain volatile impurities such as
antimony, arsenic, and bismuth.
SiiKiJ_tijri^. Roasted and unroasted materials are smelted after mixing
with suitable fluxes in reverberatory furnaces. In this reducing or
neutral atmosphere, copper and sulfur form copper sulfide and iron'
sulfide. The combination of the two sulfides known as copper matte
collects in the lower area of the furnace and is removed. Mattes
commonly contain 40 to 45 percent copper. Impurities such as sulfur,
antimony, arsenic, iron, and precious metals are in the matte. The
remainder of the molten mass containing most of the other impurities
floats on top of the matte and is drawn off and discarded as a slag.
Converting is the final stage in the process of smelting
and is accomplished by blowing thin streams of air through the molten
matte in a refractory-lined converter to oxidize the ferrous sulfide
to sulfur dioxide, to eliminate the sulfur as a gas, and also to form
a ferrous slag containing trace metal impurities. When the converter
cycle is finished, the converter is tilted to discharge the lighter
sJag and then the relatively pure copper metal, referred to as blister
copper, into ladles in which it is transferred to a holding furnace
and then to a poling furnace for fire-refining or casting into anodes
for electrolysis.
The blister copper produced by smelting is too impure for
most applications and requires refining before use. It may contain
silver and gold, and other elements such as arsenic, antimony, bismuth,
lead, selenium, tellurium, and iron. Two methods are used for refining
copper - fire refining and electrolysis.
Fire Refining. The fire-refining process employs oxidation, fluxing,
and reduction. The molten metal is agitated with compressed air.
Sulfur dioxide is liberated and some of the impurities form metallic
oxides which combine- with added silica to form slag. Sulfur, zinc,
tin, and iron are almost entirely eliminated by oxidation. Lead,
arsenic, and antimony can be removed by fluxing and skimming as a dross.
Copper oxide in the melt is reduced to metal by inserting green wood
poles below the bath surface (poling). Reducing gases formed by
combustion of the pole convert the copper oxide in the bath to copper.
II" t he original material dm1-, not contain sufficient i;oKl or silver to
w.irr.mt its recovery, or if a special purpose silver-containing copper
is desired, the fire refined copper is cast directly into forms for
industrial use. If it is of such a nature as to warrant the recovery
of the precious metals, the fire refining is not carried to completion
but only far enough to insure homogeneous anodes for subsequent elect-
rolytic refining.
-------
It I eel miy i n KM'in ing. In the electrolytic rofinlng stop tuiotlos mid
fit Hindus (ih!n copper starting shoots) f»re hung alternately In concrete
electrolytic ceils containing tho eleetroiyte which Is essentially a
solution of copper sulfate and sulfuric acid. When current is applied,
copper is dissolved from the anode and an equivalent amount of copper
plates out of solution on the cathode. Such impurities as gold, silver,
platinum-group metals, and the selenides and tellurides of metal fall to
the bottom of the tank and form anode slime or mud. Arsenic, antimony,
bismuth, and nickel enter the electrolyte. After the plating cycle is
finished, the cathodes are removed from the tanks, melted, and cast into
commercial refinery shapes. The copper produced has a minimum purity of
99.9 percent.
1.2 WASTh CHARACTERIZATION
This section contains descriptions of production technology
at typical primary copper smelters and refiners and the resultant
byproducts or wastes which are either recycled directly, shipped to
other smelters for further metallic recovery, or disposed of on land.
Estimates are given for the quantities of wastes and potentially hazard-
ous constituents thereof which are disposed of on land either in lagoons
or open dumps. Smelters with fire refining and electrolytic refiners
are considered separately.
1.2.1 Process Description
Smelting and Fire Refining. The hypothetical copper smelting
and fire refining plant is assumed to have an annual capacity of 100,000
Ml f110,000 short tons) per year. Approximately half of the U.S.
copper smelters are smaller and half larger than this figure, and the
average output ^ap.tcity of all plants is within IS percent . i' this figure.
Copper smelters are estimated to operate 350 days/year; the :^i ly
capacity is, therefore, 286 MT (314 short tons) per day.
Smelter feed material, although composed chiefly of copper
sulfide concentrates derived from porphyry ores, also includes a
variety of copper-bearing byproducts from lead and z,inc smelters as
well as diverse kinds of copper-bearing concentrates from nonporphyry
ores. Hence, the smelter feed may range in chemical composition as
follows: Copper, 25 to 35 pet; sulfur, 27 to 33 pet; iron, 20 to 30
pet; and combined gangue constituents, 10 to 20 pet. Further, typical
smelter feeds frequently contain trace to small amounts of several
or perhaps all of the following elements: thallium, bismuth, cadmium,
tellurium, selenium, platinum metals, indium, nickel, cobalt, arsenic,
antimony, lead, and zinc.
-------
A :• inip 1 i tied flow diagram identifying solid waste sources
and disposal is shown in Figure 1. The primary process steps comprise
(1) roasting to reduce sulfur content, (2) reverberatory furnace smelt-
ing to form copper sulfide matte and a siliceous slag which is discarded,
(3) oxidation (blowing) the molten sulfide matte to form molten "blister"
copper and an iron silicate slag which is returned to the reverberatory,
and (4) furnace purification (fire-refining) of the molten copper such
that anodes suitable for electrolytic refining can be cast (or copper
product ran he marketed directly).
Roaster gases are rich in SO, such that sulfuric acid can be
produced in large quantities; some of fhe acid is used in the adjacent
electrolytic refinery. Some copper smelters do not roast prior to
reverberatory furnace smelting but do make sulfuric acid from con-
verter gases; some smelters do not have adjacent electrolytic copper
refineries. Acid plant blowdown slurry effluent yields a sludge
which is not 100 percent recycled, as will be discussed in a following
sect i on.
I! 1 ret roU't ic Refining. The hypothetical copper electrolytic
refining plant is assumed to have an annual capacity of 160,000 MT
(17h,0()0 short tons) per year. This is the approximate average of all
electrolytic refinery capacities in the U.S. Copper refineries are
estimated to operate 350 days per year; the daily capacity is, there-
fore, -157 MT (503 short tons) per day.
Electrolytic refinery feed material consists solely of cast
copper anodes having a purity in the range 99.0 to 99.7 percent copper.
The cathode copper product resulting from electrolysis has a purity of
the order of 99.95 percent.
A simplified flow diagram identifying solid waste sources
and disposal is shown in Figure 2. The primary process steps comprise
(1) electrolysis in large lead or plastic lined cells, (2) melting and
refining with respect to sulfur, hydrogen, and oxygen, and (3) casting
into shapes such as wirebar,or into continuous cast rod for the wire
drawing industry.
Important auxiliary process steps are slimes recovery from
the cell bottoms, and electrolyte purification to permit electrolyte
reuse and to recover materials of value. The slimes tend to be rich
in So, Te, As, Ag, Au and Pt such that their value is very high; the
typical refinery ships the slimes filter cakes elsewhere for treatment
to recovery these metals. (Three of the existing electrolytic copper
refineries in the U.S. treat slimes from their own operations and from
other refineries). Since slimes yield is only of the order of 0.001
metrn ion per metric Ion of copper rofinod, it is apparent that the
slimes recovery "industry" in the U.S. is very small relative to copper
production. Considering this and the fact that the only solid residue
-------
COPPER SULFIDE
CONCENTRATES
SILICA
REVERTS
UNDERFLOW
RECYCLED
TO FEED
PREP
OVEHFLOW
fO TAILINGS POND
(DISSOLVED SOLIDS
CONTENT 140 kg
PER METRIC TON
COPPER PRODUCTI
MISCELLANEOUS
DUSTS NOT
IMMEDIATELY
RECYCLED (ROASTER,
REVERBERATORY
FURNACE CONVERTER)
/ STORED \
/TEMPORARILYX
\ ON LAND /
\ 17k« /
ANODE COPPER PRODUCT
NUMERICAL VALUES ARE
IN KILOGRAM/METRIC TON
OF COPPER PRODUCT
Fiquro 1 PRIMARY COPPER SMELTING AND FIRE REFINING
10
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of note produced in slimes treatment is a slag which is recycled to a
copper smelter, it is clear that the total amount of solid waste gen-
erated in slimes treatment is insignificant in the context of the
present study.
I; 1 cc t ro lytc purification consists of the following steps
per I'(irnicd on the bleed stream of impure sulfuric acid containing copper
and other elements in solution: (1) copper removal by electrolysis xvith
insoluble lead anodes in liberator cells, (2) filtration to remove an
arsenical sludge, and (3) evaporation to precipitate nickel sulfate. A
"black acid" product remains after nickel sulfate removal. This is
marketed for acid recovery.
Slags taken from the melting and refining furnace (and from
the anode casting furnace at refineries which cast anodes) are very rich
in copper (50'ii Cu) , i.e., immediate recycle to a copper smelter is
dictated and practiced.
1. 2.2 Description of Waste Streams
Smelting and i:ire Refining. Three general types of solid
residues containing potentially hazardous elements are found at copper
smelters, namely, slag, dry dusts, and wet sludges and slurries. The
amount of slag produced and disposed per unit of copper production can
be accurately estimated from the data obtained from nine smelters by
Calspan. Average factors for the estimating generation or disposal of
dusts and sludges may be departed from significantly at individual
plants for the following reasons: (1) the amount of dust generated and
collected varies greatly with smelting practice and dust collection
practice; (2) complete recycle of dusts may prevail at many or all
smelters, but there are at least some instances in which dusts are
stored on the ground prior to recycle; (3) sludges stemming fron slurries
generated by operations, such as acid plant blowdown, contact cooling
and washdown, appear to vary widely in amount and disposition. For
example, the more efTective the electrostatic precipitator preceding the
acid plant, the less will be the particulate content of the blowdown
from the acid plant, and probably the less will be the volume of the
blow down stream necessary to maintain the acid plant in operating
equjiibrium. However, even with less effective electrostatic precipi-
tator:-,, if the- thickener on the blowdown stream is effective and if
the partjculate- laden underflow from the thickener is recycled directly
to, ',ay, reverberator}- feed preparation, the amount of suspended and
tiis-.olved solids going to u settling pond, neutralisation plant, and/or
f.-i i I in}'/, pom! mav be relatively small; (4) acid effluent j'.enorat i on am!
nciiii ili aliun prjititr i'. «. urrcnt ly ill U State ol ver) i.ipul i. h.inge ,it
a high percentage of the copper ".melters of the United States. This
sjtuation arises mainly from the enforcement of air and water pollution
regulation. Many acid plants have been added to reduce SO- emission
12
-------
from stacks, and lime neutralization plants are being added to reduce
the acid and dissolved metal contents of effluents to streams and to
facilitate total recycle of water.
In view of the above, the following solid residue generation
and disposal factors arc presented as shown in Figure 1.
a. Reverberator/ Slag
The chief solid waste from the hypothetical copper smelter
is reverberatory furnace slag amounting to 3000 kg/NTT of refined copper
or 2!)'",, 000 MT/year. Molten slag is transported about 600 m (2000 ft) to
the IIKJ i n dump by refract ory- 1 ined rail-car ladles. The accumulated slag
pile lias a volume of the order of 1.5 x 10 cu m (2 x 10 cu yards).
!). Dusts l:rom Roasters, Reverbatory Furnaces and Converters
Large amounts of dust are collected from each of the primary
process operations and are generally recycled without temporary storage
on land. It is estimated, however, that 17 kg of dust/NTT of refined
copper or 4.8 MT/day are stored on land near the smelter prior to
eventual recycle. The size of the storage pile is estimated to be
approximately 150 MT.
c. Acid Plant Slowdown Slurry (Sludge)
Slowdown from the acid plant amounts to 7,940 liter/MT of
copper product or 2,270 cu m/day (600,000 gal/day). The blowdown stream
flows to a thickener which yields an underflow of 82 cu m/day (21,600
gal/day) containing 46 MT/day of suspended solids. All underflow is
recycled immediately to feed preparation, it is in no sense a waste.
Overflow from the thickener amounts to 2,200 cu m/day (570,000
gai/dayj, containing 0.77 MT/day suspended solids and 40 MT/day dissolved
solids. The suspended solids are settled in an unlined lagoon, dredged
periodically, stored on the ground near the lagoon, mid eventually
recycled.
Affluent from the lagoon goes to the main tailings pond
associated with the mine-mill-smelter complex. This effluent contains
140 kg of dissolved solids per MT copper product or 40 MT/day dissolved
solids which tend to be precipitated to an unknown degree in the
tailings pond, by reaction with the alkaline tailings and/or evaporation.
The weight or volume of tailings is overwhelmingly greater than that of
the precipitated solids, but the dissolved solids from acid plant blow-
down are nonetheless present. (No significant portion of the cost of
the tailings pond is attributable to smcltor waste disposal because of
overwhelming volume of tailings waste).
-------
d. Miscellaneous Slurries (scrubbers, contact cooling, washing)
Effluents arising from sources such as wet scrubbers, anode
cooling, other contact cooling water, and plant washdown carry dissolved
and suspended solids amounting to 4.8 MT/day. The dredged sludge amount-
ing to 17 kg/MT of copper is dried on the ground near the lagoon and
eventually recycled. It is estimated that the average size of the
accumulation drying on the ground at any time is 800 MT. The dissolved
solid content of this effluent stream is unknown; the lagoon overflow
goes to the main tailings pond where some degree of precipitation of
dissolved solids would he expected as discussed above.
I; 1 e c trp ly t_i_c_ _Ku fj rung.
(a) Miscellaneous Slurries. The only waste ultimately land
disposed or stored on land in electrolytic copper refining is a combined
dilute slurry composed of effluents from operations such as contact cooling
of cast copper anodes and/or product, non-contact cooling of furnaces,
spent anode washing, and plant washdown. This dilute acid slurry containing
suspended and dissolved solids goes to an unlined settling lagoon approxi-
mately one-half acre in area. Flow is 6,800 cu m/day, suspended solids
content is 210 kg/day, and dissolved solids content is 870 kg/day. The
suspended solids settle in the lagoon and are dredged as necessary and
disposed of on land. Dissolved solids pass on with the overflow to the main
tailings pond (associated with the nearby ore concentrator) where an unknown
degree of precipitation would he expected due to evaporation and reaction
wild tin- tailings. At some ret'inereie.s located on navigable waters and not
un-'.ifu with ore concentrators, the dilute slurry effluent goes directly to
a Kigoon of roughly four acre-ft capacity adjacent to the bay or river. An
unknown degree of settling and precipitation would occur in such a lagoon.
While this practice is presently in existence, it is believed that existing
ur soon-to-exist effluent control regulations require lime neutralization
and impoundment of the resulting sludge in such cases.
In the case of discharge to the concentrator tailings pond,
the soluis buildup attributable to the above-described dilute slurry is
insignificant relative to that due to ore tailings. These solids, how-
ever, contain potentially hazardous heavy metals.
(b) Other Residues. As indicated in Figure 2, all other
significant residues associated with electrolytic copper refining are
immediately recycled or shipped elsewhere for recovery of metal values.
The arsenical sludge resulting from electrolyte purification is stored
in concrete bins and shipped elsewhere for processing. Slimes recovered
from the electrolytic cells are of extremely high value, and the slag
from the melting furnaces is shipped to a copper smelter for reprocessing.
14
-------
Analyses of samples collected from copper smelters and refineries
are given in Appendix A. Not all copper smelting and refining wastes
arc considered as having equal degrees of potential hazard. Copper
smelter slags are large chunks of hard siliceous furnace residue. In
solubility tests described in Appendix B it was found that these slags
did not leach significantly. Therefore, copper slags (reverbatory and
electric furnace) are considered non-hazardous in view of data available
at this time.
In solubility tests it was found that sludges (acid plant
sludge) and dusts (converter dust, reverbatory dust) leached copper,
lead, zinc, and cadmium in significant concentrations (see Appendix B).
Leaching of dusts and sludges would be anticipated in view of the fine
particle size (i.e., large exposed surface area). These wastes are
therefore all considered potentially hazardous.
1.2.3 Waste Quantities
Table 5 summarizes the generation factors for each of the land
disposed or stored residuals in the primary copper smelting and fire
refining category. This table also gives typical concentrations of
potentially hazardous constituents. Table 4 gives the yearly quantity
of wastes and potentially hazardous constituents for a typical plant
producing 100,000 MT (110,000 short tons) metric tons of blister copper
(i.e. from smelter) and 100,000 MT (110,000 short tons) tons of fire
refined copper per year.
Table 5 summarizes the generation factors for each of the
land disposed or stored residuals in the primary copper electrolytic
refining category. Table 6 gives the yearly quantity of wastes and
potentially hazardous constituents for a typical plant producing 160,000
MT (176,000 short tons) of electrolytic copper per year.
Using the waste generation factors and concentration factors
contained in Tables 3 and 5 and copper smelter and refinery production
capacities, estimates have been made for the quantities of land disposed
or stored waste and hazardous constituents thereof. These estimates are
given for 1974, 1977, and 1983 on a state-by-state, regional and national
level in Tables 7 through 8. It is seen in these tables that rever-
batory slag is generated in huge amounts compared to sludges and dusts.
However, as previously discussed, this waste is not considered poten-
tial ly hazardous.
Sludges from smelters and fire refiners are generated at a
much higher rate than those from electrolytic refiners mainly as a
result of the need to scrub SO- from emissions of smelters. Electrolytic
refiners produce relatively little sludge. Increases in the quantities
of sludges from milling and fire refining and electrolytic refining will
increase moderately over the period 1974 to 1983 paralleling predicted
increases in production capacity.
15
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Dusts from smelting and fire refining are generated at a much
lower annual rate than sludges. In contrast to sludges-,;dusts are event-
ually recycled-,albeit after storage on the ground for periods of months.
There are no dusts generated in electrolytic refining.
Arizona generates the largest amount of land disposed and
hazardous wastes from smelting and fire refining by virtue of the fact
that it has the most (7) smelters and fire refiners. On the other hand
New Jersey and Maryland generates the largest quantity of electrolytic
sludges because of the large production capacities of these states for
electrolytic copper.
Projections of waste quantities for 1977 were based on addi-
tional electrolytic plant capacity now under construction in Arizona (27
MT/yr) and Texas (380 MT/yr) as given in the January 1975 Engineering
and Mining Journal. By 1983 it appears that plant effluents from
primary copper smelters and refiners will be limed, thereby increasing
the quantities of sludge per metric ton of product by an estimated, 26%
for smelters and 3% for electrolytic refiners. Thus, sludge generation
factors were increased from 19.7 kg and 2.4 kg/MT, respectively.
An additional planned electrolytic refinery capacity of 36 MT/yr in
Arizona and a 91 MT/yr smelter capacity increase in New Mexico was also
considered in 1983 projections.
1.3 TREATMENT AND DISPOSAL TECHNOLOGY
This section contains descriptions of methods presently used
for treatment and disposal of waste residuals on land and technology
which is considered for adequate health and environmental protection.
1.3.1 Current Waste Treatment And Disposal Practices
Smelting and Fire Refining. Currently reverbatory h,lags are
disposed in open dumps. This is adequate disposal since leacning is
not evident.
Acid plant sludges are stored on land before recycle. Over-
flow from acid plant hlowdown and miscellaneous slurries from scrubbers,
cooling of anodes, washing etc. are settled in lagoons or tailings
ponds. Disposal of sludges and slurries as currently done is inadequate
due 10 the danger of toxic metal leaching.
Electrolytic Refining. Slurries from electrolytic copper
refining are currently settled in unlined lagoons or tailings ponds
and may be dredged and disposed of on land. These practices are in-
adequate due to the danger of toxic metal leaching.
28
-------
The following paragraphs describe treatment and disposal
technology levels for those wastes which have been ascertained as
potentially hazardous.
1.3.2 Present Treatment and Disposal Technology (Level I)
Smelting and Fire Refining . Acid plant blowdown results in
sludge which is dredged from settling lagoons and stored on land before
eventual recycle to the reverbatory furnace. Overflow from acid plant
settling lagoons is sent to mill tailings ponds. Miscellaneous slurries
from scrubbers, contact cooling of anodes, washing, etc. are settled in
laj'.oons or tailings ponds. They arc either dredged, stored on land and
i:v<-nt u.-i 1 ly recycled or left permanently in tailings ponds. Dusts are
stored on land for periods of months before recycling for reclamation of
metallic values. The relative amounts of stored and disposed wastes
vary from location to location and in response to market fluctuations.
Electrolytic Refining. At the present time miscellaneous
slurries from electrolytic copper refining including spent acid, cooling
water blowdown, scrubber water and washings are settled in an unlined
lagoon or tailings pond. Settled sludges may either be dredged and
disposed on land, or left permanently in tailings ponds.
1.3.3 Best Technology Currently Employed (Level II)
Smelting and Fire Refining . Lined lagoons are not used
presently by the industry and thus Level II technology is unlined
lagoons for acid plant blowdown and miscellaneous slurries. Land areas
for storage or permanent disposal of dredged sludges (i.e. acid plant,
mi seellaneousj are not sealed. Therefore Level II technology is storage
or permanent disposal on unsealed ground areas.
Electrolytic Refining. Lined lagoons are not used at present
by the industry and thus Level II technology is unlined lagoons. Land
areas for storage or permanent disposal of dredged sludges are not
presently treated. Therefore Level II technology is storage or permanent
disposal on untreated ground areas.
1•3•4 Technology To Provide Adequate Health And Environmental
Protection (Level III)
Smelting and Fine Refining. Solubility tests as previously
described indicate that hazardous constituents such as lead, chromium,
antimony, cadmium, and copper are significantly leached from sludges
which are stored in lagoons (i.e. acid plant sludges, scrubber sludges,
slurries, etc ) Lagoons should therefore be lined to prevent leaching
and subsequent percolation through permeable soils to underlying water
tables.
29
-------
instead of storing dredged sludge on land prior to recycling
they should ho stored in concrete lined pits to eliminate leaching,
Altern/JtivoJy, immediate recycle of sludge without storage- may be
feasible. Immediate recycle of dusts without ground storage precludes
possible leaching and would be Level III technology.
Electrolytic Refining. Significant leaching of hazardous
constituents (e.g. lead, copper, chromium) from electrolytic refining
lagoon sediments is possible. The use of lined lagoons for permanent
storage ot' such sediments should be practiced in permeable soils with
underlying aquifers.
Tables 9 and 10 summarizus features of levels 1, IT, and III
treatment and disposal technology for smelting and refining including
assessment of the hazardous nature of residuals and adequacy of treatment
and disposal.
1.4 COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental
protection were described. The costs of implementing this technology
for typical plants is considered in this section.
1.4.1 Cost of Present Treatment And Disposal Technology (Level I)
The costs associated with the usual treatment and disposal of
residuals from typical smelting and refining plants are presented below.
Smelting and Fire Refining. The plant has an annual capa-
city of 100,000 MT operating 350 days. Production of blister copper
and fire refined copper results in a number of different potentially
hazardous waste streams, namely wet sludges and slurries, >i.:J dry dusts.
Slurry from lime treatment of SO,, gases from the acid plant
amounts to 2,270 m /day and is processed through a thickener. The
overflow from the thickener, 2200 m /day, enters a lagoon (Lagoon A).
The daily inflow contains 0.77 MT,of solids which form a sludge with
an estimated density of 1200 kg/m at the^bottom of the lagoon. This
results in a daily accumulation of 0.65 m of sludge or 227.5 m /year.
The lagoon size provides for 4 days of residency.
Miscellaneous slurries arising from sources such as wet
.scrubbers and anode cooling are deposited into a separate lagoon,
designated as Lagoon B. One large lagoon could be used if desired.
Its estimated daily inflow is 13,715 m containing 4.8 NfT of sus-
pended solids. The sludge formed is assumed to have %h& same density
as that in Lagoon A resulting in the formation of 4 m of sludge/day
and 1,400 m /yr. The lagoon size is based on 4 days of inflow.
30
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Charai t eri st ics ol both lagoons necessary ro provide retention tor
adi'(|ii;it i* settling are ^tvi-n Drlow.
Lagoon A Lagoon B
Volume . 8,80u m 55,000 m
tiottom width 34 m 91 m
Top width 46 m 106 m
Bottom length 6/ m 182 m
lop length 79 m 194 m
Total depth 3m 3m
Depth of excavation 1.2m . 6b m
Circumference 262 m 607 m
Dike Volume 3,117 m 10,987 nu
Dike Surface .5,555 m 8,y87 m
Total Width 59 m 119 m
Total length 93 m 210 m
Required area .55 ha 2.5 ha
Two test holes are drilled tor Lagoon A, four tor Lagoon b.
Both are located on rural land. The overflow from both lagoons go to
.1 t ;i i I j UK'-' pond which is part of the mine-mil 1-smcl tor complex. No
•. ixni licant portion of tin- tailings pond cost as attributable to Smelter
waste disposal because of insignificant volumes of smelter water wastes
compared to mill tailing waste water (i.e. water from ore beneficiation).
Both lagoons are dredged twice/year requiring about 65 hours of
dragline time. The dredged material is piled next to tne lagoons,
allowed to dry and t lien recycled. No provisions exist to prevent run-off
or teaching of the dredged material for Level I.
Large amounts of dust are collected from each ot the primary
process operations. About 150 MT arc stored on land near the smelter at
any given time for subsequent recycling. No measures to prevent leaching
are taken under Level 1.
Cu'.l of Level I I ri-atment and disposal technology are given in
I .ihle 11.
ll_l ec t ro lyti c He.fi" i"g • 'be plant capacity is 160,OOU MT
operating 350 ciays/yr. The only land disposed waste is a combined
dilute slurry composed of effluents from operations such as contact cooling
of copper cast anodes, spent anode washing and plant washdown.
3
The daily flow is 0,800 m which contains /10 kg of suspended
solids. This flow enters a lagoon designed tor 4 days settling time.
Characteristics of the lagoon are:
35
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TABLE 11
CUST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRlMAKY CuPPfcR-SMELTlNG AND FiRE KEFiNlNG
CAPITAL COST
Sludge
Lagoon A
Site Preparation
Survey $ 345
Test Urilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation 5 Forming 4,145
Compacting 5,765
Fine Grading 1,600
Soil Poisoning 325
iransverse Drain Fields 545
Land 965
Lagoon B
Site Preparation
Survey 1,565
Test Drilling 980
Sample Testing 500
Report Preparation 1,500
Construction
Excavation § Forming 14,615
Compacting 20,625
Fine Grading 4,045
Soil Poisoning 755
Transverse Drain Fields 3,285
Land 4, .575
Equipment
Dragline (S°0)
lotal
36
-------
TABLE 11 (cont.)
Co.ST ()(•' l.f.VI-l. ' TKl-ATMliN!' AND (MSl'OSAl. Tl',CHNUl,0(,Y
I'KlMAKY COITCK SMIiLilNCi AN1> I'lRU Wtif;INlNt;
ANNUAL COST
Land $ 535
Amortization
Construction 7,22u
Equipment 555
Operating Personnel 790
Repair and Maintenance
Construction 1,865
r-qujpment 175
l-.nerj;y
1'iiel 75
I-. lec.tr if ity 10
Taxes 135
insurance 7iO
Total $12,070
37
-------
Volume 28.UOO m
Bottom width 64 m
Top width 76 m
Bottom length 128 m
Top length 14U m
Total depth 3 m
Depth of excavation . 8b m
Circumference 443 m_
Dike volume 6,948 m_
Dike surface 6,335 m
Total width 90 m
Total length li>5 m
Required area 1.4 ha
The lagoon is located on rural land.
The density of the sludge formed in the lagoon is 134U kg/m .
Only 5b m of sludge is produced annually obviating dredging.
Costs ot Level 1 treatment and disposal technology are given
in Table 12.
1.4.2 Cost of Best Technology Currently Employed (Level IIJ
Smelting and Fire Refining. Collected dusts are recycled
immediately without ground storage. This does not affect Level I costs.
ihe technology and costs for sludge disposal are the same as Level I.
Electrolytic Refining. The technology and costs are the same
as Level I.
1.4.3 Cost of technology To Provide Adequate Health and hnvi^'onmentai
Protection (.Level 1IIJ
Smelting and Fire Refining. Jhe lagoons are lined and the
lagoon sludges are pumped to concrete pits for temporary storage prior
to being recycled. The pits are sized to hold sludge formed over a 6
month period. Pumping the sludge eliminates the need for the dragline
used in Level 1 tor dredging.
The pits for Lagoon A and Lagoon B sludges are 5 x 5 x 8 m and
3 x 10 x 24 m, respectively. The same slurry pump is used to dredge the
sludge from both lagoons. The slurry pump is required to operate 115 hrs/yr
and 150 labor hours are allocated to this task. Incremental costs of
Level III technology are listed in Table 13.
38
-------
TABLE 12
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY COPPER - ELECTROLYTIC REFINING
CAPITAL COST
Sludge
Lagoon
Site Preparation
Survey $ 875
Test Drilling 980
Sample Testing 500
Report Preparation 1,500
Construction
Excavation f» Forming 9,240
Compacting 12,855
Fine fir.vling 2,850
Soil Poisoning 550
Transverse Drain Fields 1,655
Land 2,450
Total $55,455
ANNUAL COST
Land $ 245
Amortization
Construction 3,600
Equipment
Operating Personnel
Repair and Maintenance
Construction 815
Energy
Fuel
Electricity 30
Taxes 60
Insurance 335
Total it 5,085
39
-------
TABLE 13
COST OF LEVEL III TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY COPPER - SMELTING AND FIRE REFINING
CAPITAL COST
Construction Sludge
Lagoon A liner $18,230
- Lagoon B liner 88,565
Concrete Pit (Lagoon A) 8,380
Concrete Pit (Lagoon B) 24,535
Equipment
Slurry Pump 13,730
Pipe, rigid (installed) 4,425
Pipe, flexible 880
(Dragline) (3,500)
Total $155,145
ANNUAL COST
Land
Amortization
Construction $ 16,195
Equipment* 2,470
Operating Personnel* 630
Repair and maintenance
Construction 4,190
Equipment 775
Energy
FuH
Electricity* (70)
Taxes
Insurance 1,550
Total $25,740
* Costs shown are net costs, i.e. cost of new equipment less dragline
associated costs which are no longer incurred.
40
-------
E1ectro1 y tic Refining
Level III consists of lining the lagoon. The incremental
costs are:
Capital cost
Lagoon liner $ 48,880
Total 48,880
Annual cost
Construction Amortization $ 5,675
Construction repair and
maintenance 1,465
Insurance 490
Total $ 7,630
Costs of Levels I, II and III treatment and disposal technology
are summarized in Table 14 for scrubber sludge and dust from smelting
and fire refining and in Table 15 for electrolytic plant sludges. Costs
are given for individual waste streams (i.e. sludges, dusts) and for
total waste streams in dollars per metric ton of dry and wet waste and
copper product. The costs of treating each waste stream are also expressed
as percentages of metal selling price.
Total industry costs for Levels I, II and III treatment and
disposal technology have been estimated in 1973 dollars based on 1973
production and are presented in Table 14 for smelters and fire refiners
and Table 15 for electrolytic refiners. The annualized cost for the
primary copper smelting and fire refining industry is estimated as
$190,000 for Levels I and II treatment and disposal technology and
$6,130,000 for Level III technology (i.e. technology adequate for environmental
protection). Levels I and II technology represents approximately 0.009%
of 1973 national sales. The cost of Level III technology represents
0.28% of national sales of primary copper.
The annualized cost for the primary copper electrolytic refining
industry is estimated as $50,000 for Levels I and II treatment and
disposal technology. This represents approximately 0.002% of 1973
national sales. The annualized cost for Level III treatment and disposal
technology is estimated as $120,000 or approximately 0.005% of national
sales.
41
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2.0 PRIMARY SMELTING AND REFINING OF LEAD
2.1 INDUSTRY CHARACTERIZATION
There are only 7 primary lead smelting and refining plants in
the United States owned by four different corporations. These are large
operations however ranging in size from 56,000 MT to 209,000 MT per
year. Two plants conduct smelting operations only; one plant conducts
refining operations only and the remaining four plants smelt and refine
lead ore concentrates. In 1973 total U.S. production of refined lead
was 687,739 short tons and average selling price was $0.1629/lb (Ref. 2).
The total U.S. sales are estimated as $224,065,000.
Features of individual lead smelters and refiners including
capacities, presence of acid and slag fuming plants and gas cleaning
methods are summarized in Table 16. All of these features have an
influence on the types and quantities of residuals disposed or stored on
land. Table 17 gives state by state, EPA region and national primary
lead production capacities.
The ore from which primary lead is produced contains both lead
and zinc. The lead concentrate is usually roasted in traveling-grate
sintering machines, thereby removing sulfur and forming lead oxide. The
load oxide, sinter, coke, and flux (usually limestone) are fed to the
blast furnace, in which oxide is reduced to metallic lead. The lead is
further refined in reverbatory furnaces. By product recovery from lead
refining residues include copper, zinc, antimony, arsenic, silver, gold,
bismuth, cadmium, selenium, tellurium, nickel and cobalt. Process
details are further described in the next section.
2.2 WASTE CHARACTERIZATION
This section contains descriptions of production rcciinology at
primary lead smelters and the resultant byproducts or wasces which are
either recycled directly, shipped to other smelters for further metallic
recovery or disposed of on land. Estimates are given for the quantities
of wastes and potentially hazardous constituents thereof which are
disposed of on land either in lagoons or open dumps.
2.2.1 Process Description
The hypothetical lead plant is assumed to have an annual capa-
city of 110,000 metric tons (MT) of 99.99 + % lead per year (121,000
short tons/year). Five of the six U.S. smelters have within +15 percent
of this capacity and the sixth has approximately double this capacity.
Lead smelters are estimated to operate 350 days/year; the daily capacity
is, therefore, 314 MT (346 short tons).
44
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Input (feed) to the plant amounts to approximately 163,000
MT/year (180,000 short tons/year) of flotation concentrates from sulfide
ores (and miscellaneous lead-bearing materials) averaging approximately
75 percent lead. In addition to concentrates, the feed may include
lead-bearing dusts and fumes from gas cleaning operations, sinters,
fines, slags, drosses, and residues generated in the lead smelter, and
residues such as flue dusts shipped from copper and zinc smelters. This
conglomeration of feed materials may contain such recoverable byproduct
metals as zinc, silver, arsenic, antimony, bismuth, cadmium, copper,
gold, platinum, palladium, nickel, colbalt, selenium, tellurium, and
germanium.
A representative flow diagram is shown in Figure 3. This
diagram shows the quantities of wastes or recyclable material either
stored on land or permanently disposed in lagoons or open dumps. The
principal unit smelting operations are (1) double-pass sintering to
remove sulfur and simultaneously agglomerate fine-sized feed materials:
(2) blast furnace smelting of sinter with coke and fluxes to make a lead
bullion containing most of the byproduct metal values, and a slag contain-
ing most of the zinc and worthless gangue material; (3) slag fuming to
recover byproduct zinc oxide, a lead fume for recycling, and a discardable
waste slag; (4) dressing the molten impure bullion with air and sulfur
to remove most of the copper; (5) reverberatory smelting of the copper
dross to produce an enriched copper matte for shipment to a copper
smelter, and a crude lead bullion for recycle to the dressing operation.
Auxiliary smelter operations include (1) collecting dusts from the
sintering, blast, and reverberatory furnace operations for reprocessing;
(2) precipitating fumes from the same gas streams for diverse treatment
to recovery their lead, zinc, cadmium, antimony, and arsenic contents;
and (3) recovering sulfur dioxide from the cleaned sinter plant gases in
a sulfuric acid plant.
Blast furnace slag is the primary solid waste produced at the
smelter in large quantity. Smaller amounts of particulate sludges
emanate from the acid plant and from other sources as will be detailed
below.
After dressing (i.e. floating on surface of melt) to remove
most of the copper, the crude lead bullion is refined as shown in Figure
3-a to produce marketable lead, silver, gold, bismuth, and reprocessible
byproduct slags, fumes, solutions, and residues containing zinc, cadmium,
antimony, arsenic, selenium, tellurium, copper, tin, zinc, nickel, and
cobalt.
The main steps in the pyrometallurgical refinjng of lead
bullion comprise (1) softening, using air to remove antimony, residual
copper, arsenic, and tin; (2) desilverizing the softened lead bullion
with metallic zinc; (3) dezincing and casting of pure lead. The oxidized
dross, slag, and fume from the softening furnace are subsequently smelted
47
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HOT METAL FROM
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FURNACE
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METALLIC ZINC
(3) ZINC CRUSTS
TO ZINC, RECOVERY
PRECIOUS METAL
RECOVERY
DROSS, SLAG, DUST
•PROCESSED FOR Sb.
Sn, RECOVERY
LEAD BULLION
DESILVERIZING
FURNACE (2)
REFINED LEAD
Figure 3 PYROMETALLURGICAL LEAD REFINING
49
-------
'.;> produce marketable antimonial lead, t«nvin ,. &;^
values. An estimated 95 percent of the principal mecal v^l^t,.^ ., • t-ui
silver, are recovered either in marketable liigh-quu.L;.r.% ;,<.<.• \ a )..- o:
recoverable byproducts.
Gases from the first section of the sinter r.ic.chii'e frr-a -. = , >
enough in SQ to permit sulfuric &cid reccvery b.y convent i one J ,' i< .,•••.',. -.o
The amount or acid product is 86,000 MT (95,000 short
245 MT (270 short tonrO per operating day.
2.2.2 Uesc r ip i: i on i _o_f Wab t c ^S t r earns
a< Blast F u r na c e
Molten siliceous s.ag from the olasu fuvi\ui.e
to sinter, but 30% is passed thru a sl£g fumir?,?, r'urr.a
and some residual lead, then granulated by Wh.te: qae'.i
granulated slag leaving tne sr.it it at1 and jjring <;u the iii^n ^U;;_j L
300 Kg/NCT of lead product or 9.3 M7/day 'Iu3 yiiCic conWcKy; . Sc:
do not fume slag; untreated yraim la ted s,i,.g th-,^ g'.es a:,rt"-;;y ,
b. Slag Granu1 ation Slurry
Slag granulating water amounts to approximateiy
meter/day (97,000 gal/day). This water flows to 'wo c.;. ;
coarse solids,settle out in the short residence t.imc al • ....
holding 150 m'"* (39,000 gal), are dr-edgsa period:'.-^il> = / , ' ., -\
fine slag is placed permanently on the ;:uJ.A sir., ..'!/oay '40 -;.rr;
Overflow from tne setcliaj pit ^.crr^.i^ cnvp^-jd^v] HI
particulates goes to a larger, longer r-i-irience ti.R»i y.-^r, •."
from which water is recycled to slag gra/ulat».on, T'iie. a-ro.xr.i: .. i
very fine settlings is estimated at 30 Kg/MT of lead proj. • i 01
(11 short tons/day) and the size of vhe polishing lagoo i is, appr
6000 cubic meter (i,600,000 gal).
This lagoon is dredged periodically and Jia dreugr.ngs
disposed of on the slag dump,
50
-------
c. Slurries from Primary Gas-Cleaning Operations
Sinter gases laden with SO- are partially scrubbed of part-
iculates before entering the acid plant. Gases from other primary
operations such as sinter crushing and blast furnacing are similarly
cloaned in electrostatic precipitators, wet scrubbers, and baghouses.
The bulk of the dry dusts and wet slurries resulting from primary gas
cleaning operations are recycled to sinter but a portion of the slurry
containing perhaps 6.1 MT/day (6.7 short tons/day) of solids (in an un-
known amount of water) is not immediately recycled. Gas cleaning practice
is believed to vary widely from plant to plant-
Handling of the slurry is by settling in an unlined pit of
capacity 260 cubic meters (68,000 gal). The pit is dredged periodically
and the sludge is piled on the slag dump where it resides for several
months before recycle to sinter. This sludge is produced at a rate of
l!> Kg/Ml' of lead product. It is believed that this material may not be
recycled at some plants, i.e., it is then definitely a solid waste disposed
of on land.
^• A.c*!fL Pl_a.Pt Slowdown Slurry
Necessary acid plant blowdown and lime treatment of the result-
ing effluent produces approximately 230 cubic meter/day (60,000 gal/day)
slurry effluent containing 6.3 MT/day (6.9 short tons/day) solids. This
slurry goes to a lime treatment tank and then to the acid plant lagoon
(unlined) of capacity approximately 6,800 cubic meters (240,000 cubic feet).
Miscellaneous slurries (see item e. below) join the limed acid plant blow-
down in this lagoon. The lagoon is dredged periodically and the sludge is
piled on the ground to dry before eventual recycle to sinter. The amount
of dry dredgings attributable to the acid plant blowdown is as estimated
above, i.e., 20 Kg/MT of lead product or 6.3 MT/day (6.9 short tons/day).
This material may not be recycled at some plants.
c. Miscellaneous Slurries
Miscellaneous dusts such as sweepings are slurried and combined
with wastewaters such as cleanup water and cadmium plant effluent; this
effluent is lime treated and settled in the above-mentioned acid plant
lagoon. The quantity of miscellaneous slurries entering the lagoon is
estimated to be less than 230 cubic meter/day (60,000 gal/day) and the
contained solids are estimated to be less than 20 Kg/MT of lead product
or 6.3 MT/day (6.9 short tons/day). The settled sludge is dredged and
dried on the ground in mixture with the above-described acid plant sludge.
Total dry sludge per day from acid plant blowdown and miscellaneous slurries
is 40 Kg/MT lead product.
51
-------
f . Dry Dusts From All Gas Cleaning Operations
A portion of the dusts collected from sintering, blast furnacing,
and other operations In haghouses and other dry chist collectors is not
slurricd but is recycled in dry condition to the sinter machine. It is
estimated that 19,000 MT/year (20,906 f-hort tons/year) are handled in
this manner from the typical plant. Recycle is immediate and no solid
waste disposal results. Approximately 2.000 MT or 10% of dusts are land
stored before recycle.
Chemical analyses of waste So^iples collected from v<'.rj.ous
primary lead smelters are given in Appendix A. "n soiubility tests
described in Appendix B it was found that dredged siudge originating
from the sinter dust scrubbers and acid plant leached significant concentrations
of cadmium, copper, lead, and zinc, Thic/ may be expected from the high
concentrations of heavy metals in these wastes and fine parti culate
sizes. Because of evidence indicating significant leaching of heavy
metals from sinter plant, acid plant and miscellaneous sludges, these
wastes are considered potentially hazardous,
The granulation of lead fuming slag comminutes the slag to
sand and pebble size separates, Solubiliza'cion of copper, antimony and
zinc in the order of 1 ppm was observer in leaching scudies (see Appendix
B) . This degree of solubi lizatiou was not considered sufficient to
identify the lead fuming slag as a potentially Uizarcious waste at this
t irae.
2.2.3 Waste Quantities
Table 18 summarizes the generation factors for each of the
land disposed or stored residuals in the lead smelting ar,d refining
category. This table also gives typical concentrat ions of potentially
hazardous const i tiu-nts . Table 19 gives the yearly uuantir, ..duals
and potentially hazardous constituents for a typical plart !>->,, .ucing
11 0,000 MT (l.'l, 000 short tons) of 99.99% price lead pe. L,,. .
Using the waste generation factors and concon r;U;iori ractors
contained in Table 18 and lead smelter and iefi> lag p,..ant capacities,
estimates have beer, made for the quantities of '/I'd -disposed or s.r'cd
waste and hazardous constituents thereof. These estimates ars given for
1974, 1977, and 1983 on a state-by-state, regin iai and national it-vej. in
Table 20.
The state in which the largest quant Jry of potentially hazardous
waste (i.e. sludge) is generated is Missouri wnieh is the largest producer
of primary lead (3 of the 7 primary lead smelters are in Missouri) .
Idaho and Texas also generate significant quantities of potentially
hazardous waste from lead smelting.
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There is no documented, planned plant expansion in the industry
or anticipated changes in production or pollution control technology
which would affect the quantities of land disposed wastes. Therefore,
estimates for 1977 and 1983 are the same as those for 1974.
2.3 TREATMENT AND DISPOSAL TECHNOLOGY
2.3.1 Current Waste Treatment and Disposal Practices
Currently^ discard slag is open dumped. Acid plant and sinter
plant sludges are settled in unlined lagoons, dredged and either stored
or disposed of permanently in unsealed ground areas. As discussed pre-
viously, slag is not considered hazardous at this time. Sludge is sus-
ceptible to heavy metal leaching and levels of technology for its treatment
and disposition are discussed in the following sections.
2.3.2 Present Treatment and Disposal Technology (Level I)
At the present time lagoons are unlined. Dredged acid plant
and sinter plant sludge are either permanently disposed of on unsealed
ground areas or stored on unsealed ground areas. Potentially hazardous
metals including lead, cadmium, copper, and zinc may be leached from
sludges through porous soils to underlying aquifers. Present practices
need augmentation.
2.3.3 Best Technology Currently Employed (Level II)
At one or more primary lead smelters lime sludge from the acid
plant and scrubber water sludge from the sinter plant arc immediately
recycled thereby precluding land storage or disposal. This practice is
considered Level II technology.
2.5.4 Technology To Provide Adequate Health and Environmental
Protection (Level III)
The practice of immediate recycle of acid plant and sinter
scrubwater sludges would qualify as Level III technology. In the event
that sludges cannot be immediately recycled,concrete pits and lined
lagoons will bo required as Level III technology to prevent leaching
of metals including cadmium, copper, lead and :,inc through permeable
soils into aquifers. Sludges originating from the sinter plant or the
acid plant which are permanently disposed of on land should be chemically
fixed (i.e. solidified and immobilized) before ground disposal as Level
111 technology to prevent leaching through permeable soils to aquifers.
56
-------
Table -1 summarises features of Levels I, II, and 111 treat-
ment disposal technology including assessment of the hazardous nature of
residuals and the methods and adequacy of treatment and disposal methods
for sludges.
.2.4 COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental
protection were described. The costs of implementing this technology
for typical plants is considered in this section. The hypothetical
plant is assumed to have an annual capacity of 110,000 MT of lead product
per year and operate 350 days per year.
2.4.1 Cost of Present Treatment and Disposal Technology (Leve1 I)
Slag granulation water, slurries from primary gas cleaning,
acid hlowdown and miscellaneous activities are directed into a lagoon.
Daily inflow is 1.4KO m* containing 23.8 MI' of solids. The postulated
lagoon sir.e is (>,00() in". Characteristics of the lagoon to accommodate
the daily volume of wastcwatcr and approximately 200 days of sludge
accumulation are:
Bottom width 36 m
Top width 44 m
Bottom length 72 m
Top length 80 m
Total depth 2 m
Depth of excavation .7 m
Ci rcumforonce ,"!S9 m..
Dike vo lunii1 1 , iSS.S m',
hike Mirlarr J,(>71 m"
Total width S.S m
Total length 91 in
Required area .5 ha
Two test holes and 4 analyses of soil samples to ascertain lagoon
structural stability are included in site preparation. The lagoon is
situated on rural land.
, The daily inflow of 23.8 MT of solids are estimated to form
14.3 m' of sludge in the lagoon. The lagoon is dredged when one-half
filled with sludge. This occurs every 209 days or 1.7 times/operating
year. The sludge, amounting to 5,100 m"/yr, is allowed to dry next to
the lagoon and hauled to an on-site slag dump. Dredging, loading and
hauling require about !'.)() hrs/yr of dragline and 190 hrs/yr of front
loader and truck time.
57
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The slurry from gas cleaning operations is passed through an
unlined pit with a capacity of 260 m (2 x 5.2 x 15.6). The slurry
contains about 6.1 MT/day of solids, 4.1 MT (67%) of which are estimated
to settle in the pit. (The solids form a sludge with an estimated
density of 1650 Kg/m , and about 3 m of sludge accumulate daily).
Dredging, loading and hauling of sludge from the slag granula-
tion pits is estimated to require an average of 2 hr/day of backhoe and
frontloader and 1 hr/day of truck time. Comparable requirements for the
gas cleaning slurry pit are 60 hr/yr of backhoe and front loader time
and 30 hr/yr of truck time. The costs of Level I sludge disposal are
given in Table 22.
2.4.2 Cost of Best Technology Currently Employed (Level II)
The only change from Level I is the immediate recycling of
wastes dredged from the gas cleaning slurry pit. Costs remain as in
Level I.
2.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III)
The lagoon is lined; dredgings from the lagoon are chemically
fixed prior to being hauled to the slag dump.
Additional costs are as follows:
Capital Costs
I.agoon liner
Total
Annual Costs
Construction Amortization $ 2,040
Construction Maintenance
and Repair 525
Chem-Fix of sludge 67,320
Insurance 175
Total $ 70,0(>0
60
-------
:_ .:esT OF LEVEL i TREATMENT \xo DISPOSAL
rS- IMA*:' LEAI FLV.T
CAPITAL COST
Sludge
Lagoon
Site Preparation
Survey $ 315
Test Drilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation £, Forming 2,510
Compacting 3,495
Fine Grading 1,200
Soil Poisoning 320
Transverse Drain Fields 505
Land 875
Sump Construction
Slag Granulation Pits (2) 400
Primary Gas Cleaning 345
L-quipment
Dragline (10",) 7,000
Dump Truck (70%) 7,875
Front Loader § Backhoe (90%) 10,440
Bulldozer 1,120
Dump
Survey 1,335
Land 3,735
To t;il $43,410
ANNUAL COST
Land $ 460
Amortization
Construction 1,435
Equipment 4,210
Operating Personnel 20,695
Repair and Maintenance
Construction 265
Equipment 1,320
Energy
Fuel 2,325
hlee trieity 230
T.-uos 115
435
Total $M .-UK)
61
-------
Costs of Levels I, II and III treatment and disposal technologies
are summarized in Table 23 for sludge residuals from primary lead smelting
and refining. Costs are given in dollars per metric ton of dry and wet
waste and dollars per metric tons of lead product. Costs for each level
of technology are also expressed as a percent of metal selling price.
Total industry costs for Levels I, II and III treatment and
disposal technologies have been estimated in 1973 dollars and are presented
in Table 23. The annualized industry costs for Level I and II treatment
and disposal technologies is estimated as $180,000 which is 0.08% of
estimated national sales. The annualized industry cost for Level III
treatment and disposal technology (i.e. adequate for environmental pro-
tection) is estimated as $560,000 which is 0.25% of estimated 1973
national sales.
62
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3.0 PRIMARY SMELTING AND REFINING OF ZINC (SIC 3333)
3.1 INDUSTRY CHARACTERIZATION
There are two major processes employed for recovery of zinc
metal from ore concentrates. They are electrolytic recovery and distillation
(pyrometallurgical). Of the seven operating primary zinc plants in the
United States three use electrolytic processes while four employ various
modes of pyrometallurgical distillation (vertical retort, horizontal
retort, electrothermic). Table 24 gives the location of primary zinc
plants, types of operation, and zinc producing capacities. This table
also indicates the presence or absence of byproduct acid plants and
identifies major products produced at the plants. Table 25 gives state
by state, HPA region and national primary lead production capacity by
process type. Total U.S. production of refined primary zinc in 1973 was
628,785 short tons and the average selling price was 20.66 cents/lb
(Ref. 2). National sales are therefore estimated as $259,814,000 in
1973.
The reduction of zinc ores and concentrates to zinc is accomplished
by electrolytic deposition from a solution or by distillation in retorts
or furnaces. For either method the zinc concentrate is roasted to elim-
inate most of the sulfur and for conversion to impure zinc oxide called
roasted concentrates or "calcines".
At electrolytic zinc plants, the roasted zinc concentrate is
leached with dilute sulfuric acid to form a zinc sulfate solution. The
pregnant solution is then purified and piped to electrolytic cells,
where the zinc is electrodeposited on aluminum cathodes. The cathodes
are lifted from the tanks and are stripped of the zinc, which is then
melted in a furnace and cast into slab form. Zinc concentrates shipped
to electrolytic plants commonly contain lead, gold, and silver. The
zinc acid leaching and electrolytic tank residues become enriched in
these metals by the extraction of the zinc and are usually shipped to a
lead smelter. There the lead, gold, and silver content is recovered in
lead bullion.
Distillation retort plants are classified as horizontal retorts,
vertical retorts that are externally heated by fuel, or continuous
vertical retorts heated electrothermally. All employ coal or coke as
the reducing agent. Quantities of coal or coke required range from
about 0.5 to 0.8 ton per ton of slab zinc output. The zinc vapor and
carbon monoxide from the retorts pass into condensers where the zinc is
collected us liquid metal ready for casting into slab form. Zinc produced
by distillation, normally the lower commercial grade, may be upgraded by
refining to reduce the quantities of impurities. Refining, by redistillation
is accomplished by vertical fractionating columns which separate the
impurities contained in the feed zinc and can produce zinc of 99.995
plus purity.
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WASTE aiAKAm.KlZAHON
This section contains descriptions of production technology at
primary zinc smelters and the resultant byproducts or wastes which are
either recycled directly, shipped to other smelters for further metallic
recovery, or disposed of on land. Estimates are given for the quantities
of wastes and potentially hazardous constituents thereof which are dis-
posed of on land either in lagoons or open dumps.
3. 2. 1 Process rjgscripti on
Because of major differences in production processes and waste
generation between electrolytic and pyrometallurgical (distillation) zinc
plants, the two types of plants wilJ be discussed separately.
ijlectrol^tic. The hypothetical electrolytic zinc plant is
assumed to have an annual capacity of 100,000 MT (110,000 short tons) per
year. Two of the three operating electrolytic zinc plants in the U.S.
are within ten percent of this size and the third is very nearly three-
fourths of this size. .Electrolytic zinc plants are assumed to operate
350 days/year; the daily capacity is, therefore, 286 MT (314 short
tons/day).
I-. 1 ectrolyt ic reduction nl ,-.inc from sulfide ore concentrates
bruin-, with rousting (as duo.-, pynnnrt .t 1 lurgieul ~inc smelting). However,
liu' ni.i'it ed concent rut t?s are di.ssolved in sulfuric acid and processed in
aqueous solution rattier than being thcrmochemically treated to reduce
and volatilize zinc. The electrolytic method is considered to be more
efficient than the retort method in the recovery of lead, copper, and
precious metal values associated with the zinc sulfide concentrates
which constitute the primary feed to zinc reduction plants of both types.
A simplified flow diagram identifying solid waste sources and
disposal is shown in Figure 4. The primary process steps in the electro-
lytic method comprise (1) roasting the zinc sulfide concentrate to reduce
the sulfur content, (2) leaching the calcine with acidic stripping electro-
lyte, (3) thickening and filtration to remove the leach residue, (4)
oxidation of the dissolved iron followed by neutralisation of acid with
lime, calcine, rind metallic ;:inc to precipitate copper, iron, arsenic,
antimony, cadmium, germanium, nickel, and cobalt, (.5) electrolysis of
tin- purified /inc .uid an acidic .stripped electrolyte for reuse in leach-
ing fresh i-ah-ine. The cathode deposit of rinc is melted and casted
into high-purity .slab zinc. (0) The auxiliary operations include (a)
collecting and returning all dusts and fumes to the roasters, (b) wash-
ing and drying the zinc leach residue with its lead and precious metal
content for reprocessing at a lead smelter, (c) selective leaching of
the solution purification residue to recover a cadmium-rich solution
for electrolysis with zinc as a returnable solution, and copper as a
67
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residue for shipment to a copper smelter. In addition to these opera-
tions, part of the spent e>loctrolvu> must periodteaUv or Continuously
be bled from thi circuit for periodic purification or discharge lo.
to remove impurities, such as chlorine, fluorine, cobalt, magnesium, and
sodium that gradually build up in the electrolyte.
Pyrometallurgical. The hypothetical pyrometallurgical zinc
plant is assumed to have an annual capacity of 107,000 MT (118,000 short
tons) of 99.+% zinc metal. This figure is based on the approximate size
of the smaller of the two actual vertical retort plants operating in the
U.S. and on the fact that the larger of the two operating plants has
approximately double this capacity at 227,000 MT (250,000 short tons).
No consideration is given to the size of the two remaining horizontal
retort zinc plants in the U.S. because both are scheduled to cease opera-
tions.
Pyrometallurgical zinc smelters are estimated to operate 350
days/year; the daily production and assumed capacity of the hypothetical
plant is, therefore, 310 MT (338 short tons).
Input (feed) to the plant consists of roughly 214,000 MT
(236,000 short tons) annually of sulfide concentrates, oxide ores, and
zinc scrap averaging roughly 50% zinc in the aggregate.
The major operational steps in the vertical retort process
arc (1) Rousting to doorcase the sulfur content of the zinc concentrate;
(2) sintering to remove residual sulfur and volatilize cadmium, lead,
indium, germanium and selenium; (3) retorting a thoroughly mixed charge
of sinter, and carbon reductant in either a vertical retort or electro-
thermic furnace; and (4) condensing the volatilized zinc. Roaster
gases are scrubbed free of dust, and the sulfur dioxide content converted
to sulfuric acid (5). Dust and fumes from the roasters and sinter hearths
are leached with sulfuric acid to recover cadmium (6). Oxide ores, scrap
and reverts are pretreated in a Waelz kiln to upgrade the zinc content (7).
Figure 5 is a smiplified flow sheet showing the sources of
solid wastes and their disposal. The main retort residue is composed of
residual coked coal, residual zinc, ferrosilicon, and a siliceous slag
component.
3.2.2 Description of Waste Streams
Electrolytic. a. The chief solid residue in the electrolytic
reduction of zinc from sulfide ore concentrates is the leach residue (i.e.,
that portion of the calcined feed which fails to dissolve in the sulfuric
acid leach solution). This residue does not, however, constitute a
waste at the typical installation; it is shipped immediately to a lead
smelter for recovery of lead and precious metal values.
69
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k- Miscellaneous Acidic Slurries. Acidic residue from
sources such as s»as scrubber bleed, preleach bleed, dross and dust leach
I)Iced, anode washing, and spent acid bleed contain suspended and dissolved
solids, as well as sulfuric acid and heavy metals (Cd, Cu, Pb, Zn).
These water slurries in the amount of 1,480 cu m/day (391,700 gal/day)
are settled in an unlined lagoon (A) of capacity 2000 cu m (72,000 cu
ft.)- Settled solids are dredged from the lagoon and dried on land near
the plant prior to eventual shipment to a lead smelter. The amount of
sludge dredged (on a dry basis) is 9.1 Kg/MT of zinc product or 2.6
MT/day. Wet sludge generation will be in the order of 23 MT/day.
Overflow from the above lagoon goes to a liir.e treatment facility
whore it joins with acid plant blowdown slurry as discussed below.
c. /\cj.d Plant Blowdown Slurry. The amount of calcium
sulfute- sulfite sludge resulting from lime treatment of acid plant
blowdown as dry weight is approximately 31.5 Kg/MT zinc product (approximately
105 Kg/NTT wet weight). Table 26 contains estimates of the daily flow
volumes and solids contents of limed slurry from acid plant blowdown
plus limed overflow from the lagoon (A) in which the above-described
miscellaneous acid slurries are settled. These solids are dredged from
a second unlined lagoon (B).
Solids dredged from lagoon B (17 Kg/MT of zinc product or 4.8
MT/day) are dried on land for eventual shipment to a lead smelter.
TABLE 26
SOLIDS IN TWO LIMED EFFLUENT STREAMS PRESENT AT AN
ELECTROLYTIC ZINC PLANT
Flow Solids (Dredged from Lagoon B)
_S^t£eani_ cu m/day gal/day MT/day S. ton/day MT/year
Lagoon (A) Overflow
At id I'lanl Ulowdown (B)
Combined
1,480
1,380
2,860
3.92,000
364 , 000
756,000
8.5
9.0
17.5
9.4
9.9
19.3
2,960
3,150
6,110
d. Other Residues. As shown in Figure 4 other residues yielded
by the component processes of electrolytic zinc reduction are recycled
immediately or shipped to copper or lead smelters. Since these materials
are not stored or dried on the ground, they are not considered to be a
part of the solid waste disposal scenario.
71
-------
Pyrometallurgical
u. l Acid Plant Blowdown Slurry. The amount of calcium sulfate-
sulfite sludge resulting from lime treatment of acid plant blowdown as
dry weight is 31 Kg/MT zinc product (approximately 100 Kg/MT wet weight).
Table 27 contains estimates of the daily flow volumes and solids contents
of limed slurries from acid plant blowdown and primary gas cleaning
operations which are combined prior to liming and then settled as an
intimate mixture in the lagoon before dredging.
TABLE 27
soi.rns IN TWO I.IMI-D INFLUENT STREAMS PRESENT AT A
I'YKOMhTALI.URGICAL ZINC PLANT
Flow Solids (Dredged from Lagoon)
Stream cu m/day gal/day MT/day Short ton/day MT/year
Primary Gas 1,090 288,000 28 31 9,950
Cleaning Operation
(Sinter, Roasters, Crushers)
Acid Plant Blowdown 1,310 345,000 9.0 9.9 3,150
Combined 2,400 633,000 37 41 13,100
72
-------
The total amount of dredged sludges from these sources amounts to 122
Kg/MT of zinc product or 37 MT/day or 13,100 MT/year (37,430 MF/year wet
weight). The dredged sludges are either stored for long periods of time
(i.e. months) before recycling or disposed permanently on land.
ll- Ketort ('.as Scrubber Bleed Slurry ("Blue Powder") . A
small portion of the zinc product vapor produced in the retorts passes
through the zinc condensers and is recovered as "blue powder" from wet
scrubber bleed .slurry and settled in two concrete pits. Slurry volume
flowing into the concrete pits is 136 to 272 cu m/day (1500 to 3000 gal/day)
tiach pit is approximately 20' x 40' in size. It is estimated that 10
Kg/MT of zinc product or 3.1 Ml/day (3.4 short tons/day) of blue powder
is dredged from the pits and that the amount stored on the ground for
months for drying and oxidation before eventual recycle is roughly 535 MT
at any given time.
°- Cadmium Plant Residue. Some fume and dust from primary gas
cleaning operations is treated in a hydrometallurgical plant to recover
cadmium metal. A filter cake residue amount to 1,8 Kg/MT of zinc product
or 0.514 MT/day (0.566 short tons/day) is trucked to the main retort
residue dump for disposal.
Chemical analyses of samples obtained from primary zinc plants
are given in Appendix A.
Retort residue consists of friable 3-5" diameter gravel like
lumps. In solubility tests as described in Appendix B the only heavy metal
found to leach significantly was zinc (230 ppm in leachate). Zinc is
not a highly toxic metal and a plant and animal trace nutrient. Retort
residue, in view of available information, is not considered potentially
hazardous at this time.
Cadmium plant residue contains high concentrations of cadmium,
copper, lead and zinc. Lead was found to leach to the extent of 9 ppm
and zinc to the extent of 4,000 ppm from cadmium plant residue. This
residue is therefore considered potentially hazardous.
Sludges from gas cleaning, acid plant blowdown lime treatment
(i.e. gypsum sludges) and anode sludges contain high concentrations of
the highly toxic heavy metals cadmium and lead and high concentrations of
zinc and copper also. These elements were found to leach significantly in
solubility tests described in Appendix A. The silt to colloidal sizes of
individual sludge particulates make sludges susceptible to weathering
processes. For the above reasons sludges are considered potentially
hazardous when disposed or stored on land for periods of months or greater.
73
-------
3.2.3 Waste Quantities
Tables 28 and 29 summarize the generation factors for each of
the land disposed or stored residuals in the electrolytic and pyrometallurg-
ical primary zinc smelting and refining categories. These tables also
gives typical concentrations of potentially hazardous constituents.
Table 30 gives the yearly quantity of wastes and potentially hazardous
constituents for a typical plant producing 100,000 MT (110,000 short
tons) of zinc metal by the electrolytic process. Table 31 gives this
information for a typical plant producing 107.000 MT (118,000 short
tons) of zinc metal per year using pyrometallurgical processes (i.e.
vertical retort or electrothermic).
Using the waste generation factors and concentration factors
contained in Tables 28 and 29 and smelter capacities estimates have been
made for the quantities of land disposed or stored residuals and hazardous
constituents thereof. These estimates are given for 1974, 1977 and 1983
on a state- by-state, regional and national level in Tables 32 and 33.
It will be noted that the amount of potentially hazardous
sludge from pyrometallurgical zinc production is expected to decrease
over the period 1974-1983 while the amount of potentially hazardous
sludge from electrolytic refining is expected to increase significantly,
especially from 1977 to 1983. This is because pyrometallurgical zinc
capacity is expected to decrease over this period while electrolytic
capacity is expected to increase. The net result of a technology shift
to electrolytic refining of zinc from pyrometallurgical refining should
be a marked decrease in sludge generation since pyrometallurgical opera-
tions produce approximately 125 Kg of sludge/MT of zinc product and
electrolytic produces approximately 25 Kg/NTT of product.
Projections of land disposed waste for 1977 are based on
increased electrolytic plant capacity now under constructior in Oklahoma
(51,000 MT). Further planned increases in electrolytic plant capacity
in Kentucky (163,010 MT) and Tennessee (145,000 MT) (Reference 9) account
for estimated waste increases by 1983.
On the other hand the closing of pyrometallurgical plants in
Oklahoma and Texas will result in no wastes from pyrometallurgical
plants in these states in 1977 and 1983. Pyrometallurgical plant expansion
in Pennsylvania will increase land disposed waste by an estimated 11
percent.
3.3 TREATMENT AND DISPOSAL TECHNOLOGY
3. 3. J Ciirrent Jtes,te Treatment and Disposal Practices
For purposes of clarity, treatment and disposal of residues
will be discussed separately for electrolytic and pyrometallurgical
plants.
74
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85
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SOURCE: C/
82
-------
Table 33-d
ESTIMATED STATE, REGIONAL, AND NATIONAL WASTE FROM
PRIMARY PYROMETALLURGICAL ZINC SMELTING AND REFINING
TOTAL RETORT RESIDUE -1977,1983 (METRIC TONS) *
STATE
PENNSYLVANIA
EPA REGION
III
NATIONAL TOTAL
TOTAL
DISPOSED
120,990
120,990
120,990iA)
TOTAL
POTENTIALLY
HAZARDOUS
0
1
F
DISPOSAL
METHOD
OPEN DUMP
CONSTITUENTS
Cd
103
103
103
Cr
5.4
5.4
5.4
Cu
555
555
555
Pb
289
289
289
Se
1.0
1.0
1.0
Zn
12,987
12,987
12,987
NOTES:
(A) A SMALL PORTION OF THIS MATERIAL IS SOLD FOR CONSTRUCTION USES.
•THIS WASTE BELIEVED NONHAZARDOUS BASED ON CALSPAN
SOLUBILITY TESTS DESCRIBED IN APPENDIX B.
MULTIPLY BY 1.1 TO CONVERT TO SHORT TONS.
SOURCE: CALSPAN CORPORATION
83
-------
Hlectrolytic. Acid plant gypsum sludge and sludge from scrubber
bleeds, anode cleaning, spent acid anc other bleeds is currently dredged
from unlined lagoons and either disposed of or stored on the ground for
variable periods of time (i.e. a few too many months) before shipment to
lead smelters for recovery of lead and other metals. These practices
are inadequate because of the danger of heavy metal leaching as discussed
previously.
Pyrometallurgical. Currently all but one pyrornetallurgical
plant disposes of retort residue and cadmium plant residue (iron press
residue) by open dumping. This practice is considered adequate at the
present time for retort residue but not cadmium plant residue which may
be leached of heavy metals.
Gypsum sludge from the acid plant and blowdown bleed slurry
(i.e. "blue powder") from the retorts are kept temporarily in unlined
lagoons which are periodically dredged. Dredged sludges are stored on
land and may eventually be recycled. As discussed previously these sludges
contain potentially hazardous heavy metals including arsenic, lead,
cadmium, zinc and copper and released some of these constituents in
solubility tests.
The following sections describe levels of technology for those
residuals from electrolytic or pyrometallurgical smelting and refining
which are considered potentially hazardous.
3.3.2 Present Treatment and Disposal Technology (Level I)
1: lectrolytic. Settling of sludges and bleeds (i.e. acid
plant gypsum sludge, scrubber bleed, anode cleaning spent acidj in
unlined lagoons with disposal or storage on unsealed soil areas is environ-
mentally inadequate if heavy metals leach through permeable soils to
ground water.
Fj^rqnieta_llur,^i_caJ . Open dumping of cadmium plant residue is
IMIV i roiimeiit ,i! ly i n.ideqii.i t e bec.mse of the possibility of heavy metal
leaching. Temporary storage of acid plant gypsum sludge and blowdown
bleed slurry (i.e. "blue powder") in unlined lagoons and storage of dredged
.sludge.s on J.jnd for considerable time periods (one month to man)' months)
is environmentally inadequate because of the threat of heavy metal
leaching and percolation through permeable soils to groundwater.
3.3.3 Best Technology Currently Employed (Level II)
Electrolytic. The technology employed for the settling and/or
storage of sludges and bleeds is the same as that employed for Level I.
84
-------
a I lur^ii-u I . Ai Itu-.t our plain is Known to ".hip iron
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3.4. 1 Cost of Present Treatment and Disposal Technology (Leve1 I)
hloctrolytic Reduction Plants. Two lagoons are used for the
settling of solids. One is relatively small, 2000 m ; the other,
1.1,440 m , is sized to allow four days of residency. Characteristics
of the lagoons are shown below:
Lagoon
Bottom width 14 m 39 m
lop width 26 m 51 m
lint torn length 27 m 78 m
Top length 39 m 90 m
Total depth 3m 3m
Depth of excavation 1.75 m 1.1 m
Circumference 142 m, 295 m
Dike Volume 975 nfl 3,808 m!?
Dike Surface 1,761 m 4,071 m
Total width 37 m 65 m
Total length 50 m 104 m
Required area .2 ha .7 ha
The lagoons are constructed on rural land.
The daily inflow into Lagoon A is 1,480 m, which contain 9.1 MT
of solids. These solids are estimated to form 12 m of sludge daily.
The daily inflow into Lagoon B is 2,860 m which contain 17.5 MT
of .soijil.s. These solids are estimated to form 23 m of sludge.
Both lagoons are dredged periodically. The sludge is deposited
next to the lagoons, allowed to dry and then recycled. Only the cost of
lagoon dredging is included.
The lagoons are assumed dredged at 25 day intervals. The amount
of sludge removed each time totals 875 m or 12,250 m for the year. A
3/4 yard dr i;, line is employed. It has an hourly capacity of 27 m (1).
Consequently about 460 hours (25%) of equipment operating time are
required annually. Costs of Level I treatment and disposal technology
are given in Table 37.
Pyrometallurgical^
The annual capacity of the plant is 107,000 MT operating 350
days/year. Effluents from primary gas cleaning operations and acid blow-
down totaling 2,400 m /day go to a lagoon. The lagoon is designed for
4 day residency. Its design characteristics are:
92
-------
TABU; 37
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ZINC PLANT - ELECTROLYTIC REDUCTION
Capital Cost
Lagoon
Site Preparation
Survey
Test Drilling
Sample testing
Report Preparation
Construction
Excavation tj Forming
Compacting
Fine Grading
Soil Poisoning
Transverse Drain Fields
Land
Lagoon
Site Preparation
Survey
Test Drilling
Sample Testing
Report Preparation
Construction
Excavation § Forming
Compacting
Fine Grading
Soil Poisoning
Transverse Drain Fields
Land
Equipment
Dragline (25%)
Total
Sludge (from acid plant
anode washings,
spent acid § bleeds)
125
490
250
1,200
1,295
1,805
790
175
340
350
440
490
250
1,200
5,065
7,045
1,830
365
600
1,225
93
-------
TABLE 37 (continued)
Annual Cost
Sludge
Land $ 160
Amortization
Construction 2,760
Equipment 2,780
Operating Personnel 5,590
Repair and Maintenance
Construction 580
Equipment 875
Energy
Fuel 525
Electricity 55
Taxes 40
Insurance 430
Total $13,795
94
-------
Volume 9,600 m
Bottom width 36 m
Top width 48 m
Bottom length 71 m
Top length 83 m
Total depth 3 m
Depth of excavation 1.2 m
Circumference 273 m
Dike volume 3,241 m
Dike surface 3,713 m
Total width 61 m
Total length 96 m
Required area .6 ha
The lagoon is located on rural land since primary zinc smelters are
located in rural areas.
The daily lagoon inflow contains 37 MT of solids. These are
estimated to result in the formation of 48.7 m of sludge. It is assumed
that the lagoon is dredged when the sludge accumulation equals 50 percent
of the lagoon volume. Thus the lagoon is dredged every 99 days or 3.5
times/year. This requires 625 hr/yr of dragline time. The sludge is piled
adjacent to the lagoon, dried, and then hauled to the retort residue dump.
A small portion of the zinc product vapor produced in the
retorts passes through the zinc condensers and is recovered as "blue powder"
from wet scrubber bleed slurry settled in two concrete pits. Each pit
measures about 12 x 6 x 2.5 m. Wall thickeness is about .4 m requiring
36 x 2.5 x .4 m or 34.2 m of concrete in place for wall construction.
About 3.1 MT of blue powder is dredged daily from the pits,
stored on the ground for drying and oxidation and eventually recycled.
One hour of backhoe time is assigned daily for this task. The average
amount stored at any given time is about 535 MT.
Approximately 16,800 m of dried sludge are generated annually.
Transport requirements to load and haul the sludge to a dump are estimated
to be 895 hours of truck and front loader time.
The dump area, providing space for the accumulation of 20 years
of dried sludge extends over 3.4 hectares. The dump is located on rural
land. About 200 hours of bulldozer time arc assigned at the dump for
spreading and compacting the sludge.
A samll amount of waste, about 0.5 MT/day, is generated by the
associated cadmium plant. This waste is presently disposed with the
wastes from the retort furnaces. Its disposal cost is negligible.
Cost of Level I treatment and disposal are given in Table 38.
95
-------
TABU! ^H
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ZINC PLANT - PYROMETALltlRGlCAl. REDUCTION
Capital Cost
Lagoon Sludge
Site Preparation
Survey $ 375
Test Drilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation £ Forming 4,310
Compacting 5,995
Fine Grading 1,670
Soil Poisoning 340
Transverse Drain Fields 560
Land 1,050
Sumps 19,360
Equipment
Dump Truck (50%) 12,500
Frontloader (65%) 13,000
Dragline (35%) 24,500
Bulldozer (10%) 1,600
Dump
Survey 2,025
Land 5,950
Total $95,175
Annual Cost
Land $ 700
Amortization
Construction 4,245
Equipment 8,205
Operating Personnel 33,730
Repair and Maintenance
Construction 965
Equipment 2,580
Energy
Fuel 3,765
Electricity 375
Taxes 175
Insurance 950
Total $55,690
96
-------
.3.4.2 Cost of Best Technology Currently Employed (Level II)
^cctrolyticReduction Plants. The technology and costs for
Level JJ are the same as Level I.
I'yromotul hirgical Plants. The hluo powdtM- Jrodged from the pit
is recycled without prolonged ground storage. A 20 x JO m concrete pad
is constructed adjacent to the pit to prevent leaching during the temporary
storage. The cadmium containing waste is immediately sent to a lead
smelter. Associated capital and annual costs are shown below.
Capital Costs
Concrete Pad $ 6,140
Total $ 6,140
Annual Cost
Construction Amortization $ 710
Construction Repair 6 Maint. 185
Insurance 60
Total $ 955_
Cost of TochnoJojjy To Provjjjje_jUkKuaate Health mid I^vironjwontal
I'rotorUon (.Level 111)
Electrolytic Reduction Plants. Level III consists of lining
both lagoons. Additionally, the lagoon sludges are pumped into concrete
pits prior to being recycled. This procedure eliminates the requirement
of the dragline used for lagoon dredging in Level I and its attendant costs.
The pits are sufficient to accommodate 25 days of accumulated sludge.
Net incremental costs are shown in Table 39. The pit dimensions
are 3 x 6 x 10 m and 3 x 10 x 20 m, respectively. The slurry pump has a
capacity of 236 1/min and is operated 740 hrs/yr; 1000 hrs/yr of labor
are assigned. The assumed, required piping consists of 200 m of installed
pipe and 100 m of flexible pipe, both 7.6 cm in diameters.
Pyrometallurgical Plants. The lagoon is lined and the lagoon
sludge is chemically fixed. The costs are the same for rail and truck
transport of the residue. Costs are given in Table 40.
Costs of Levels I, II and III treatment and disposal technology
are summarized in Table 41 for electrolytic plants and in Table 42 for
pyrometallurgical plants. Costs are given in dollars per metric ton of
dry and wet waste and dollars per metric ton of zinc product. Costs for
each level of technology and each type of waste are also expressed as a
percent of metal selling price.
97
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TABLE 39
COSTS OF LEVEL III TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ZINC PLANT - ELECTROLYTIC REDUCTION
Capital Cost
Construction Sludge
Lagoon A Liner $ 6,020
Lagoon B Liner 22,515
Concrete Pit (Lagoon A) 11,440
Concrete Pit (Lagoon B 21,420
Equipment
Slurry Pump 13,730
Pipe Rigid (Installed) 4,425
Pipe Flexible 440
(Dragline) (24,500)
Total $55,490
Annual Cost
Land
Amortization
Construction $ 7,130
Equipment I/ (940)
Operating Personnel I/ 4,560
Repair and Maintenance
Construction 1,840
Equipment I/ (295)
Energy
Fuel I/
Electricity (700)
Taxes
Insurance \J 555
Total $12,150
\j Costs shown are not costs, i.e. costs of new equipment less dragline
associated costs which are no longer incurred.
98
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TABLE 40
W/i OF LEVEL III TREATMENT AND DISPOSAL TECHNOLOGY PRIMARY ZINC PLANT
PYROMETALLURGICAL REDUCTION
Construction
Lagoon Liner
Total
CAPITAL COST
Sludge
$ 19,800
$ 19.800
ANNUAL COST
Land
Amortization
Construction
Equipment
Operating Personnel
Repair and Maintenance
Construction
Equipment
Energy
Fuel
Electricity
Taxes
Insurance
Chemical Fixation
$ 2,295
595
Total
200
221,760
$224,850
99
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Total estimated industry annualized costs for treatment and
disposal of sludge from electrolytic zinc production in 1973 dollars are
given in Table 41 while the estimated 1973 costs for sludge disposal
from the pyrometallurgical process are given in Table 42. The annualized
industry cost for Level I and Level II treatment and disposal of sludge
from electrolytic zinc smelters is $30,000 or 0.03% of national sales.
The annualized cost for Level III treatment and disposal technology
(i.e. adequate for environmental protection is $50,000 or 0.06% of
national sales.
The annualized industry cost for Levels I and II treatment and
disposal technology for sludge from pyrometallurgical zinc production is
$160,000 or 0.11% of national sales. The annualized industry cost for
Level III treatment and disposal technology is $800,000 or 0,58% of
national sales.
102
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4.0 1'RJMARY SMhLTlNC AND RLF1NING OI; ALUMINUM (SIC 3334)
4.1 INDUSTRY CHARACTERIZATION
There are 31 primary aluminum plants in the United States owned by
thirteen different firms. Major producers are Aluminum Company of America
(ALCOA), Anaconda Aluminum Company, Kaiser Aluminum § Chemical Corporation
and Reynolds Aluminum. Total U.S. plant capacity is 4,418,000 MT/year.
Distribution of plants by capacities is shown in Table 43. Output from
primary aluminum reduction cells is molten aluminum which is usually cast
into pure and alloy ingots and billets before shipping for product manufac-
ture. The total U.S. value of primary aluminum production is given as
$2,206,440,000 for 1973 (Ref. 2).
Aluminum metal is manufactured by the Hall-Heroult process, which
involves the electrolytic reduction of alumina dissolved in a molten salt
bath of cryolite (a complex of NaF-AlF3) and various salt additives:
Elecfrolysis
A1 o > 4A1 -I- 30
23 2
AJumiii.i Aluminum Oxygen
The electrolysis is performed in a carbon crucible housed in a
steel shell, known as a "pot." The electrolysis employs the carbon crucible
as the cathode (negative pole) and a carbon mass as the anode (positive pole).
The three types of anode configurations used are: prebaked (PB), horizontal-
stud Soderberg (HSS), and vertical-stud Soderberg (VSS).
The major portion of aluminum produced in the United States is pro-
cessed in cells using prebaked anodes. In this type of cell, the anode con-
sists of blocks that are formed from a carbon paste and baked in an oven
prior to their use in the cell. These blocks—typically 14 to 24 per cell--
are attached to metal rods and serve as replaceable anodes. As the reduction
proceeds, the carbon in these blocks is gradually consumed at a rate of about
1 inch per day by reaction with the oxygen by-product.
The second most commonly used anode configuration is of the hori-
zontal-stud Soderberg type. In this type of anode, pitch and carbon "paste"
are added periodically at the top of the superstructure and the mixture is
baked in place as it descends closer to the molten bath. The entire anode
assembly is moved downward as the carbon is consumed. The cell anode is
contained by aluminum sheeting and preformed channels through which electrode
connections, called studs, are inserted into the anode paste. As the baking
anode is lowered, the lower row of studs and the bottom channel are removed,
and the electrical connectors are moved to a higher row.
103
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The vertical-stud Soderberg anodo is similar to the horizontal-stud
anode, with the exception that the studs are mounted vertically in the cell.
The studs must be raised and replaced periodically. Representative raw mate-
rial and energy requirements for aluminum reduction cells are presented in
Table 44. A schematic representation of the reduction process is shown in
Figure 6 . Geographical distribution of U.S. primary aluminum plant capacity
according to the types of processes employed (i.e., prebake, Soderberg) is
given in Table 43 .
4.2 WASTE CHARACTERIZATION
The hypothetical plant is assumed to have an annual capacity of
153,000 MT (169,000 short tons) of aluminum. This figure was obtained by
averaging the 1973 production levels for those plants using wet-type primary
air pollution control equipment and having a scrubbing liquor treatment sys-
tem that is likely to require a settling basin or lagoon. Since aluminum
reduction plants operate around the clock, 365 days a year, the daily capa-
city of the hypothetical plant is 420 WT (463 short tons).
4.2.1 Process Description
Essentially, all primary aluminum reduction plants use the Hall-
Heroult electrolytic process (discussed previously) for producing free alumi-
num metal from alumina (A^Oj). The aluminum is produced in individual cells,
which for the assumed typical plant total 700, each having a capacity of
600 Kg/day (1320 Ib/day). The cells (or pots) are arranged in lines, called
potlines, and are housed in potrooms. Fumes from the cells are collected
by hoods located directly over the cells (primary control). In the hypothe-
tical plant considered, the fumes collected by the hoods are scrubbed in a
wet-type system. The assumed plant has no system for controlling the fumes
that are not captured by the primary system. These uncollected fumes are
vented to the atmosphere through ventilators located in the roofs of the
potrooms.
Input materials for the potrooms consist of carbon anodes, carbon
potliners (the cathodes), alumina (A^O^) , Cryolite (Na3AlF6), aluminum
fluoride (A1F3), and calcium fluoride (CaF2). The anodes are prepared in
the carbon plant where petroleum coke is crushed, screened, classified and
mixed with a pitch binder. The mixture is pressed into the desired shape
and baked. For Soderberg anodes, the pressing and baking operations arc
omitted, and the prepared paste is introduced directly into the anode. The
hypothetical plant is of the probaked anode configuration. The cathodes,
or potliners, ure also made in the carbon plant. Anthracite and pitch are
used in their preparation.
When a sufficient quantity of aluminum metal has formed in a cell,
it is drawn off into a transfer crucible and sent to the cast house where
it is placed directly in a holding furnace or cast into pigs. Usually the
molten metal is fluxed before casting to remove minor impurities.
105
-------
Table 44
RAW MATERIAL AND ENERGY REQUIREMENTS FOR ALUMINUM PRODUCTION
Parameter
Representative value
Cell operating temperature
Current through pot line
Voltage drop per cell
Current efficiency
Energy required
Weight alumina consumed
Weight electrolyte fluoride consumed
Weight carbon electrode consumed
-1740 F (~9508C)
GO.OOOto 125,000 amp
4.3 to 5.2
85 to 90%
6.0 to 8.5 kwh/lb aluminum
(13.2 to 18.7 kwh/kg aluminum)
1.89 to 1.92 Ib AL203/lb aluminum
(1.C9to 1.92 kg Al_203/kg aluminum)
0.03 to 0.10 Ib fluoride/lb aluminum
(0,03 to 0 10 kg fluoride/kg aluminum)
0.45 to 0.55 Ib electrode/lb aluminum
(0.45 to 0.55 kg electrode/kg aluminum)
Source: Compilation of Air Pollutant Emission Factors,
2nd Ed., U.S.E.P.A., 1973 (AP-42)
106
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COAL TAR PITCH I SOFT PITCH
PETROLEUM COKi ANTHRACITE
PASTE PLANT
T
ANODE I CATHODE
I
CRYOLITE
CALCIUM
FLUORIDE
]
ALUMINA
ELECTROLYTIC REDUCTION CELL (POTROOM)
SODERBERG CELL
PREBAKED CELL
Figure 6 PROCESS DIAGRAM FOR PRIMARY ALUMINUM SMELTING
107
-------
4.2.2 Description of Waste Streams
Compounds of fluorine are the major potentially hazardous materials
found in the wastes generated in primary aluminum smelting. They are present
in spent anode butts, in spent potlines, and in dusts deposited in the pot-
rooms. At present, there are nine plants that employ dry systems for primary
control of emissions. Of these nine plants, two employ wet systems for con-
trol of fugitive emissions which escape from furnaces into work areas. The
remainder require no fugitive emissions controls. Seventeen plants have wet
primary control systems, and at least four of these have wet fugitive emissions
controls. The remainder have no fugitive emissions controls. There are four
plants that control part of the potlines with dry primary systems and the
remaining potlines with wet systems. None of the four plants have fugitive
emissions controls. Figure 7 illustrates the aluminum smelting and refining
processes and associated wastes.
Only those plants that employ prebaked anode require anode bake
plants. Of 21 plants of the prebake type, four employ wet-type systems for
control of bake plant emissions, three use dry-type systems, and nine have
no bake plant emission control systems. The controls for the remaining five
plants are not known.
Sludges
Emissions from the electrolytic cells amount to 36.6 Kg/MT aluminum
and contain 11.1 Kg fluoride. Emissions from the anode bake furnace amount
to 1.5 Kg/Mr with 0.5 Kg fluoride. Wet scrubbers are assumed for emissions
control in the electrolytic and anode bake furnaces. The scrubber waters are
thun limo treated resulting in sludges which are discharged or possibly
recycled to the scrubbers. It is the treatment of this scrubber water used
for primary air pollution control that accounts for the major portion of
potentially hazardous sludge wastes generated at primary aluminum smelting
plants. As shown in Figure 7 , 117 Kg of particulates per metric ton of
molten aluminum containing 18 Kg of fluoride are discharged to lagoons after
lime treatment. The lagoons are periodically dredged and dredged material
disposed on land.
For the hypothetical plant wet scrubber,systems were assumed for
emission control in the potrooms and the anode bake plant. In practice, the
type of emission control varies widely within the industry. Some plants use
dry-type control systems throughout, some use a mixture of wet and dry sys-
tems, and other use wet-type systems only. For example, sludge generated in
the recovery of cryolite from the potroom scrubber water amounts to less
than 20 Kg per metric ton of aluminum produced. In plants that recover cryo-
lite from spent potliners, as well as from potroom scrubber water, sludge
generation amounts to about 125 Kg per metric ton of aluminum. On the other
hand, in plants that do not recover cryolite, sludge generated in the treat-
ment of potroom scrubber water averages an estimated 95 Kg per metric ton of
aluminum and can exceed 200 Kg per metric ton of aluminum for once-through
lime treatment systems. In general, the use of dry systems reduces the amount
of waste destined for disposal because dry collection, such as alumina bed
adsorption allows direct recycle of captured materials. Baghouses are already
108
-------
COM
PARTICIPATES
,,GF ft GASES
..« WICELLNOOO
92PF EFFICIENCY)
INGOTS, tic
NOTE: ALL VALUE! GIVEN A$ K| Of WASTE PER METRIC TON Or ALUMINUM PRODUCED
TV - TOTAL PARTICULATft; f • f LUOHINE; ff • PARTICULATE FLUOHIDE8.
OF • QASEOUI FLUORIDES. CN • CYANIDES
Figure 7 PRIMARY ALUMINUM PRODUCTION
109
-------
commonly used in carbon plants, and the collected dust (mostly carbon) can
be directly recycled in the anode preparation operations.
With respect to wet scrubber systems, it should be noted that there
are several alternatives for treating the scrubber water. In the hypothetical
plant, once-through lime treatment was assumed. Other alternatives include:
recycling of the lime-treated scrubber water, and recycling of the scrubber
water accompanied by recovery of the cryolite. The amount of sludge generated
in scrubber water treatment systems that incorporate a cryolite recovery capa-
bility can be significantly less than that generated in once-through lime
treatment systems.
Spent anode butts. During the electrolysis of the aluminum
to obtain molten aluminum, the carbon anodes are consumed to the point that
they can no longer be used. At this point, the anode remnants or "butts" are
broken up and sent back to the carbon plant for reclamation of carbon to pre-
pare new anodes. Large chunks of anode butts are either immediately sent to
the carbon plant for reprocessing or stored for short periods of time before
reprocessing.
Spent pot liners. Cathode potliners become partially saturated
with cryolite (Na.,AlF,). After continued use, they crack and the molten
aluminum becomes contaminated with iron from the outer shelf of the
cathode. At this point the cathodes must be replaced. Approximately 53 Kg
of spent potliners containing 9 Kg of fluoride are produced per metric ton
of molten aluminum. The spent potliners coated with cryolite are stored
outside to await shipment to a purchaser who will process them to recover
cryolite or to await later processing by the smelter itself if facilities
are available. Storage may be for periods of many months to a year or more.
Potroom skimmings. Potroom skimmings are residues skimmed from
the hot metal after tapping. These residues, consisting almost entirely of
hath material, are generated at the rate of about 5.5 Kg prr IT..:trie ton of
aluminum produced and contain about 2.0 Kg of fluorides per metric ton of
aluminum. Generally, the skimmings are returned directly to the pots. In
some cases, the skimmings are handled in the same manner as spent potliners
(i.e., stored and reprocessed).
Shotblasting dust. In the recovery of spent anodes, the steel
rods in the center of the anodes are sand blasted to remove attached carbon
and cryolite. Dust from this operation is generated at a rate of about
5 kilograms per metric ton of aluminum and contains 0.5 Kg of fluoride.
This waste is normally landfilled.
Skim processing dust. Alloying furnaces in the cast houses pro-
duces a skim on the surface of the molten aluminum metal. This skim is
removed and processed for metal recovery in small furnaces. Dust emissions
collected from this operation amounts to approximately 2.5 Kg/MT of aluminum
product.
110
-------
Chemical analyses of waste samples obtained from primary aluminum
plants are given in Appendix A . Spent pot liners consist of large and small
chunks of carbon coated with cryolite. In solubility tests as described in
Appendix B , high concentrations of fluoride and cyanide were found to leach
from spent potliners. Potliners are therefore considered potentially hazard-
ous. Although solubility tests were not conducted on pot skimmings, soluble
fluorides are- expected to leach from them. Skimmings are therefor*; considered
potentially hazardous.
Sludges originating from wet scrubbing of emissions from the anode
bake plant or aluminum reduction cells were found to release soluble fluoride
and/or cyanide in solubility tests described in Appendix B . They are there-
fore considered potentially hazardous.
Although shot blast dust was found to contain a high concentration
of copper, very little was solubilized in solubility tests. It is not con-
sidered potentially hazardous. Cast house dust contains significant concen-
trations of the heavy metals Cu, Zn, Pb, Cr, and Ni. Solubility of these
metals was not ascertained. Cast house is considered potentially hazardous
at this time because of the high concentration of heavy metals and small
particle sizes.
4.2.3 Waste Quantities
Table 45 summarizes the generation factors for each of the land-
disposed or stored residuals in the primary aluminum smelting and refining
category. This table also gives typical concentrations of potentially
hazardous constituents. Table 46 gives the yearly quantity of wastes and
potentially hazardous constituents for a typical plant producing 153,000
metric tons (169,000 short tons) of aluminum per year.
Using the waste generation factors and concentration factors con-
tained in Table 45 and smelter capacities, estimates have been made for the
quantities of land-disposed or stored residuals and hazardous constituents
thereof. These estimates are given for 1974, 1977, and 1983 on a state-by-
state, regional and national level in Table 47.
Projections for increased quantities of potliners and dust for 1977
were based on planned plant expansions as given in the March 1975 issue of
Engineering ami Mining Journal. Projections for 1983 were computed from the
1977 data assuming a four percent yearly growth indicated by representatives
from the industry.
Our projections for 1977 and 1983 also take into account the trend
toward dry emission control systems. Note that the projected total amounts
of sludge for 1977 and 1983 are less than the national total for 1974. The
1977 projections were based on published information on plant expansions.
Because similar information was not available for 1983, the 1983 levels for
ducts and potliners were computed from the 1977 data assuming a four percent
yearly growth. No increase in sludge amounts are shown for 1983 over 1977
because all expansions are assumed to use dry systems for potroom emissions.
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4.3 TREATMENT AND DISPOSAL TECHNOLOGY
4.3.1 Current Waste Treatment and Disposal Practices
Potliners and Pot Skimmings. Currently spent potliners and pot
skimmings are stored en the ground before cryolite recovery either on site
or at an outside location.
Sludges. Sludges originating from lime treatment of pot room
emissions scrubwater, carbon plant scrubwater,and potliner storage pile runoff
are settled in unlined lagoons and may or not be periodically dredged with
permanent dispostion in open dumps.
Dusts. Dust from shot blasting and skim processing is currently
open dumped. .
As discussed previously all of the above wastes are considered
potentially hazardous. Levels of technology for their treatment and disposal
and evaluation of environmental adequacy thereof are discussed in the next
sections.
4.3.2 Present Treatment and Disposal Technology (Level I)
Potliners and Pot Skimmings. Spent potliners and pot skimmings
are stored on the ground for months to years before cryolite recovery
either on site or at an outside location. This practice is inadequate because
of possible fluoride and cyanide leaching.
Sludges. For the hypothetical plant, scrubber water from the
potrooms is combined with scrubber water from the carbon plant and runoff
from the potliners storage pile. The combined stream is then treated
with lime and pumped to an unlined lagoon. The lagoon deposits primarily
contain carbon particles, alumina, fluorine and calcium (from the lime
treatment). Most of the fluoride and calcium will occur as calcium fluoride.
Periodically, the bottom deposits in the lagoon may be pumped out or clammed
out in order to allow continued use of the lagoon. At some plants sludge is
left permanently in lagoons and new lagoons built as lagoons are filled to
capacity. These deposits amounting to some 117 kg/MT aluminum are disposed
of on land, probably adjacent to the lugoon. When the material is first
pumped it is thick and watery. liventually it dries out into a firm black
mass. Level I Technology for disposal of lagoon dredgings is disposal in open
dumps or permanent disposal in lagoons.
Many of the primary aluminum plants are using the above methods or
variations thereof for treatment and disposition of process and pollution
control residuals. Fluorides and cyanides could present environmental
threats if they leach through permeable soils to groundwater or are carried
in overland water runoff to surface streams.
117
-------
Dusts. Dust from shot blasting(rodding room) and skim processing
(cast house) is open dumped. As discussed previously shot blast dust did not
leach in solubility tests and open dumping is considered adequate. Cast
house dust is suspected of leaching and open dumping therefore is not con-
sidered adequate.
4.3.3 Best Technology Currently Employed (Level II)
Pot Liners and Pot Skimmings. A segment of the primary aluminum
industry immediately processes spent potliners and skimmings for cryolite
recovery without storing on the ground for any significant period of time.
This practice will preclude any leaching or runoff of fluorides or cyanides
and is therefore considered as Level II technology. It is noted, however,
that there is a trend in the industry toward dry-type cell emission control
systems. In plants having such systems, sodium fluoride emitted by the
cells is intercepted by the emission control system, along with gases and
other particulates, and returned to the cells, As a result of this recovery
process, the loss of sodium fluoride is greatly reduced and recovery of
aluminum fluoride from spent potliners is of greater interest than the
recovery of cryolite. Thus cryolite could become in excess supply and be
of little or no value.
Sludges. At the present time no primary aluminum plants are
using lined or sealed sludge disposal areas. Therefore, Level II technology
is the same as Level I for containment of sludges from the carbon plant and
cryolite recovery. This technology may not be adequate for ground and
surface water protection if leaching or runoff of fluorides is significant.
Dry alumina absorption beds are used at some plants for the collection of
reduction cell emissions as previously described. The collected emissions
are recycled directly to the potlines thus precluding sludge generation
from pot lines.
[)u->jts. Open dumping of shot blast dust from anode plants and
skim proc
-------
Dusts. Sealing of ground areas where dust from skim processing
(cast house) is disposed is needed if trace metals (Cu, Zn, Pb, Cr, Ni)
leach through permeable soils to groundwater. Open dumping of shot blast
dust is considered adequate at present in view of insignificant leaching of
trace metals in solubility tests.
Tables 48 through 50 summarize features of Levels I, II and 111
treatment and disposal technologies for primary aluminum smelters including
assessment of the hazardous nature of residuals and the methods and adequacy
of treatment and disposal methods.
/\A COST ANALYSIS
In tho last hiH-tion various treatment and disposal technologies
currently employed or coiusidorod for adoquate health and environmental pro-
tection wore described. The costs of i nip J omen ting this technology for
typical plants is considered in this section. Tho hypothetical plant is
assumed to have an annual capacity of 153,000 NfT of aluminum per year and
operates 365 days/year.
4.4.1 Cost of Present Treatment and Disposal Technology (Level I)
Scrubber liquor for potrooms and the carbon plant together with
run-off water from the potliner storage pile and other water are lime treated
and put into a lagoon. The daily inflow is 105,980 m . Lagoon character-
istics to accommodate this flow for adequate retention time are:
Volume 4 J.r., 00(1 nf"1
ttcttoin width -'62 m
Top width J7-I m
Bottom length 523 m
Top length 535 m
Total depth 3 m
Depth of excavation .25 m
Circumference 1,630 m
Dike volume 38,088 m
Dike surface 25,591
Total width 291 m
Total length 552 ha
Required area 16 ha
Eight test holes are drilled and 16 soil samples analyzed. The lagoon is
situated on rural land.
Approximately 42 MT of solids a day or 17,900 MT/yr flow into the
lagoon daily. These form approximately 113 m of sludge. The sludge is
removed by pumping. A 2HP slurry pump with a capacity of 236 1/min is
employed. The sludge is pumped through a 7.6 cm pipe to a sludge dump.
The flow rate is 0.85 m/sec. The sludge dump is located some distance from
the lagoon. Pipe costs include 200 m of installed pipe and 300 m of flexible
pipe. Pumping operations arc conducted on the average of 9.5 hr/day.
119
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About 41,390 m of sludge are generated annually which amounts to
827,800 m over 20 years. The sludge, when dried, reduces in volume by 40
per cent, thus requiring storage for about 500,000 m of waste. The area
requirement is 5 ha assuming sludge build-up to a height of 10 m. The area
is initially excavated to a depth of .5 m to prevent excessive run-off.
Twelve hours of labor time/day is assigned to the pumping operation.
Solid waste consisting of discarded potliners is generated at a
rate of 22 Ml/day or 8030 MT/yr based on 52.5 Kg/NT of aluminum produced.
The potliners have an estimated density of 2404 Kg/M . The potliners are
eventually recycled, however, the average amount stored represents 3 years
of production which amounts to about 10,000 m (24,000 MT) of waste. This
requires an area of about .25 ha for storage assuming the potliners are
piled to a height of 5 m.
One hour/day of frontleader and truck time is allocated for
transporting the potliners to the dump.
Other sources of solid waste are dusts from shot blasting and
processing of cast house skim. Approximately 540 NTT of shot blasting dust
and 330 MT of skim processing dust are,produced annually. The densities of
the dusts are 1285 Kg/m and 1045 Kg/m , respectively. This represents
870 m of waste which are disposed on land each year. Loading and trans-
porting this material requires 48 hours of frontloader and 48 hours of truck
time each year. An area measuring about 0.2 ha is needed to store 20 years
of waste given that the waste is piled to a 10 m height.
Capital and annualized costs of Level I treatment and disposal are
summarized in Table 51. The annual cost per metric ton of aluminum
produced is estimated as $0.62.
4.4.2 Cost of Best Technology Currently Employed (Level II)
The use of lined lagoons is not presently practiced in the industry
and thus is not costed as Level II technology. Level II sludge and dust
disposal is the same as Level 1.
Potliners are recycled within a few weeks. This eliminates the
need for the potliner dump and reduces the proportionate use of the truck
from 20 to 2 1/2 percent. A 20 x 20 m concrete pad is provided for temporary
storage of 2 - 4 days accumulation of used potliners. The resultant incre-
mental costs are;
Capital Cost
Concrete pad $ 6,140
Potliner dump (595)
Dump truck (4,375)
Total $ 1,170
126
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TABLE 51
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ALUMINUM PLANT
Capital Cost
Sludge Potliners Dust
Lagoon
Site Preparation
Survey $10,000
Test Drilling 1,970
Sample Testing 1,000
Report Preparation 2,400
Construction
Excavation S Forming 50,655
Compacting 70,465
Fine Grading 11,515
Soil Poisoning 2,020
Transverse Drain Fields 24,095
Land 28,000
Dump
Survey 3,125 $ 155 $ 125
Land 8,750 440 350
Excavation *i Forming 33,250
Equipment
Slurry Pump 13,730
Pipe, Rigid (installed) 4,425
Pipe, Flexible 1,320
Dump Truck (20%) 4,300 700
Frontloader (20%) 3,440 560
Total $266,720 $8,335 $1,735
Annual Cost
Land $3,675 45 35
Amortization
Construction 24,415 20 15
Equipment 3,095 1,230 200
Operating Personnel 42,130 7,885 1,035
Repair and Maintenance
Construction 4,765
Equipment 975 385 65
Energy
Fuel - 1,000 130
Electricity 80 100 15
Taxes 920 10 10
Insurance 2,665 85 15
Total $82,720 $10,760 $1,520
127
-------
Annual Cost
Land $ (45)
Construction Amortization 645
Equipment Amortization (695)
Operating personnel
Construction maintenance
and repair 165
Equipment maintenance
and repair (220)
Energy
Fuel (585)
Taxes (10)
Insurance 10
Total (735)
4.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III).
In addition to immediate recycling of potliners, the lagoon is
lined, the ground is sealed at the sludge and dust disposal areas, and
collection ditches, pumps and piping installed. Costs incurred are
shown in Table 52.
Costs of Levels I, II and III treatment and disposal technologies
are summarized in Table 53 for sludge, potliners and dust residuals from
primary aluminum smelting and refining. Costs are given in dollars per
metric ton of dry and wet waste and dollars per metric ton of aluminum
product. Costs for each type of waste and each level of technology are
also expressed as a percent of metal selling price.
Total industry costs for Levels, I, II and III treatment and
disposal technologies have been estimated in 1973 dollars and are presented
in Table 53. The annualized industry cost for Level I and II treatment
and disposal technologies is estimated as 2,550,000 or 0.09% of national
sales.
The annualized industry cost for Levels III which is adequate
for environmental protection is $5,720,000 or 0.211 of estimated national
sales. The large increase in cost from Level IT to II is a result of using
lined luyoons for scrubber sludge.
128
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TABLE 52
COST OF LEVEL III TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ALUMINUM PLANT
Capital Cost
Sludge Potliners Dust
Construction
Lagoon Liner $609,795
Land Sealing 100,000 - $4,000
Collection Ditches I/ 9,360 - 355
Equipment
Pump and Piping 2J 14,625 - 585
Total $733.780 - $4,940
Annual Cost
Land - - -
Amortization $
Construction $ 83,420 - $ 505
Equipment 2,325 - 95
Operating Personnel -
Repair and Maintenance
Construction 21,575 - 130
Equipment 730 - 30
Energy
Fuel
Electricity 250
Taxes
Insurance 7,340
Total $115,640
JV Ditches are installed around perimeter of sludge dump and 1.5 x 1 side
of dust disposal area.
2J Common pump and piping is used for run-off from sludge and dust dump.
129
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5.0 PRIMARY SMELTING AND REFINING OF ANTIMONY (SIC 3339)
5.1 INDUSTRY CHARACTERIZATION
The primary antimony industry is relatively small. Production of
antimony from domestic smelters in 1972 was 12,106 MT of which only 3,480 MT
was antimony metal (Ref. 1 ). The remainder consisted of oxides and sulfides
of antimony byproduct, antimonial lead, and other antimony residues. The
selling price of antimony metal in 1973 was given as 59 cents/lb. Thus
national sales are estimated at $4,528,000 (Ref. 2).
Plants producing antimony metal are the NL Industries, Inc. smelter
at Laredo, Texas, the Sunshine Mining Co. electrolytic plant at Big Creek,
Idaho, and the U.S. Antimony Corporation plant at Thompson Falls, Montana.
Some antimony is recovered as a lead smelter byproduct (~7% of production
in 1972) at three primary lead smelters. Over 90% of byproduct antimony
produced at primary lead smelters was consumed at the smelter in manufactur-
ing antimonial lead.
Approximately 80% of domestic antimony is produced at the Laredo,
Texas smelter. Most of the remainder is produced electrolytically in Idaho.
The U.S. Antimony Corporation produced 142 tons in 1972 (Ref. 1 ). Table 54
gives the geographical distribution of primary antimony plants by state,
region, and nationally, according to process type.
The methods selected for the production of antimony metal or oxide
are largely determined by the grade of the ore or concentrates. The lowest
grade ores or concentrates (5 to 25 percent antimony) are generally treated
by volatilization processes; those containing 25 to 40 percent antimony
are smelted in a blast furnace. High-grade ores containing 45 to 60 percent
antimony may be treated by a process called liquation, in which the antimony
is reduced, melted, and drained from the gangue material. Special concen-
trates, such as silver-antimony concentrate, are treated by leaching and
electrowinning. Condensed oxides from volatilization processes, liquated
antimony sulfide, and some intermediate-grade ores are smelted in reverbera-
tory furnaces.
5.2 WASTE CHARACTERIZATION
5.2.1 Process Descriptions
Antimony metal is produced by pyrometallurgical and electrolytic
processes as described below.
Py romet a11urgica1
The operation at Laredo, Texas (Flff. 8") utilizies blast furnace smelt-
inc. The charee to the blast furnace consists of mixed oxide and sulfide ores
from mines in Mexico or the United States, and byproducts of other smelting
operations: mattes, slags, flue dusts, or residues from lead and zinc
132
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MIXED OXIDE
ANDSULFIDE
ORES
DUST RECYCLE-
BAG
HOUSE
DUST
1
AIR
CHARCOAL &
FLUX
BLAST
FURNACES
COKE, FLUX
DUST
IMPURE
ANTIMONY
REVERBATORY
FURNACE
SLAG
RECYCLE
PURE ANTIMONY
NUMERICAL VALUE IS IN KILOGRAM/MT
OF ANTIMONY PRODUCED
Figure 8 PYROMETALLURGICAL ANTIMONY SMELTING AND REFINING
134
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refineries. The blast-furnace operation is somewhat similar to that of lead
blast-furnace practice. A vertical or shaft furnace with a rectangular cross
section is used, with low-pressure air being supplied. The ore and residues,
charcoal, and fluxing additions are charged at the top of the furnace and
the products of molten antimony metals and slag are tapped from the hearth
at the bottom of the furnace. The blast-furnace process is capable of pro-
ducing extremely pure metal with minimal antimony loss.
Elect rob/tic
Electrolytic primary antimony metal is produced by a leaching-
electrolysis process. The process is depicted in Figure 9 . In this process,
a complex copper-antimony sulfide concentrate is leached with sodium sulfide
solution which dissolves the antimony as sodium thioantimonate (Na3SbS4).
The leach solution is clarified by settling and filtration, and electrolyzed
in diaphragm cells to yield antimony of 9.3 to 99 percent purity. During the
electrolysis, some sulfide sulfur is oxidized to thiosulfate or sulfate.
With recirculation of the electrolyte, the concentration of these oxidized
forms of sulfur would build up to the point where they would interfere with
the solubilization of antimony in the leaching step. By treating spent
electrolyte with barium sulfide, these sulfates are kept at acceptable levels;
and the sulfide concentration is maintained. The barium sulfate and thio-
sulfate precipitate is filtered and heated with coal in a rotary kiln to
reconstitute barium sulfide for reuse.
5.2.2 Description of Waste Streams
Pyrometallurgi cal
The only waste produced in the pyrometallurgical process for
antimony metal production is a hard siliceous blast furnace slag produced
at a rate of 2,800 Kg/NTT of antimony metal. All other byproducts are recycled
to the blast furnace. Analysis of a blast furnace slag from an antimony
smelter is given in Appendix A . Antimony and zinc are present in high con-
centrations (i.e., Sb -18,000 ppm; Zn - 500 ppm). Leaching of trace metals
from the lurgo chunks of hard siliceous slag would not be anticipated, llow-
ovor, in solubility tests as described in Appendix B arsenic, antimony,
copper, and zinc were found to leach in significant concentrations. Blast
furnace slag is therefore considered potentially hazardous at this time.
Hloctrolytic
The only discard waste from electrolytic antimony production is
spent anolyte solution which is discharged to a tailings pond which receives
wastes from the mining and milling operations also. Approximately 13m3 of
anolyte solution containing 210 Kg/NfT of antimony metal is produced. The tail-
ings pond will accumulate 190 MT of dry solids per year from the electrolytic-
operation.
Analyses of solids disposed of in tailings ponds is given in
135
-------
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Appendix A . Solids contained traces of Pb, Cu, Zn, Ni, Mn, Cr, As and Cd.
Antimony was the highest concentration (27,000 ppn). In solubility tests de-
scribed in Appendix B, antimony, copper, cadmium and lead were found to leach
to a minor extent. For this reason along with the very fine particulate nature
of the sludge exposed to weathering processes, electrolytic sludge from primary
.-intimoiiy plants is considered potentially hazardous.
5.2.3 WAST!! QUANT! I'll S
Table 55 summarizes the generation factors for each of the land
disposed residuals in the primary antimony melting and refining category (i.e.,
blast furnace slag, electrolytic sludge). This table also gives typical con-
centration of potentially hazardous constituents as derived from the analysis
of blast furnace slag and spent anolyte solids given in Appendix A . Table 56
gjyes the yearly quantity of wastes and potentially hazardous constituents
for a typical pyrometallurgic.il plant producing 2900 MT/year of antimony metal.
Table 56 gives yearly wastes for a typical plant producing 900 MT/year of anti-
mony metal by the elctrolylic process.
Using the waste generation ("actors and concentrations contained in
Tables 55 and plant capacities, estimates have been made for the quanti-
ties of land disposed residuals and ha/.ardous constituents thereof. These es-
timates are given for 1974, 1977 and 1983 on a state-by-state-regional and
national level in Table 57.
Primary antimony metal production has fluctuated from a low of
2594 MT in 1973 to a. high of 3462 MT over the period of 1969 to 1973 (Ref. 2).
Thus, no expansion in the industry can be predicted over the period 1974 to
1983 with waste generation based on a total original production of 2900 MT for
each of the predictive years [1974, 1977, 1983). Texas is the largest producer
of primary antimony and generates the largest quantity of potentially hazardous
waste.
5.3 TREATMENT AND DISPOSAL TECHNOLOGY
S. ^. 1 Current TrcaTineiit and jji .sposa] Technology
At the current time discard slag from smelting furnaces of pyro-
metallurgica1 plants is open dumped and spent anolyte solution from the elec-
trolytic plant is discharged to the tailings pond. Since both of these wastes
are considered potentially hazardous levels of technology for their treatment
and disposal are discussed in more detail in the following sections.
5.3.2 Present Treatment and Disposal Technology (Level I)
Blast furnace/reverberatory slag. At the present time discard slags
from reverberatory and/or blast furnaces from both pyrometallurgical plants
are open dumped on site. This practice is inadequate if heavy metals leach
through permeable soils to groundwater.
137
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Spent Anolyte Solution. At the present time spent anolyte solution
from the one electrolytic antimony plant is discharged to the on-site unlined
mill tailings pond. This practice is inadequate if heavy metals leach through
permeable soils to groundwater.
5.3.3 Best Technology Currently EmpJLo/ed (Level II)
Blast Furnace/Reverberator:) Slag. The practice and adequacy of Level
II slag disposal is the same as Level I, since open dumping is the only prac-
tice employed.
Spent Anolyte Solution. Tue practice and adequacy of Level II spent
anolyte solution disposal is the same as Level I since discharge to unlined
tailings ponds is the only practice empJoyed.
5.3.4 Technology Necessary To Provide Adequate Health and Environmental
Protection (Level IIIj _„__
Blast Furnace/Reverberator/ Slag, In the event that leaching of heavy
metals through permeable soils to grcundwater occurs (i.e., Sb, Zn) sealing
of ground at slag disposal areas with collection and treatment of runoff will
be required to provide adequate health and environmental protection.
Spent Anolyte Solution. Sediments and water discharged to the tail-
ings pond from the electrolytic antimony refining contain appreciable concen-
trations of antimony, sulfide, thallium, tellurium, and other potentially
hazardous heavy metals. If significant leaching of contained hazardous metals
from the fine sediments occurs through permeable soils to water tables a sep-
arate lined lagoon should be constructed for permanent containment of sludges
as Level III technology. Only treated water would be discharged to> the main
tailings pond. Tables 58 and 59 summarize features of Levels I, II and III
treatment and disposal technology for pyrometallurgical and electrolytic re-
fining of antimony including assessment of the hazardous nature of residuals
and the methods and adequacy of treatment and disposal methods.
5.4 COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental protec-
tion were described. The costs of implementing this technology for typical
plants is considered in this section. Pyrometallurgical plants and electro-
lytic plants will be considered separately because of differences in production
technology and types and quantities of land disposed wastes.
5.4.1 Cost of Present Treatment and Disposal Technology ("Level II
Pyrometallurgical Plant. The typical plant produces 2,700 MT of
metal annually operating 365 days/year. In the process, 7,500 MT of slag are
generated. The slag is deposited on an on-site dump.
One hour/day of front loader and backhoe time are assigned for
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hauling the slag to the dump. A dump area of about .9 ha is required to store
slag produced over a 20 year period, assuming it is piled to a 10 m height.
The dump is located on rural land and 45 hours/year of bulldozer time are
assigned at the dump for spreading and compacting. The costs of Level I treat-
ment and disposal technology are given in Table 60.
Electrolytic Plant. Annual production is 900 NfT based on 365 days of
operation.
Waste discharge consists of 13.2 m^ of waste water daily containing
193 NfT of solids per year which is discharged into a tailings pond. The tail-
ings pond is associated with a mine-mill complex. The refinery waste repre-
sents less than 1 percent of the total discharge into the tailings pond. Con-
sequently, no cost is assigned to the refinery discharge.
5-4.2 Cost of Best Technology Currently Employed (Level II)
Pyrometallurgical Plant. The technology and costs of Level II treat-
ment and disposal of blast furnace slag, the only land disposed waste, is the
same as Level I.
Electrolytic Plant. The technology and costs of Level II treatment
and disposal of spent anolyte, the only land disposed waste, is the same as
Level I.
5.4.3 Cost of Technology to Provide Adequate Health § Environmental
Protection
Pyrometallurgical Plant. The land is sealed at the slag dump;
collection ditches and pump and piping are installed. Additional costs in-
curred are given in Table 61 . Additional costs for this level amount to
approximately one half of present costs (i.e., Level III costs are 1 1/2 times
present costs). Since the amount of slag is large relative to the amount of
product, costs are less per ton of waste than product produced.
Electrolytic Plant. The waste water from the .electrolytic plant
(spent anolyte) is directed to a lined settling pond before the overflow enters
the tailings pond. About 528 Kg/day of dry solids are discharged into lagoon
where they form .32 m^ of sludge. The yearly sludge accumulation is 116.8 m-*.
The lagoon is designed to retain 20 years accumulation of sludge,
fi.e. 2336 m3), containing 3860 NTT of dry solids or Ml,000 MT of wet sludge
a lagoon is generally effective until it is one half filled with sludge. Thus,
the selected lagoon size to hold the volume of sludge generated in 20 years
is 4,800 m . Its design characteristics are:
146
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TABLE ft()
COST OF LEVEL 1 TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ANTIMONY 1>LANT - PYROMETALLURGICAL REFINING
Capital Cost
Reverbatory slag
Equipment
Dump Truck (20%) $ 5,000
Front Loader (20-n) 4,000
Bulldozer (3%) 480
Dumj)
Survey 560
Land 1,580
Total $11,620
Annual Cost
Land $ 160
Amortization
Construction 65
Equipment 1,510
Operating Personnel 8,430
Repair and Maintenance
Construction
Equipment 475
Energy
Fuel 1,040
Electricity 105
Taxes 40
Insurance 115
Total $11,940
147
-------
TABLE hi INCREMENTAL COSTS FOR LAND SEALING AND RUNOFF COLLECTION FOR SLAG
DISPOSAL, PRIMARY ANTIMONY PRODUCTION
Capital Costs
Land Sealing $18,000
Collection Ditches 1,435
Pump § Piping 10,010
Total $29.445
Annual Cost
Construction Amortization $ 2,255
Equipment Amortization 1,595
Construction Maintenance 5 Repair 585
Equipment Maintenance § Repair 500
Energy
Electricity 50
Insurance 295
Total $ 5,280
148
-------
Bottom width 24 m Circumference 201 m
Top width 36 m Dike Volume 1,900 m3
Bottom length 47 m Dike Surface 2,615 m^
Top length 59 m Total Width 48 m
Total depth 3 m Total Length 71 m
Depth of Excavation 1.45 m Required Area .35 ha
Costs are shown in Table 62.. Since a new separate lined lagoon would be required
for anolyte,all costs would be incremental.
Costs of Levels I, II and III treatment and disposal technologies for
pyrometallurgical and electrolytic plants are summarized in Tables 63 and 64
Costs are given in dollars per metric ton of dry and wet waste and dollars per
metric ton of antimony product. Costs for each level of technology are also
expressed as a percent of metal selling price.
Total industry costs for Levels I, II, and III treatment and disposal
technologies have been estimated in 1973 dollars and are presented in Table 63
for pyrometallurgioal antimony production and Table 64 for electrolytic antimony
production. The annualizcd industry cost for Level I and II treatment and dis-
posal of pyrometallurgical antimony waste slag is $10,000 or 0.29% of national
sales value. The annualized cost of Level III technology (i.e., adequate for
environmental protection) is 10,200 or 0.42% of national sales value.
Since sludge from electrolytic antimony production comprises approxi-
mately only one percent of a larger waste stream from a mine-mill complex,
costs under Levels I and II treatment and disposal technology are negligible.
The settling of the sludge in a separate lined lagoon before discharge to the
main tailings pond results in an industry capital investment of - $10,000 and
2,000 annualized costs which represents 0.25% of national sales.
149
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TABLE 62
COST OF LEVEL III TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY ANTIMONY PLANT - ELECTROLYTIC REFINING
Capital Cost
Spent Anolyte
Sludge
Lagoon
Site Preparation
Survey $ 220
Test Drilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation $ Forming 2,525
Compacting 3,515
Fine Grading 1,175
Soil Poisoning 250
Transverse Drain Fields 440
Lagoon Liner 11,190
Land 615
Total $21,870
Annual Cost
Land $ 60
Amortization
Construction 2,470
Equipment
Operating Personnel
Repair and Maintenance
Construction 575
Equipment
Energy
Fuel
Electricity
Taxes 15
Insurance 220
Total $ 3,340
150
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Ci.O t'KlMAKY SMll.TlNi. -V\l> KUININU OS MUU UK\
(».l INDUSTRY CHARACn RIZMION
'ihu primary mercury smelting ami refining industry has been centered
i/i California, Nevada, and Oregon. Mercury has also been recovered from ore
in Arizona, Alaska, Idaho, Texas, and Washington and is recovered as a
byproduct from gold ore in Nevada and zinc ore in New York. Due to low prices
and slackened demand the mercury industry has been in a decline during recent
years. Table 65 illustrates the decline in the industry (Ref. 3). In 1969
there were 109 producing mines. In 1973 there were only 6 producing mines.
It has been predicted (Kef. 3j, given such variables as market prices and
effects of emission standards promulgated by EPA in April 1973, that production
of primary mercury could range from a high of 20,000 flasks to a low of 3,000
flasks by 1985 but is expected to remain steady at about 3,000 flasks through
1985. Table 66 shows the geographical distribution of primary mercury
production capacity as of late 1973. It must be kept in mind that operations
open and close on short notice and resultant production can fluctuate greatly.
U.S. sales of primary mercury was given as $1,601,000 in 1972 (Ref. 1).
TAB LI- 65. SALIENT DOMESTIC MERCURY STATISTICS
1969 1970 1971 1972 1975
Producing Mines 109 79 56 21 6
Production-flasks* 29,640 27,296 17,883 7,286 2,200
* 1 flask - .M.47 kg (70 lb)
Mercury is extracted from ore and concentrate by henting in retorts or
furnaces to liberate the metal as a vapor, followed by cooling of the
vapor and collection of the condensed mercury. Recovery of mercury is
high, averaging about 98 percent for retort installations and 95 percent
for furnace plants.
6.2 WASTE CHARACTERIZATION
This section contains descriptions of production technology at
primary mercury smelters and the resultant byproducts or wastes which are
either recycled directly or disposed of on land. Estimates are given for
the quantities of wastes and potentially hazardous constituents thereof
which are disposed of on land.
6.2.1 Process DCScription
As discussed previously mercury is normally recovered directly
from ores either in retorts or furnaces. Retorts are inexpensive installa-
tions for small operations and require only simple firing and condensing
equipment. They are best adapted to operations treating 500 pounds to 5 tons
per day of high-grade sorted ore.
153
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For larger operations, either rotary or multiple-hearth furnaces
with mechanical feeding and discharging devices are preferred. Furnace
capacities range from 10 to 100 MT per day with larger sizes to meet
requirements.
Mercury-laden gases pass from the furnaces through dust collectors
into the condensers, where the vapor is cooled and the mercury collected.
Final traces of mercury in the gases from the condensers are removed in
washers, and the stripped gases are discharged through a stack into the
atmosphere. The mercury from the condenser always contains some dirt and
soot, which are separated by hoeing the mixture with lime.
Mercury can also be leached from its ores and concentrates with a
solution of sodium sulfide and sodium hydroxide and can be recovered as
the metal by precipitation with aluminum or by electrolysis. Leaching of
mercury ores has not been practiced extensively in the past. However use
of these methods may increase in the future.
6.2.2 Description of Waste Streams
Calcine Residue. Figure 10 shows the various operations in
pyrometallurgical recovery of mercury using retorts or furnaces along with
the evolution and disposition of wastes. The largest source of waste is the
furnace or retort residue amounting to some 207 MT per metric ton of metallic
mercury produced. In solubility tests described in Appendix B leaching of
toxic trace metals from furnace residue was insignificant (less than 0.5 ppm).
Furnace or retort, residues (i.e. calcine residues) are not considered poten-
tially hazardous, therefore, at this time.
Condenser Cooling and Washing Water. There is a small volume of
water (43 m /MT metal) discharged from condenser cooling and wash-
down. This amounts to about 5 m /day. Samples or analyses of condenser waste-
water were not available but traces of free mercury are expected. This waste-
water is therefore considered potentially hazardous.
As noted in Figure 10 dirty mercury may be filtered or cleaned to
obtain a purer grade. Triple distillation may even be done for purification.
In some cases nitric acid may be used for cleaning. In such cases the small
laboratory scale volume of waste acid will contain soluble mercurous nitrate
and should not be disposed on on land or in lagoons. All traces of mercury
should be removed from the acid by distillation.
The typical plant representing the most prevalent plant size within
the industry produces 500 flasks of mercury per year (76 Ib/flask) or 17.3 MT
per year. The typical plant operates about 150 days per year. Daily
production of mercury is therefore about 0.115 MT of mercury. Daily discharge
of i-oiulririor w.'itiT I •: :ipprox imutoly '1. i> m' to a Injjoon or the slag pile. There is
u culcino rosldu^ of iipproxlnuit oly 24 MT per day. Yearly discharge of water to
155
-------
CRUSHED ORE
GASES
TO STACK
99.9% MERCURY
Figure 10 PRIMARY MERCURY PRODUCTION
156
-------
unlined lagoons is approximately 147,000 rn . About 3,580 MT of calcine
residue is open dumped per year from the typical plant. Two samples of
calcine waste were obtained for chemical analyses. Results of analyses
are given in Appendix A.
6.2.3 Waste Quantities
Table 67 summarizes the generation factors for each of the
land disposed residuals in the primary mercury smelting and refining
category (i.e. calcine residue condenser water). This table also gives
the typical concentrations of potentially hazardous consitutents in a
furnace calcine residue based mainly on the average analyses of two
furnace calcine residues given in Appendix A. It is reported that
mercury recovery from ore is about ->5% using furnaces and 98% using
retorts because of the higher efficiency of retorts. Calcine waste can
vary widely in mercury content. A typical concentration of 200 p'pm
mercury in furnace or retort residue was used based on 98% recovery from
a l"o ore.
Samples of solids from condenser wastewater and analyses
thereof were not available, and therefore concentration factors could
not be developed.
Using the waste generation factor and concentration factors
for calcine residue contained in Table 67 estimates have been made for
the quantity of land disposed residue and hazardous constituents thereof
for a typical plant producing 500 flasks or 17.3 MT of mercury per year.
These estimates are given in Table 68.
Table 69 gives state-by-state, regional and national quantities
of land disposed residuals from mercury smelting based on production
capacities and generation and concentration factors contained in Table
67. Estimates are given for 1974, 1977 and 1983. Estimates were based
on the assumption that yearly national production by traditional methods
will remain level at approximately 3,000 flasks through 1985 as forecast
by U.S. Bureau of Mines.
6.3 Current Treatment and Disposal Practices
Caly inc. Currently calcine from retorts or furnaces is open
dumped. The greatest percentage of the primary mercury industry is
located In arid or semi-arid regions. In addition insignificant leaching
of trace metals was observed in solubility tests described in Appendix
B. There appears to be little chance of groundwater contamination from
calcine residue. Open dumping is considered environmentally adequate at
this time and will not be discussed further.
157
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Condenser Water. Wastewater from condenser cooling or wash-
down is either sent to unlined lagoons or spread in the calcine dump. This
waste is considered potentially hazardous due to the suspected presence of
free mercury. Levels of technology for this wastewater are discussed in the
following sections.
6.3.2 Present Treatment and Disposal Technology (Level I)
The practice of disposing of condenser cooling and wash water in
unlined lugoons or in calcine dumps is not considered environmentally
adequate.
6.3.3
Best Technology Currently Employed (Level II)
Level II treatment and disposal technology and adequacy thereof
are the same as Level I.
6.3.4 Technology To Provide Adequate Health and Environmental
Protection (Level III)
Lined lagoons should be used to store condenser wastewater and
any other process wastewater.
Table 70 summarizes features of Levels I, II and III treat-
ment and disposal technologies including assessment of the hazardous nature
of residuals and the methods and adequacy of treatment and disposal methods.
6.4
COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental pro-
tection were described. The costs of implementing this technology for
typical mercury furnaces or retorting operations is considered in this
section. The typical plant is assumed to produce 500 flasks (17.3 MT) of
metallic mercury per year and operates 150 days/year.
6.4.1
Cost of Present Treatment and Disposal Technology (Level I)
Condenser cooling and washdown water which may contain traces
of metallic mercury flows into a lagoon. The flow is at a daily rate of
The lagoon is designed to hold 20 years of effluent. Its character-
istics are:
Volume
Bottom Width
Top Width
Bottom Length
Top Length
Circumference
2,000 m
19.4 m
27.4 m
38.8 m
46.8 m
160.4 m
3
Depth of Excavation 0.95 m_
Dike Volume 858 m-
Dike Surface 2,426 m
Total Width 37.6 m
Total Length 57.0 m
Required Area 0.21 ha
160
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Practically all mercury plants arc located in arid regions. The
sizing of the lagoon assumes a net annual evaporation rate of about 63 cm/m
of surface area.
Capital and operating costs are shown in Table 71.
(>.<1.2 Cost, of Bc-.i 'ri'i-luio_logy_ Current l_y_ hni^loy_qd (Level II)
The technology and costs of Level 1J technology are the same as
for Level 1.
6.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III)
The lagoon is lined. Associated capital and annual costs are
shown below.
Capital Cost Slurry
Lagoon Liner $ 7,375
Total $ J7.375
Annual Costs
Construction Amortization $ 855
Construction Repair and
Maintenance 220
Insurance 75
Total $ 1,150
Costs of Levels I, II and 111 treatment and disposal technologies
for condenser wastewater slurries are summarized in Table 72. Costs are
given in dollars per metric ton of dry and wet waste and dollars per metric
ton of mercury product. Costs for each level of technology are also expressed
as a percent of metal selling price.
Total industry costs for Levels I, II and III treatment and disposal
technologies have been estimated in 1973 dollars and are presented in
Table 72. The annualized industry cost for Level I and II treatment and disposal
technology is 40,000 or 0.7?o of estimated 1973 national sales. The
annualized cost of Level III technology for the industry is $210,000 as a
result of the need for lined lagoons for adequate environmental protection.
This represents 1.5"<; of estimated 1973 national sales value.
163
-------
Table 71
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY MERCURY PLANT
Capital Cost
Lagoon
Site Preparation
Survey $ 130
Test Drilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation & Forming 1,140
Compacting 1,585
Fine Grading 1,090
Soil Poisoning 200
Transverse Drain Fields 345
Land 370
TOTAL $ 6,800
Annual Cost
Land $ 35
Amortization
Construction 745
Equipment
Operating Personnel
Repair and Maintenance
Construction 130
Equipment
Energy
Fuel
Electricity
Taxes 10
Insurance 70
TOTAL $ 990
164
-------
Table 72
COST SUMMARY FOR PRIMARY MERCURY SMELTING AND REFINING
Annual Production: Model Plant 17.3 MT Industry 74.0 MT
Waste (type)
Slurry
Amount (MT/MT of
44
production)
Cumulative Unit Waste Disposal Costs:
Waste (type)
Slurry
Capital Cost
Annual Cost
Total
I
$/MT of $/MT of
Waste Prod.
Wet
8.93 393.06
1.30 57.25
Capital Cost $3.93.06
Annual Cost 57.25
Level
II
$/MT of $/MT of
Waste Prod.
Wet
8.93 393.06
1.30 57.25
$393.06
57.25
III
$/MT of
Waste
Wet
18.62
2.81
$/MT of
Prod.
819.36
123.69
$819.36
123.69
Cumulative Industry Waste- Disposal Costs ($ million):
Waste (type)
Slurry
Total
Metal Price:
Treatment Cost
Slurry
Total
I
Cap. Ann.
$0.03 $0.004
$0.03 $0.004
$8,396.08/MT
as a Percent of Metal
I
Cap. Ann.
4.7% 0.7%
4.71, 0.71
Level
II
Cap . Ann .
$0.03 $0.004
$0.03 $0.004
Selling Price:
Level
II
Cap . Ann .
4.7% 0.7%
4.7% 0.7°.
III
Cap.
$0.06
$0.06
III
Cap.
9.8%
9.8%
Ann.
$0.009
$0.009
Ann.
1.5%
1.5%
165
-------
7.0 PRIMARY SMELTING AND REFINING OF TITANIUM (SIC 3339)
7.1 INDUSTRY CHARACTERIZATION
There are only two primary titanium, (i.e., titanium sponge)
producers in the U.S. One is Reactive Metals, Niles, Ohio and the other
is Titanium Metals Corporation, Henderson, Nevada (jointly owned by
National Lead and Allegheny Ludlum). Total U.S. plant capacity is
approximately 15,230 MT of titanium sponge metal. State, regional and
national distribution of titanium sponge metal plants is given in Table
73. Approximately 9 plants in the United States take the sponge metal
and purify it to ingots in electric furnaces. There are virtually no
wastes associated with this operation. For this reason this section of
the report is restricted to sponge metal production. Domestic production
of titanium sponge is not published. The selling price of sponge in
1973 was given as $1.42/lb. Assuming that the above two plants were
operating at capacity, national sales are estimated as approximately
$47,800,000.
Production of titanium sponge consists of chlorination of
rutilc (Ti02) concentrates to produce titanium chloride gas followed by
reduction to titanium sponge.
7.2 WASTE CHARACTERIZATION
The typical plant produces approximately 7600 MT of titanium
sponge per year and operates 350 days per year. This is the average
production of the two plants.
7.2.1 Process Description
Feed to the plant is estimated at 13,670 MT/yr of rutile
(Ti02) concentrate containing 94% Ti02. The concentrates are treated
(See Figure 11) with chlorine gas to produce TiCl. gas which is then
condensed, reduced with magnesium metal and purified with aqua regia to
produce titanium sponge. The titanium sponge is then melted and purified
in an electric furnace to produce pure titanium ingot. Magnesium chloride
from the reduction process is electrolyzed to produce chlorine gas and
magnesium metal which are recycled as shown in Figure 11.
7.2.2 Description of Waste Streams
Chlorinator and Condenser Sludges. Impurities in the rutile
feed are mainly aluminum, columbium, iron, silicon, vanadium, and zirconium.
These elements plus some of the carbon, chlorine, and titanium present
in the chlorination mix and the condenser form sludges which are settled
in a pond. Sponge melting is carried out in a vacuum furnace which
produces no slag. Much of the chlorine used in chlorinating the rutile,
and the sodium or magnesium used in reducing TiCl. to metallic titanium
metal are recovered by electrolysis and recycled in the process.
166
-------
TABLE 73
UtiOGItAI'lllCAI, DISTRIBUTION OF PRIMARY TITANIUM
SPONGIi METAL PRODUCTION CAPACITY
State
No. Of Plant Distribution
Plants by Capacity (MT)
' 56SO 9550
Plant Distribution Capacity Per
by Process Process (MT)
CR* CR*
Ohio
Nevada
EPA Region
V
IX
National
1 1
1
1 1
1
2
1
1 1
1
1 1
2
5680
9550
5680
9550
15,230
* CR - Chlorination, reduction
167
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168
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Sludges from the chlorinator and the condenser are the only
significant solid waste. As shown in Figure 11 these sludges, amounting
to 330 Kg/NfT of titanium sponge (dry weight) or 2508 KfT/yr, along with a
small quantity of distillation effluent and spent acid from leaching of
titanium sponge is sent to impoundment ponds, which are dredged frequently.
The amount of combined sludge dredged from the settling pond is estimated
at 2520 MY/yr or 7.2 Ml/day (dry basis). This sludge is trucked to a
landfill site located within a few miles of the plant by a contractor.
It has been stated that chlorinator and condenser waste sludge is about
40 percent water soluble and that the water soluble portion contains
potentially hazardous chloride and chloride-oxide complexes of chromium,
titanium, vanadium and other heavy metals (Ref. 4). In addition, HC1
gas may be released to the atmosphere in an anaerobic environment.
These sludges are therefore considered potentially hazardous.
7.2.3 Waste Quantities
Table 74 gives the generation factors for the chlorination-
condenser sludges which are lagoon disposed. This table also gives the
typical concentrations of potentially hazardous constituents. Concentrations
of potentially hazardous constituents were obtained from chlorinator
sludge analyses contained in Bureau of Mines Report of Investigations
7221 (Ref. 4). Average analyses of 8 samples showed the following
percentages of chlorine and various metals:
Cl 18.7
Cr2°3 1'7
Ti02 17.4
V205 4.6
Zr02 4.7
Chloride and titanium are relatively non-toxic compared to many other
substances. However their very high concentrations and high solubility
present definite groundwater pollution hazards.
Table 75 gives the yearly generated quantity of sludge and poten-
tially hazardous constituents thereof for a typical plant producing 7600
MT of titanium sponge operating 350 days per year. Generation and
concentration factors contained in Table 74 were used in calculations for
Table 75.
Table 76 gives estimates for state-by-state, regional and national
sludge generation and hazardous constituents for 1974, 1977, and 1983.
The only two states generating hazardous waste from titanium sponge pro-
duction are Ohio and Nevada with Nevada generating twice as much as a
169
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result of larger plant capacity. There is no basis for estimating plant
capacity changes or changes in pollution control technology which would
change the quantities of land disposed waste in 1977 and 1983 as compared
to 1974 since production quantities of sponge metal from the two domestic
plants aro not published.
7..1 TREATMENT AND DISPOSAL TECHNOLOGY
7.3.1 Current Treatment and Disposal Practices
The only significant solid waste is the chlorinator-condenser
sludge. At the plant located in EPA Region IX, sludge is reacted with
water and settled in a pond. The settled sludge is clammed out and
hauled to a commercial landfill site. Daily cover of waste is practiced
at this site. At the plant located in Region V sludge is trucked by a
contract disposer to large lagoons constructed in highly impermeable
glacial till and clay underlain by shale. Since this waste is considered
potentially hazardous it is discussed in more detail in the following
sections.
7.3.2 Present Treatment and Disposal Technology
As discussed previously chlorinator-condenser sludge is disposed
of either in a commercial landfill or a industrial sludge lagoon. This
practice is inadequate if toxic constituents (i.e. chromium, vanadium,
others) are found to percolate through soils to underlying aquifers.
7.3.3 Best Technology Currently Employed (Level II)
At the present time, chlorination condenser sludge containing
minor amounts of solids from distillation and leaching is either put
into a daily covered landfill or into sludge lagoons. Either of these
practices will qualify as Level II technology. If significant movement
of the chloride salts of chromium, titanium, vanadium and others into
water tables occurs, then these practices could not be considered environ-
mentally adequate.
7.3.4 Technology to Provide Adequate Health and Environmental
Protection (Level III)
As discussed previously, it has been shown that chlorinator
waste sludge is highly water soluble and the water fraction contains
chloride salts and complexes of chromium, vanadium, titanium and other
heavy metals. For this reason, lined lagoons should be used for ultimate
disposal of these sludges.
Features of Levels 1, II and III treatment and disposal tech-
nologies including assessment of the hazardous nature of residuals and
the methods and adequacy of treatment and disposal methods are summarized
in Table 77.
172
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7.4 COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental
protection were described. The costs of implementing this technology for
typical plants is considered in this section. The representative plant is
assumed to have an annual production of 7,600 MT titanium sponge metal and
operates 350 days per year.
7.4.1 Cost of Present Treatment and Disposal Technology (Level I)
Land disposed solid waste consists of chlorinator sludge which
accumulates in settling pits. The daily flow is 1135 m containing 7.16 NTT
solids. The settling pits are dredged by an outside contractor and carried
by tank truck to an off-site lagoon. An average of 27.5 m of sludge are
hauled daily.
Two settling pits, each designed to accommodate a 1 day flow are
provided. Their dimensions are 30 x 8 x 5 m. The contractor tank trucks
are assumed equipped with slurry pumps. It is estimated that 2 hrs/day of
truck time are required for dredging and transporting the sludge.
The only capital costs incurred by the plant are the construction
of the settling pits. Annual costs are based on the contractor's construction
and operation costs. The latter are computed similarly as for on-site
operations and then increased by a factor of 2 to account for transportation,
general administrative expenses and profit. The outside contractor is
assumed to utilize a lagoon with a volume of 192,500 m , sufficient to con-
tain 20 years of sludge wastes.
The capital cost to the plant for excavating and forming the
settling pits is $3,190. Other costs are shown in Table 78.
TABLE 78
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
Primary titanium (Chlorinator and Condenser Sludges)
Capital Cost $ 3,190
Annual Costs
Construction Amortization 235
Insurance 30
Contractor Charges 45,560
Total $49,015
175
-------
7.4.2 Cost of Best Technology Currently Employed (Level II)
The technology and costs of Level II are the same as those for
Level I.
7.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III)
The lagoon used for potentially hazardous chlorination sludge
disposal is lined. This results in an additional annual cost of $89,500
to the plant.
Costs of Levels I, II and III Treatment and Disposal Technology
are summarized in Table 79 for sludge residuals from primary titanium
refining. Costs are given in dollars per metric ton of dry and wet
waste and dollars per metric ton of titanium product. Costs for each
level of technology are also expressed as a percent of metal selling
price.
Total industry costs for Levels I, II and III treatment and
disposal technolgoies have been estimated in 1973 dollars and are presented
in Table 79. The annualized industry costs for Levels I and II treatment
and disposal technology is estimated as $70,000 or 0.16% of estimated
1973 sales values. The annualized industry cost for Level III treatment
and disposal technology (i.e. adequate for environmental protection) is
estimated as $230,000 or 0.54% of estimated 1973 sales.
176
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8.0 PRIMARY REFINING OF TUNGSTEN (SIC 3339)
8.1 INDUSTRY CHARACTERIZATION
The tungsten producing and processing industry is small and
rather specialized compared with major nonferrous metals industries.
Production of metal powder was 3,404 MT in 1971, 4,322 MT in 1972 and
5,634 MT in 1973 (Ref. 2). Geographical distribution of production
capacity by state and EPA region is given in Table 80. The selling
price of tungsten metal powder varied from $4.50 to $6.70 per pound.
Using a price of $5.00 the national sales are estimated as $61,974,000
for 1973.
8.2 WASTE CHARACTERIZATION
From the limited information available, it is estimated that
the average production of tungsten metal at the 15 plants in the United
States which produce metal or the final precursor to metal (i.e., ammonium
paratungstate) is 300 MT/year. Assuming 350 days/year operation, daily
capacity is 0.86 MT. It is believed that wet (hydrometallurgical)
processing is universal in the industry, but with very substantial
variations in process details.
8.2.1 Process Description
A flow diagram showing tungsten production processes and
sources of land disposed waste is shown in Figure 12. Ore concentrates
containing 60-70 percent tungsten expressed as W0_ are treated with hot
HC1 to solubilize tungsten and precipitate impure tungstic acid. The
crude tungstic acid is leached with sodium hydroxide to produce a solution
of sodium tungstate which can he further purified by crystallization,
ro-dissolution, re-precipitation, final digestion in HC1, and reaction
with ammonia to form pure ammonium paratungstate (APT). The crystallized
APT is then reduced pyrometallurgically with hydrogen gas to form pure
tungsten metal powder.
8.2.2 Description of Waste Streams
Digestion Residue As shown in Figure 12, a solid digestion
residue (Filter cake) remaining from the initial digestion of the ore
constitutes the primary residue from the tungsten production process.
This material in the amount of approximately 15 MT/year (50 Kg per
metric ton of tungsten product) is stored on land (for many years) at
some plants and is stored in drums at other plants. It apparently
contains sufficient metal values to warrant eventual re-treatment.
Analysis of a sampled digestion residue is contained in Appendix A.
Copper is present in significant concentration (3.8%) and traces of zinc
and lead are also present. In solubility tests as described in Appendix
B copper, nickel and zinc were found to leach significantly (90, 60, and
1.5 ppm respectively). Digestion residue is therefore considered potentially
hazardous.
178
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EFFLUENTS(DILUTE
SLURRIES) 480,000 GPD
LAGOON
482 Kg
(SLUDGE)
ORE
CONCENTRATE
HYDROCHLORIC
ACID
•EFFLUENT-
D.QE8T.ON
TUNGSTIC ACID (SOLID)
NaOH LEACH
OH
1
SODIUM TUNGSTATE SOLUTION
CRYSTALLIZATION
SODIUM TUNGSTATE
f FILTRATION!
FILTRATION
•HCI
FINAL HCL
DIGESTION
TUNGSTIC ACID (SOLID)
WASH & FILTER
DIGESTION \
RESIDUE )
50 kg /
STORED ON LAND
FOR EVENTUAL
RETREATMENT
(NOTE: AT LEAST ONE
PLANT STORES RESIDUE
INDOORS IN DRUMS.)
±
AMMONIA
REACTOR
AMMONIUM PARATUNGTATE SOLUTION
NUMERICAL VALUES ARE IN
KILOGRAMS PER METRIC TON
OF METAL PRODUCED
CRYSTALLIZATION [-NHa RECYCLE-!
1
— H2
— NH3 RECYCLE —I
H2 REDUCTION
PURE TUNGSTEN
METAL PRODUCT
Figure 12 PRIMARY TUNGSTEN REDUCTION
180
-------
Wastewaters . Dilute slurries from each process containing
dissolved and suspended solids flow into a lagoon. Daily flow to the
lagoon is 1817 m (480,000 gallons) and carries approximately 395 kg of
solids (482 Kg/MT tunsten product) . Yearly accumulation of wet sludge
in the lagoon is approximately 345 MT containing 138 MT of solids.
About once a year sludge is dredged from the lagoon and dumped on land.
Samples of this sludge were not obtained but consideration of
the composition of ore concentrate and processing reagents indicate that
soluble salts of sodium and ammonium will be present in addition to
traces of arsenic, copper, lead and zinc. The fine particulate nature
of wastewater sludges and solubility of sodium salts indicates that
significant leaching could occur. This waste is therefore considered
potentially hazardous at this time,
8.2.3 Waste Quantities
Table 81 gives waste generation factors for a primary Tungsten
plant and Table 82 gives the yearly generation of wastewater sludge and
digestion residue from a typical plant.
Based on waste factors for digestion residue and wastewater as
previously given, state by state production estimates, and material
balance calculations, estimates of land disposed wastes and potentially
hazardous constituents thereof were made. These estimates are given on
state by state, regional and national levels in Table 83 for 1974, 1977
and 1983. Production and pollution control technology in 1977 and 1983
are assumed to not change significantly, thereby producing about the
same quantities of land disposed wastes as in 1974.
8.3 TREATMENT AND DISPOSAL TECHNOLOGY
8.3.1 Current Waste Treatment and Disposal Practices
Digestion residue is usually stored on land before recycling.
Wastewater from filtration processes is settled in unlined lagoons,
dredged once a year and open dumped on land. As discussed previously
hoth of these wastes are considered potentially hazardous. Levels of
technology for their treatment and disposal are discussed 1n the following
sen- 1' I
S.3.2 Present Treatment and Disposal Technology (Level I)
Digestion Residue The digestion residue remaining after
treatment of tungsten ore concentrate is stored on land for a period of
one or more years before eventual recycling. Storage on land constitutes
Level I practice. This is environmentally inadequate if trace metals
(Cu, Ni, Zn, others) percolate through permeable soils to groundwater.
181
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Wastewater The wastewater from various filtration and washing
processes as shown in Figure 12 is sent to unlined lagoons. Solids from
this water produce sludge at a rate of 482 Kg/MT of tungsten metal
produced. The wet sludge amounting to 345 NfT (138 NfT dry weight) per
year is dredged from the pond once a year and open dumped on land. This
practice is inadequate if heavy metals leach and percolate through
permeable soils to groundwater.
8.3.7) Best Technology Currently Employed (Level II)
Digestion Residue At least one plant is storing digestion
residue in drums rather than on land. Since this practice precludes any
leachnij', of potentially hazardous constituents, it constitutes Level II
prud i Cf.
Wasjtewa.ter Sludge At least one company is chemically fixing
lagoon sludges before disposing on land. Since this practice will
preclude or greatly inhibit leaching of potentially hazardous constituents,
it is considered a Level II practice and environmentally adequate.
8.3.4 Technology to Provide Adequate Health and Environmental
Protection (Level III)
Digestion Residue Storage of digestion residues in drums or
equivalent storage will preclude any leaching thereby constituting Level
III practice. This is the same as Level II technology.
Wastewater Sludge If significant leaching of hazardous consti-
tuents from wastewater sludge is demonstrated then the use of lined
lagoons will be needed. Chemical fixation of sludge removed from lagoons
before land disposal will be Level III practice if leaching of hazardous
constituents is demonstrated.
Features of Levels I, II and III treatment and disposal technology
are summarized in Tables 84 and 85.
8.4 COST ANALYSIS
In the last section various treatment and disposal technologies
currently employed, or considered for adequate health and environmental
protection were described. The costs of implementing this technology
for typjcal plants is considered in this section. The typical plant
produces 300 NfT of tungsten annually and operates 350 days per year.
8.4.1 Cost of Present Treatment and Disposal Technology (Level I)
Tungsten digestion residue is accumulated at a rate of 15
MT/year. It is stored on land for months to years before eventual retreatment,
Hauling and storage costs are insignificant.
Various filtration wastewaters are sent to a lagoon. Daily
inflow is about 1,800 m . Three hundred and forty-five MT of wet sludge
is formed containing 138 MT of solids.
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The settling pond characteristics needed to accommodates one
year input of sludge with adequate retention time are:
Volume 7,500 m
Bottom Width 31 m
Top Width 43 m
Bottom Length 62 m
Top Length 74 m
Total Depth 3 m
Depth of lixcavation 1. 25 m
Circumference 245 m_
Dike Volume 2,783 nu
Dike Surface 3,308 m
Total Width 56 m
Total Length 87 m
Required Area .5 ha
The pond is dredged once a year at which time 345 MT of wet
sludge arc removed, deposited adjacent to the pond, allowed to dry and
then hauled to an on-situ dump. About 14 hrs/yr of dragline time arc
required for dredging and 16 hours of frontloader and truck time to haul
the sludge to the dump.
The dump extends over a .5 ha area, which is sufficient to
contain 20 years of sludge waste. Forty hours/yr of bulldozer time are
assigned to the dump for grading and compacting.
Level I capital and annual costs are shown in Table 86.
8.4.2 Cost of Best Technology Currently Employed (Level II)
The digestion residue is stored in 0.2 m (55 gal) drums for a
year before being reprocesses. The sludge from filtration wastewaters
is chemically fixed before land disposal. Associated capital and annual
costs arc given in Table 87.
8.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III)
Level III technology is the same as Level II except that the
lagoon is lined. Additional costs incurred are:
Capital Cost
Lagoon Liner $16,045
Total $16,045
189
-------
Table 86
Cost of Level I Treatment and Disposal Technology
Primary Tungsten plant
Capital Cost
Sludge
Lagoon
Site Preparation
Survey $ 315
Test Drilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation § Forming 3,700
Compacting 5,150
Fine Grading 1,490
Soil Poisoning 305
Transverse Drain Fields 515
Land 880
Equipment
Dump Truck 500
Frontloader 400
Dragline 490
Bulldozer 100
Dump
Survey 375
Land 1,055
Total $16,895
Annual Cost
Land $ 195
Amortization
Construction 1,600
Equipment 1,279
Operating Personnel 1,082
Repair and Maintenance
Construction 335
Equipment 100
Energy
Fuel 130
Electricity 130
Taxes 50
Insurance 150
$ 5,001
190
-------
TABLE 87
Cost of Level II treatment and Disposal Technology
Primary Tungsten Plant
Capital Cost
Equipment
Storage Drums
Total
Sludge
Digestion
Residue
Annual Cost
Land
Amortization
Construction
Equipment
Operating Personnel
Repair and Maintenance
Construction
Equipment
Energy
Fuel
Electricity
Taxes
Insurance
Chem. Fix.
Total
$ 374
$ 374
$ 175
55
10
$ 240
191
-------
Annual Cost
Construction Amortization $ 1,865
Construction Maintenance § Repair 480
Insurance 160
Total $ 2,505
Costs of Levels I, II and III treatment and disposal technologies
are summarized in Table 88 for sludge and digestion residuals from
primary tungsten refining. Costs are given in dollars per metric ton of
dry and wet waste and dollars per metric ton of tungsten product. Costs
for each waste at each level of technology are also expressed as a
percent of metal selling price.
Total industry costs for Levels I, II and III treatment and
disposal technologies have been estimated in 1973 dollars and are presented
in Table 88. The annualized industry costs for Level I treatment and
disposal technology of sludge and digestion residue is estimated as
$90,000 for 1973 or 0.13% of estimated national sales. The estimated
1973 annualized industry cost is $420,000 or 0.57% of estimated national
sales for Level II technology. The estimated 1973 annualized industry
cost for Level III technology adequate for environmental protection is
$500,000 or 0.68% of estimated national sales.
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9.0 PRIMARY SMELTING AND REFINING OF TIN (SIC 3339)
9.1 INDUSTRY CHARACTERIZATION
The only primary tin smelter in the United States is the Gulf
Chemical § Metallurgical Company, Texas City, Texas plant. 1971, 1972 and 1973
productions as given by the U.S. Bureau of Mines (Ref. 2 ) were 4,000, 4,300 and
4,500 MT respectively. Marketable by-products of the plant are cathode melting
dross, lead-tin alloy solder, copper-silver cement and ferric chloride solution.
The average selling price of primary tin in 1973 was $2.28/lb. Thus, national
sales are estimated as $22,572,000 in 1973.
9.2 WASTE CHARACTERIZATION
This section relates production processes and air and water effluent
controls to the emanation of solid waste residuals. The disposition of these
solid waste residuals is described.
9.2.1 Process Description Figure 13 is a flow diagram of the tin and
associated by-product production at Texas City. Dispostion of process and
emissions control solids is shown.
The major production process sequence employed in the production of
tin ingot as shown in Figure 13 include the following operations:
1. Roasting of concentrate in a multiple hearth furnace
to drive of S07 and produce an oxidized concentrate.
2. Acid leaching of roasted concentrate to remove iron,
copper, silver and other metallic impurities.
3. Two stage smelting to produce 95% pure tin bullion
and slag by-product.
4. Electrolytic refining of 95% tin bullion to produce
99.9% pure tin.
Effluent gases produced during roasting are treated by dry cyclone
and electrostatic precipitator and sent to the acid leach. The degree of
stack gas treatment for S02 removal is not known but may involve wet scrubbers
with resultant scrubwater solids.
Acid leach is normally done with hydrochloric acid to remove iron,
copper, silver and other metallic impurities. Leach solution is treated by
thickening (with solids returned to process stream) and cementation to produce
a copper/silver cement product and ferric chloride solution both of which are
salable commodities.
In the two stage smelting operation,coal is used as a reductant.
Dusts are removed from flue gases and returned to the smelting circuit. Iron
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tin alloys are treated, in part, by acid leach to provide tin oxide and addi-
tional ferric chloride solution for sale. Tin metal is routed to electrolytic
cells for further purification.
Electrolytic refining of crude tin from the smelter is carried out
in sulfuric acid solution to produce pure tin cathodes. The cathodes are
melted and cast into ingots. Drosses from the melting operation are either
sold for custom recovery or returned to the process stream. Anode slimes from
the electrolytic circuit are treated separately to produce a lead/tin alloy;
no details on wastes from slime treatment is currently available. The electro-
lytic circuit is reported to be operable without the necessity for periodic
bleed of electrolyte solution.
9.2.2 Description of Waste Streams
Smelting Slag From the previous discussions of tin production it
is seen that much of the residue from tin ore concentrate leaching and electro-
lytic refining are further processed for recovery of copper, silver, lead and
other by-products. Residuals of no value and a small quantity of nonrecover-
able metallics end up in a discard slag which is open dumped. From long term
materials balance studies at the Texas City smelter (Ref. 5) it has been shown
that 26% of smelter input or 915 Kg/MT of electrolytic tin output ends up as
discard slag.
Discard slag is in granulated form and therefore individual aggre-
gates are expected to range in size from sand size grains to a few inches in
diameter. The tin content of discard slag is said to average 1% (Ref.5 ).
Lead content is given as .01% (100 ppm). Iron content is given as 10.3%.
Traces of antimony are reported. Calspan was not able to obtain slag samples
for further analyses. However, mineralology of input concentrates indicate
that copper, arsenic, and zinc should be present in trace amounts. In the
absence of solubility data discard slag is considered potentially hazardous
because of lead and other heavy metal content and fine particle size resulting
from granulation.
Table 89 gives the generation factor for discard slag from the
tin smelter along with estimated concentration of tin and lead. Arsenic,
copper, zinc and antimony are believed present also but since slag samples
were not made available their concentrations could not be ascertained. Table
90 gives the quantities of slag and hazardous constituents thereof for an
annual production of 4,000 MT tin metal.
Dujjts. Dusts collected in emissions control devices are recycled
to the process.
i.i.2.'S Waste Quantities llused on production of 4,000 MT/year in recent
years and using a factor of 915 Kg discard slag/MT, it is estimated that
3,660 MT of slag was land disposed from the Texas City smelter in 1974. This
slag will contain 0.36 MT of lead and 36.6 MT of tin, potentially hazardous
substances. Production and consequent slag generation is expected to remain
at present levels through 1985 since production has varied little from 1970 to
1973, (Ref. 2 ). Table 91 gives estimated slag disposal and toxic constituents
196
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thereof for 1974, 1977, and 1983.
9.3 TREATMENT AND DISPOSAL TECHNOLOGY
9.3.1 Current Waste Treatment and Disposal Technology
The current method of waste disposal is open dumping of discard
slag. In the absence of further data to indicate otherwise, this waste is
considered potentially hazardous. Levels of technology for treatment and
disposal of slag are described in the following sections.
9.3.2 Present treatment and Disposal Technology (Level I)
Discard Slag In the absence of information from the plant it is
assumed that the present method of slag disposal is open dumping. In the
absence of samples for solubility tests it must be assumed that lead or other
toxic metals may be released in significant concentrations. If solubilized
heavy metals percolate through permeable soils to ground water protective
measures as described under Level III will be needed.
Dusts Recycling of dusts to the process is done and is a sound
environmental practice.
9-3'3 Best Technology Currently Employed (Level II)
Level II practices are the same as Level I for slag and dust.
9.3.4 Technology To Provide Adequate Health and Environmental Protection
(Level III)
Discard Slag If leaching of solubilized toxic metals to ground
water or runoff occurs then protective measures are needed. These measures
include provision for collection and treatment of slag pile runoff and sealing
of the ground surface at slag disposal sites to prevent percolation of leachate.
Dusts Recycling of dusts to the process is an environmentally sound
and commendable practice.
Features of Levels I, II and III treatment and disposal technologies
including assessment of the hazardous nature of residuals and the methods and
adequacy of treatment and disposal methods are summarized in Table 92.
9.4 COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental protec-
tion were described. The cost of implementing this technology for a typical
tin smelter is estimated here.
199
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9.4.1 Cost of Present Treatment and Disposal Technology (Levol I)
The plant produces 4,000 MT of tin annually which results in the
generation of 3,660 MT of slag (915 kg/MT tin). The slag is disposed of on an
on-site dump.
Loading and hauling are estimated to require 200 hrs/yr of front
loader and truck tine. An additional 20 hrs/yr of bulldozer time is assigned
at the dump for spreading and compacting the slag.
The slag has an estimated density of 1675 Kg/m ; thus 2,185 m of
waste are disposed annually. A .5 ha dump site is required to accommodate
20 years of slag assuming it is piled to an average height of 10 m. Capital
and annual costs are summarized in Table 93.
9.4.2 Cost of Best Technology Currently Employed (Level II)
The technology and costs for Level II technology are the same as
Level I.
9.4.3 Cost of Technology to Provide Adequate Health 5
Environmental Protection (Level III)
The land is sealed at the slag dump, collection ditches, and pump
and piping are installed. The incurred and incremental costs are:
Capital cost
Land sealing $ 10,000
Collection ditches 1,070
Pump and piping 8,210
Total $ 19,280
Annual Cost
Construction Amortization $ 1,285
Equipment Amortization 1,310
Construction Maintenance 5 Repair 330
Equipment Maintenance § Repair 410
Energy
Electricity 30
Insurance 195
Total $ 3,560
Costs of Levels I, II and III treatment and disposal technology are
sujmnarized in Table 94 for smelter discard slag. Costs are given in dollars
per metric ton of waste and dollars per metric ton of tin product. Costs for
each level of technology are also expressed as a percent of metal selling price.
202
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TABLE 93
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
PRIMARY TIN SMELTING AND REFINING PLANT
CAPITAL COST
Equipment Slag
Equipment
Truck (10%) $ 2,500
Front loader (10%) 2,000
Bulldozer (1%) 160
Dump
Survey 375
Land 875
Total $ 5,850
ANNUAL COST
Land $ 90
Amortization
Construction 35
Equipment 740
Operating Personnel 4,565
Repair and Maintenance
Construction
Equipment 235
Energy
Fuel 565
Electricity 55
Taxes 20
Insurance 60
Total $ 6,365
203
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Total industry costs for Levels I, II and III treatment and disposal
technologies have been estimated in 1973 dollars and are presented in Table 94
The annualized industry costs for Levels I and II treatment and disposal tech-
nology is estimated as 10,000 or approximately 0.03% of estimated 1973 material
sales value. The annualized industry cost for Level III technology (i.e.,
adequate for environmental protection) is estimated as 11,000 or 0.05% of
estimated 1973 national sales value.
205
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10. PRIMARY SMELTING AND REFINING OF MAGNESIUM (SIC 3339)
Reported production of primary magnesium metal in the United States
has steadily increased from 89,246 MT in 1968 to 109,611 MT in 1972 (Ref. 1).
Nearly the entire U.S. production of magnesium in 1972 was attributed to the
Dow Chemical Company at Freeport, Texas. American Magnesium Company, Snyder,
Texas and NL Industries, Rowley, Utah also produced minor amounts of metal.
Total domestic production capacity for magnesium metal is expected to be about
213,000 MT/year by 1975 (Ref. 1). The 1973 production of primary magnesium
metal was 124,538 MT. The selling price was 38.25 cents/lb., (Ref.2 ). Thus,
national sales can be estimated as $105,017,000. Because of insufficient data
collection it was not possible to quantify the types of wastes associated with
primary magnesium production or the costs associated with waste disposal. Con-
sideration of the raw materials and processes used in magnesium metal produc-
tion indicates the absence of hazardous components in significant concentrations
or volumes. Very little or none of this industries' residuals will be land
disposed.
Production processes and qualitative assessment of the nature and
disposition of wastes are presented in the following paragraphs.
The major method employed for production of magnesium metal is elec-
trolysis of seawater (Dow Chemical, Freeport, Texas) or other high salt water
such as the Great Salt Lake (NL Industries, Rowley, Utah). This process is
shown in Figure 14, Minor production of magnesium is from dolomite limestone
by reduction with ferrosilicon at high temperatures (Northwest Alloys, Inc.,
Addy, Washington). This process is shown in Figure 15.
Neither of the above processes can be expected to produce significant
quantities of potentially hazardous wastes. Seawater which is the raw material
for the Dow process contains less than 0.1 ppm of each of a number of heavy
metals and is discharged back into the ocean after precipitation of magnesium
as magnesium hydroxide. The principal waste is a filter cake consisting mainlv
of sodium chloride and calcium sulfate which are nonhazardous. This cake is
slurried off for waste disposal in the ocean.
The only significant solid wastes from the ferrosilicon process are
slags from the retort and smelting furnaces. These slags are generated at
rates of 7055 kg and 64 kg per metric ton of magnesium produced for retort and
smelting furnaces, respectively. The slags are open dumped.
Dolomitic limestone which is the principal raw material for produc-
tion of magnesium by the ferrosilicon process contains only small amounts of
trace metals. Typical analysis of limestone for trace metals is given in
Table 95.
Limestone is used extensively for soil pH adjustment in agriculture.
Discard slag would not be expected to differ significantly in concentration
of heavy metals and, therefore, is considered nonhazardous.
206
-------
Mg(OH)z
FIGURE 14 PRODUCTION OF MAGNESIUM
FROM SEAWATER
207
-------
Raw materials
Dolomite ore
(mjgnenum cwbonilt +•
Mfcium c«rbon*ti)
Calcined dolomite
(magnesium o>id« + calcium oxide)
Calcined dolomite powder
1
Pulverized
(errosilicon
Ctteliwd dolomltt -f
l*frotilicon powdw
Briquetting machine
Culcitwd dolomite -
Itrroiilicon pelM>
Magnesium
cryjtali
Calcium silicate
+ iron
FIGURE 15 PRODUCTION OF MAGNESIUM
BY FERROSILICON PROCESS
208
-------
TABLE 95
TRACK P.MiMUNT COMPOSITION OF LIMESTONE
Element Concentration (ppm)
As i
Be 1
Cd 0.04
Cr 11
Cu 4
Pb 9
Hg .04
Se 0.08
Zn 20
209
-------
11.0 PRIMARY SMELTING AND REFINING OF CADMIUM (SIC 3339)
There is no discrete primary cadmium industry in the United States.
Primary cadmium metal and compounds are byproduct materials produced by
companies principally engaged in the recovery of zinc either as a major value
or as a byproduct itself. This is because there is no separate ore of cadmium
and because it occurs in significant quantities in nature exclusively in
association with zinc sulfides. Major cadmium producers in the United States
are shown in Table 96. Total cadmium metal production in the country is small,
Production for the years 1971 to 1975 was as follows: (Ref. 10)
1971
1972
1973
1974
1975
3580 MT
3765 MT
3400 MT
?020 MT
1990 MT
TABLE 96 Producers of Cadmium Metal
American Smelting § Refining Company Denver, Colorado
American Smelting § Refining Company Corpus Christi, Texas
Amax Zinc Company East St. Louis, Illinois
The Bunker Hill Company Kellogg, Idaho
National Zinc Company Bartlesville, Oklahoma
The New Jersey Zinc Company Palmerton, Pennsylvania
St. Joe Minerals Corporation Monaca, Pennsylvania
United Refining & Smelting Company Franklin Park, Illinois
These figures include secondary cadmium as well as primary since individual
breakdowns are not published. Table 97 gives the state and regional distribu-
tion of cadmium production for the year 1973. Plant capacities are not
available.
Examination of the various alternative methods for recovery of
cadmium metal from zinc smelters reveals that relatively small quantities of
nonrecyclable waste are produced. These wastes are considered potentially
hazardous, however. Since cadmium production is always a minor operation at
primary zinc smelters and occassionally at primary lead or copper smelters the
waste streams become minor inputs to the larger waste streams.i These waste
streams are covered previously in this report. Cadmium recovery wastes will
always report either as sludges or dusts. As such, lined lagoons or sealed
soil areas would be required for their disposal as stipulated in the sections
on primary zinc, lead and copper. For these reasons and unavailability of
210
-------
TABLE 97
GEOGRAPHICAL DISTRIBUTION OF PRIMARY CADMIUM PRODUCTION
State
Colorado
Idaho
Illinois
Oklahoma
Pennsylvania
Texas
KPA Region
3
5
6
8
10
National
Ho. of Plants
1
1
1
1
2
1
Plant Distribution by Production (MT)
NA
522 (1973)
123 (1973)
NA
662 (1973)
NA
662 (1973)
123 (1973)
NA
NA
552 (1973)
3440 (1973)
211
-------
data on quantification of cadmium wastes from the many alternative processes
for recovery details on quantities of cadmium plant wastes and costs of
treatment and disposal are not presented in this section.
Contained in the following is a description of the alternative
methods for cadmium metal production and identification of possible sources
of potentially hazardous residuals which may be disposed on land.
Process Description
The starting material for cadmium-recovery processes are by-products
of other metal-production operations, all involving zinc. These are:
(1) Fumes and dusts from the roasting and sintering of
zinc concentrates;
(2) Zinc metal containing relatively high concentrations
of lead and cadmium;
(3) Dusts from the smelting of lead-zinc ores or copper
lead-zinc ores;
(4) Purification sludge from zinc electrowinning plants.
Cadmium from Zinc Concentrates. In the recovery of zinc from zinc concen-
trates by the retort process, zinc sulfide concentrates are roasted, sintered
(see zinc section), and introduced into retorts with reducing agents such as
coal or coke. When the mixture is heated, the zinc oxide is reduced to zinc
metal which is volatilized, condensed, and cast into ingots. Unless special
provisions are taken, much of the cadmium and lead almost always present in
zinc concentrates would accompany the zinc. Figure 15 outlines a process used
to separate these elements from the zinc and at the same time recover cadmium.
Roasting is done in multiple-hearth furnaces under oxidizing condi-
tions to eliminate most of the sulfur (1). At the temperature of the roast-
ing operation, cadmium tends to volatilize preferentially and concentrates
to some extent in the flue dust. This is collected in suitable equipment
such as cyclones, baghouses, and electrostatic precipitators and is returned
to join the roasted product. The roasted concentrates and flue dust are then
charged onto the surface of a sintering machine (2) which agglomerates the
calcine, removes residual sulfur, and removes most of the lead and cadmium
from the zinc. Lead-cadmium removal is accomplished by chloridizing the
charge with the zinc chloride solution returned from filtration step or by
the addition of sodium chloride solution. Under proper conditions of temper-
ature and chloride addition, cadmium and lead chloride volatilize from the
i h.-irgc «uul ;irc drawn down through the bed and into the fume-collecting facil-
ities for subsequent treatment. The sintered zinc oxide product is sent to
the retorting furnaces for the recovery of zinc metal.
212
-------
ZiNC SUCfiOfc CONCENTRATES
ROASTING*1'
RETURN SINTER
1
t '
». FLUtOUST —
t
t *
FINISHED ZINC CADMIUM AND
SINTER LEAD FUME
i 1
TO RETORTING DUST COLLECTION
t
1 r"
[_ LEACHING*31
t
LEAD SULFATE
f CADMIUM
j LEAD RECOVERY j SOLl
,
,
noN*4>
SULFATE
TION
f f
PRECIPITATION16'
1
F
FILTRATION*6'
SODIUM CHLORATE
SULFURICACIO
CADMIUM SPONGE
f
DRYING (7>
1
r
MIXING*8'
1
r
REDUCTION*91
LIME
FUEL
CHLORIDE
SOLUTION
RESIDUE
METALLIC CADMIUM
CASTING
PRIMARY NONFERROUS METALS INDUSTRY •
V/l COPPER, ZINC AND LEAD INDUSTRIES,
EPA-R2-73-247*. SEPTEMBER 1973
Figure 16 CADMIUM RECOVERY IN THE ZINC RETORT PROCESS
213
-------
After collection, the cadmium and lead fumes and dusts are leached
with dilute sulfuric acid and sodium chlorate (3) to ensure the complete
dissolution of cadmium sulfide which would otherwise remain undissolved. In
the leaching operation, cadmium and lead are converted to sulfates and/or
chlorides. Cadmium sulfate and chloride remain in solution, but lead is
almost completely converted to insoluble lead sulfate. This compound, toget-
her with other insoluble materials (quartz, silicates, etc.). is filtered out
ami sent to lead recovery, (4). The .solution of cadmium sulfate (and/or
chloride) is then treated with powdered zinc to precipitate cadmium (5). The
slurry is filtered (6). The solution from filtration, containing practically
all of the zinc addad and about 10 percent of the cadmium as chlorides and
sulfates, is returned to the sintering operation.
The cadmium filter cake, contains about 30 percent moisture and
only small amounts of lead and zinc. This sponge is then steam dried, (7) and
mixed with coal or coke and lime, (8) and transferred to a retort furnace
where the cadmium is reduced to metal, vaporized, recondensed, and cast into
required shapes. (9) Residue from the reduction furnace or retort containing
some cadmium is returned to the sintering'operation.
Cadmium from Zinc Mstal. Some zinc made by the retort process contains both
cadmium and lead in appreciable quantities. An example is the so-called
prime Western zinc which may contain about 1.5 percent lead and several tenths
of a percent cadmium. Such a material can be highly1 refined by the process
shown in Figure 17 . This process involves two fractionation stages> each at
different temperatures. In the first, (1) operated at a temperature in excess
of 1000 to 1150 C, cadmium and zinc are boiled off, leaving a high lead-zinc
alloy. The cadmium-zinc vapors are condensed and refractionated at about
800 C. (2) In this operation, cadmium is volatilized and a high-purity zinc
product is withdrawn at the bottom of the fractionation column.
The cadmium product from this process generally requires further
lr
-------
Zinc V.ctal
(Impure)
Lead-
Zinc -*-
Alloy
Fracticmation (1)
Zinc-Cadmium
Vapors
Condensation
Zinc-Cadmium
Liquid
Fractionation (2)
Refined Zinc
99.99% Pure
Cadmium Vapors
I
Condensation
Cadmium-Zinc Metal
or Cadmium Dust
To Purification
FIGURE 17 CADMIUM RECOVERY FROM ZINC METAL
SOURCE: Water Pollution Control in The
Primary Nonferrous Metals Industry, VOL.1
Copper, Zinc and Lead Industries,
EPA-R2-73-247a, Sept. 1973
215
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Another route is shown in the center column of Figure 18. By this
route, the baghouse dust containing about 3 to 4 percent cadmium is refumed
in a reverberatory furnace (5) in several batch stages in order to obtain a
fumed product of optimum grade for subsequent processing. This fume, (6)
which may contain 40 to 45 percent cadmium is contacted with water and oxidi-
zing agents KMnO. and NaC10, (7) in a ball mill and leached by agitation with
additional sodium chlorate,'ferrous suJfate, and sulfuric acid. (8) The
purpose of the oxidising ag „;•*•'• ar>r! *hf ferrous sulfate is to ensure that
cadmium is all dissolved and that arsei, n. is oxidized and precipitated as
ferric arsenate. Die slurry is iilteroo, v?) the residue, consisting of lead
sulfate and ferric arscnute, 1.-. aorit. Is&ck : o lead smelters.
Precipitation with zinc dust under controlled conditions of acidity
(10) produces a fairly high-grade "sponge", fll) This is filtered, washed
by repulping with water, refiltereJ., ar.u uiied. (12) The filter cake is then
granulated and formed into s;r.al: lu'u;jettes. (13) The briquettes are then
retorted batchwise in graphite crucibles to volatilize cadmium. (14) The
cadmium vapors are condensed (15) and the liquid metal cast into bars. These
are remelted under a layer of sodium hydroxide flux is removed and ammonium
chloride is added to remove thallium from the cadmium. The cadmium is then
recovered with sodium hydroxide flux and the purified metal, which may contain
99.95 percent of cadmium, is a<,t ir^o '•>TT:-.
The third route, shown at the right side of Figure 18 is also applic-
able to lead and zinc smelter fumes. In this process the fume or dust is
leached by repulping in water (17) and the slurry is treated with chlorine
gas. (18) Subsequent steps of filtration, (19) precipitation, (20) and
isolation of cadmium sponge are similar to those previously described. The
resultant sponge may be subsequently treated by electrolytic or pyrometal-
lurgical methods to produce pure cadmium.
Cadmium Recovery from Purification Sludge in Zinc Elcctrowinning. The electro-
winning process for zinc is described in the section of this report dealing
with that metal. The successful electrodeposition of zinc requires extremely
pure solutions. Cadmium, among other metals, such as copper, nickel, cobalt,
and iron, must be virtually completely removed. In zinc metallurgy, both
copper and cadmium are eliminated from solution by treating the zinc electro-
lyte with powdered zinc. Generally this is done stagewise. In the first
stage, only enough zinc metal is added to precipitate copper, which, reacts
preferentially with the zinc powder. After removal of the copper precipitate
by filtration, the zinc electrolyte is then treated with additional zinc dust
to precipitate the cadmium. This precipitate, high in zinc, contains virtually
all the cadmium originally in the electrolyte and constitutes the purification
sludge. Its processing is shown in Figure 19 .
Purification sludge (1) is first allowed to oxidize in air (2) with
or without heating, to render the cadmium readily soluble. The oxidized
material is then leached with spent electrolyte from the electrolysis step.
(3) After filtration (4) to remove insoluble copper introduced from the zinc
electrolyte, the filtrate is stage-treated with zinc dust (5-8) to minimize
217
-------
PURIFICATION SLUDGE (1)
I
OXIDATION (2)
I
LEACHING
(3) }•
FILTRATION (4)
COPPER CAKE TO
COPPER SMELTER
FILTRATE
FIRST PRECIPITATION (S) }••»•
SPENT
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FILTRATE
CADMIUM SPONGE
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T
GLUE, ETC.
CADMIUM CATHODES
i
MELTING (13)
SOURCE: WATER POLLUTION CONTROL
IN THE PRIMARY NONFERROUS METALS
INDUSTRY, VOL I, COPPER, ZINC AND
LEAD INDUSTRIES, EPA-R2-73-247a,
SEPT. 1973
CAST SHAPES
Figure 19 CADMIUM RECOVERY FROM PURIFICATION SLUDGE OF ZINC
ELECTROWINNING OPERATIONS
218
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zinc concentration in the cadmium sponge. This sponge containing about 80
percent cadmium with 5 percent zinc, is again oxidized by steam drying (9)
to enhance the solubility of cadmium. It is leached in spent electrolyte (10)
from the electrolysis circuit and filtered. (11) The filtrate, containing
about 200 grams per liter of cadmium as sulfate, is then electrolyzed (12) to
produce cadmium cathodes.
Electrolysis is ci'-ri°d ov1: i banks of cells similar to zinc cells.
After passage through the ceLls, the spent electrolyte is returned to various
leaching stages in the circuit. Stripped e.Jthode metal is washed, dried, and
melted (13) under a flux and cast into various shapes.
It is seen from the flow diagrams and previous discussion that most
residuals from cadmium production are recycled for metallic recovery. A
sample of a cadmium plant filter cake residue was obtained from one zinc
smelting facility (cadmium is recovered from flue dust). This waste amounted
to 514 kg/day. Analyses of this land dumped waste is given in Table 98.
TABLE 98
ANALYSIS OF CADMIUM PLANT RESIDUE
Cd 280 ppm
Cr 24 ppm
Cu 1,150 ppm
Pb 215,000 ppm
Zn 39,000 ppm
In solubility tests described in AppendixB this waste was found
to leach significant concentrations of lead (9.0 ppm) and zinc (4,000 ppm)
and is considered a potentially hazardous waste as are all cadmium plant
residues.
Since cadmium is recovered from the residual dusts or sludges
primarily from zinc smelters and to a lesser extent at lead or copper
smelters, and environmentally acceptable methods of residual disposal have
been identified for these smelting categories the safe disposition of cadmium
processing residuals is covered under these categories.
Any standards regarding the environmentally sound methods of
residual disposal of cadmium plant wastes on land should be handled as part
of overall standards for primary zinc smelters or specialty smelting plants.
219
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12. PRIMARY SMELTING AND REFINING OF ARSENIC (SIC 3339)
The production of arsenic in the United States is almost exclusively
MS a by-product from the smelting of copper and lead ores. In the roasting
and smelting of such ores, arsenic oxidized to As 0 , is volatilized, and is
recovered in fumes and flue dust. These contain varying amounts of the tri-
oxide (up to 30 percent) associated with the oxides of other metals such as
copper, lead, antimony, zinc, or tin, and various other particulate materials.
The most marketable form of arsenic is the pure trioxide, with little
demand for pure elemental metallic arsenic. Only about 3% of arsenic is pro-
diu-ed as the metal. (Ref.2 ) In the United States, the only primary metal
producer now making either arsenic trioxide or arsenic metal is the Tacoma
smelter of the American Smelting and Refining Company. It recovers arsenic
from the flue dust of its own and other copper and lead smelters. (Ref. 7)
In earlier years Anaconda also produced arsenic in Montana. No production data
have been published in recent years. The U.S. Bureau of Mines estimated U.S.
production of arsenic compounds and metal as 9253 MT in 1973. Only 277 MT
(i.e3%) is arsenic metal. The price of arsenic metal in 1973 was 1.00/lb,
(Ref. 2). Thus, the national sales of arsenic for 1973 can be estimated as
$554,000.
In order to characterize and quantify the total and hazardous waste
associated with arsenic production it will be necessary to visit and sample
such wastes (if there are any) from the one U.S. producer (ASARCO, Tacoma,
Washington). At that time treatment and disposal technology and environmental
adequacy thereof may be identified and associated costs estimated. A brief
description of flue dust processing for arsenic recovery and associated wastes
follows.
Processing consists of roasting the flue dusts to volatilize the
contained As 0 , which concentrates in the fume from the roasting operation.
The collected Fume is again heated in a reverberatory furruu/c, and the fume
from this operation is passed through a series of chambers ^- collect fractions
of various purity from the volatilized furnace output.
Potentially hazardous arsenic compounds are always present in the
collected dust emissions from copper, lead, zinc and other base metal roasting
operations. These dusts are frequently recycled for recovery of arsenic or
other metallies. Where they are not recycled, but disposed of or stored on
land for significant time periods, attention must be given to possible leaching
of arsenic. Sealing of disposal or storage areas may be necessary.
220
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13.0 PRIMARY SMELTING AND Ri-l-'INING OF SELENIUM AND TELLURIUM (SIC 3339)
13.1 INDUSTRY CHARACTERIZATION
The only, raw material for the recovery and refining of selenium
and tellurium in the United States is electrolytic slimes, a by-product
of copper refining. The precious metals gold, silver and platinum are
also recovered from slimes.
Production of these metals is small. Production of selenium
has ranged from 287 MT to 566 MT from 1968 to 1972, while production of
tellurium has ranged from 55 MT to 117 MT over the same time period
(Ref. 1).
U.S. producers of selenium find tellurium metal and compounds
are listed in Table 99.
TABLE 99
United States Producers of Selenium and
Tellurium Metals and Compounds
American Metal Climax, Carteret, N.J.
American Smelting and Refining Co.,
Baltimore, Md.
International Smelting and Refining Co.
Perth Amboy, N.J.
Kennecott Copper Corp., Garfield, Utah
Anaconda Co., Perth Amboy, N.J.
Se
X
X
Te
X
X
X
X
Major electrolytic copper refiners ship anode muds and slimes to the above
locations for recovery of precious metals, selenium and tellurium.
The wastos associated with selenium and tellurium rscovery will
be minor constituents of the total wastes from copper electrolytic refiners.
As shown in Figure 2 (pg. 11,1 previously for every 1000 kg (1 MT) of electrolytic
copper produced 1.1 kg of slimes are produced. These slimes are in turn
processed for selenium, tellurium and precious metals recovery. Thus any
quantity of residual waste remaining would be small. These residuals in the
form of water and sludges would be disposed of along with the other waste
sludges from the electrolytic copper plant. In Section 1.0 of this report
dealing with electrolytic copper refineries the use of lined lagoons has been
221
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stipulated for adequate environmental protection. This practice would
insure adequate disposal of the residues from selenium and tellurium and
precious metals recovery.
For the above reasons and lack of specific data from the
plants processing electrolytic slimes for selenium and tellurium re-
covery, quantification of wastes and treatment and disposal costs are
not contained in this report.
Production processes are briefly described here and possible
types and sources of hazardous wastes are identified.
13.2 WASTE CHARACTERIZATION
13.2.1 Process Description
Selenium Smelting and Refining. More than 90 percent of the
domestic output of selenium is as a byproduct of copper refining. The
selenium is recovered from the anode mud or slime that falls from the
anode during electrorefining of copper. It also is recovered from some
lead refining operations.
All commercial processes may be considered as modifications or
combinations of three fundamental methods: Smelting with soda ash,
roasting with soda ash, and roasting with sulfuric acid. In the first
two methods, the selenium in the anode slimes is converted to sodium
selenate which is soluble in water. Tellurium is precipitated from the
solution by neutralization with acid and recovered by filtration, and
then elemental selenium is precipitated with sulfur dioxide. In the
third process, the anode slimes are roasted with sulfuric acid. Selenium
dioxide volatilizes and is recovered in a wet Cottrell precipitator as
selenious acid from which red amorphous selenium is precipitated by the
addition of sulfur dioxide. Subsequent treatment of the various products
resulting from the initial operations of all three processc recovers
additional amounts of selenium. Overall recovery of seleniwn in the
slimes is 60 to 70 percent (Ref. 7).
Tellurium Smelting and Refining. Tellurium-refining methods
are closely guarded trade secrets, and modifications which different
companies make in the standard flowsheet are not disclosed. Processing
begins with tellurium-bearing anode residues and slags from the refining
of copper and lead.
The anode slime is leached in acid solution to remove copper
and then fused with sodium bisulfate to convert residual copper to
copper sulfate and to volatilize part of the selenium. The fused cake
is granulated and leached with water to remove the copper sulfate. A
subsequent leach with caustic soda solution extracts the tellurium.
Residue from the double leach is refined in a Dore furnace for recovery
of precious metals. The tellurium is recovered from the caustic solution
by acidifying to precipitate a tellurium oxide sludge that is subsequently
refined to metal.
222
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About 0.17 pound of tellurium is produced in the United States
for each ton of copper produced. Inasmuch as copper anodes contain
about 0.4 pound of tellurium per ton, only about 50 percent of the
contained tellurium is recovered in a marketable form (Ref. 7).
With respect to the possible land disposal of potentially
hazardous waste from seleniuir, and tellurium (and precious metals which
are also recovered from anode siimes arm muds), it is important to note
that estimated recovery of selenium am.; tellurium from copper anode
slimes is 60 to 70 and 50 percent, i-espectively. It may be expected
that since recovery is only 50 to 70?o -Ha;'if leant concentration of these
potentially hazardous metals, along with arsenic, copper and lead and
other heavy metals, will end up in wastewater and disposal slags from
specialty smelters which process copper anode slimes and mud. A study
of the specialty smelting plants listed in Table 99 would be needed to
determine the exact quantities and disposition of potentially hazardous
residuals from selenium and tellurium production.
223
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14.0 PRIMARY SMELTING AND REFINING OF GOLD AND SILVER (SIC 3339)
M.I 1NDUSTRY CHARACTERIZATION
Silver
Less than one percent of the new silver produced in the U.S.
comes from simple silver of gold-silver ores amenable to direct processing
by amalgamation and/or cyanidation. That is, practically all primary
production is from ores bearing lead, zinc, and/or copper; silver-
bearing concentrates floated from these ores are smelted at the facilities
designed to produce primary lead, zinc, and copper as discussed elsewhere
in this report.
A total of 36,494,000 troy 02. of silver were produced in 1973
(Ref. 2) at an average selling price of $2.558 per troy ounce. National
sales of primary silver can then be estimated as $93,352,000.
The one percent share alluded to above amounts to roughly
200,000 troy ounces of silver and is produced at installations which are
primarily gold smelters because the lode gold-silver ores from which the
silver is produced contain much more gold than silver. The largest of
these smelters is discussed below.
Gold
The 1973 U.S. production of gold metal was 1,322,000 troy
ounces (Ref. 2). The average selling price of gold was $97.81 per troy
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Gold
As shown in Figure 20 the only solid material not totally recycled
or sold Is that portion of the slag from the scavenger furnace which goes
bat* to the ore treatment mill. This slag "bleed" amounting to less than
O.I MT/day ultimately leaves the plant with the ore tailings and is insigni-
ficant compared to the mine tailings which amount to 4700 MT/day.
14.3 TREATMENT AND DISPOSAL TECHNOLOGY
Silver
As discussed previously primary silver is almost entirely re-
covered as by products of lead, zinc, and copper smelting or from gold ores.
The minor amount of residues attributable to silver recovery at these
primary smelters makes it unnecessary to describe Level I, II and III treat-
ment and disposal technology.
(iold
As discussed previously, the only disposable waste from recovery
of gold from concentrates is slag from the scavenger furnace. Less than
0.1 MT/day of this slag ultimately leaves the plant with ore tailings amount-
ing to 4700 MT/day. Description of treatment and disposal technology for
environmental protection is, therefore, not warranted.
14.4 COST ANALYSES
Since the quantities of land disposed wastes from primary silver
and gold smelting and refining are negligible, costs of treatment and
disposal technology are negligible.
226
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15.0 PRIMARY SMELTING AND REFINING OF PLATINUM
15.1 INDUSTRY CHARACTERIZATION
As one would expect, the production of primary platinum in the
United States is small. Production is reported in conjunction with
palladium, iridium, osmium, rhodium and rutlemium which comprise the
platinum-group metals. Prediction of *he nlatinum-group metals from
primary refiners ranged from 12,000 troy ounces to 21,000 troy ounces
(0.373 MT to 0.653 MT) from 19t.8 to 19/2 tRef 1). The major part of
U.S. output is recovered as a specialty refining byproduct of copper
refining in Maryland, New Jersey, Te,'.,'E '".ah and Washington. Raw
material for recovery of platinum-group metals are anode muds or slimes
from electrolytic copper plants. Selenium, tellurium, gold and silver
are also recovered from copper anode muds and slimes.
The wastes associated with platinum recovery will be minor
constituents of the total wastes from copper electrolytic refiners. As
shown in Figure 2 (page 11) previously, for every 1000 kg (1 MT) of electro-
lytic copper produced 1.1 kg of slimes are produced. These slimes are in
turn processed for selenium, tellurium and precious metals recovery includ-
ing platinum. Thus any quantity of residual waste remaining would be
small. These residuals in the form of water and sludges would be disposed
of along with the other waste sludges from the electrolytic copper plant.
In Section 1.0 of this report dealing with electrolytic copper refineries
the use of lined lagoons has been stipulated for adequate environmental
protection. This practice would insure adequate disposal of the residues
from platinum selenium and tellurium and other precious metals recovery
from slimes.
For the above reasons and lack of specific data from the plants
processing electrolytic slimes for platinum recovery, quantification of
wastes and treatment and disposal costs are not contained in this report.
Processes and possible sources of land disposed wastes with potentially
hazardous constituents thereof are briefly described in the following
paragraphs.
Copper anode muds containing platinum and other precious metals
are washed and filtered to remove soluble copper which is recycled to
copper electrolysis cells or disposed of to tailings impoundment if metal
content is too low for recovery. Washed muds are roasted with sulfates
and sulfuric acid, leached with water, and treated with copper powder to
promote precious metal precipitation. Flue gases from the roaster are
scrubbed (wet) and subjected to electrostatic precipitation to yield a
recovered, marketable so Ionium and selenium oxide product and wastewater
which is routed to disposal ponds.
227
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The crude cemented precious metal concentrate, rich in copper,
is isolated by filtration and the filtrate is routed to the copper electro-
winning circuit. The concentrate is melted in special furnaces known as a
Core furnace with a variety of fluxing agents. Crude silver anodes and a
slag which is routed to the copper smelter are derived from the Dore furnaces.
Fumes from the Dore furnace are routed through both wet and dry control
systems and all solids are returned to the roaster while clarified solution
is routed to waste ponds.
Dore metal anodes are electrorefined to recover silver^ and anode
muds are routed to a leaching circuit, induction furnace, and anode casting
module for the preparation of gold anodes for electrorefining in Wohlwill
cells. Wastes generated in this circuit are either recycled to the silver
recovery circuit (if solids) or to the platinum/palladium recovery circuit
(in the case of spent or contaminated Wohlwill cell electrolyte).
Spent Wo'hlwill electrolyte is treated with ammonium chloride and
chlorine to precipitate ammonium chloroplatinate which is isolated by filtra-
tion, dried, and reduced to platinum metal. Filtrate from the ammonium chloro-
platinate isolation is transferred to the palladium recovery circuit and gases
from the reduction furnace are vented to atmosphere through caustic scrubbers
•vith scrubbing liquors going to wastewater impoundments.
As seen in the previous discussions, possible sources of land
disposed wastes include: (1) filtrate from initial washing of copper anode
muds, (2) roaster scrubwater which has been subjected to flotation for recovery
of selenium, (3) clarifier water from Dore' furnaces, and (4) scrubber liquor
from reduction furnaces. It is presumed that these wastes are put into small
lagoons. The volumes of liquid discharge to lagoons and resultant sludge
quantities are not known but are relatively minor. Potentially hazardous
constituents include As, Te, Se, Pb and Cu.
228
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16.0 PRIMARY SMELTING AND REFINING OF BISMUTH
The quantity of primary bismuth metal produced in the United
States is very small. Production figures are not divulged. Refined
primary bismuth is produced or>ly by American Smelting and Refining
Company in Omaha, Nebraska (Ref. 1). Domestic production of approximately
381,000 Ibs bismuth in 1973 <«os estimated by subtracting imports from
exports and consumption (Ref, 3). ».io a/.iage 1973 selling price of
bismuth was $4.92/lb. Thus national 5al ;s of domestic production for
1973 was estimated as $1,875,000.
As will be described the quantities of wastes associated with
bismuth production are also expected to bo small. In order to obtain
accurate estimates of the quantities nf t>tal wastes and hazardous
constituents thereof from bismuth sjiwl^l.j^ it would be necessary to
visit and sample bismuth procosh wu.io streams from the Omaha plant.
This was not done on this program. !''<>*' those reasons bismuth refining
was not studied in detail during this study.
Bismuth is always recovered as a byproduct of copper and lead
smelting. In copper smelting, bis ,itn persists with copper until the
converting operation, in which much of it volatilizes with lead, antimony,
etc., as fumes. These are collected and sent to lead smelters. The
portion of bismuth which escapes volatilization during converting is
finally removed in the electrolytic copper refinery in the slimes.
Treatment of these slimes for the recovery of gold and silver leaves
bismuth associated with lead.
These slime residues are processed at lead or specialty smelters
for recovery of bismuth and lead. Bismuth associated with lead ore
accompanies lead through all steps of processing and is removed by
special refining processes. When bismuth has been removed from lead by
pyrometallurgical procedure, it is collected in a dross containing
calcium, bismuth, magnesium, and lead. This dross is melted and chlorinated
to remove calcium and magnesium as a molten chloride slag and a bismuth-
lead alloy. The alloy is fluxed with molten sodium hydroxide to remove
traces of arsenic and tellurium that may be present. Residual silver
and gold are removed by treatment with zinc. The resultant lead-bismuth-
zinc alloy is then subjected to chlorination treatment with chlorine gas
under controlled conditions. Zinc and lead combine with chlorine preferentially
over bismuth and are removed as molten chloride slags. The molten
bismuth is finally treated with air and caustic soda and cast. This
method of refining yields 99.999 per cent pure bismuth.
During the electrolytic refining of lead, bismuth enters the
slimes with any arsenic, silver, gold, or antimony that may have been
present. These slimes are filtered, dried, and melted to produce a slag
and metal. The metal phase is blown with air to remove arsenic and
antimony. It is then heated under oxidizing conditions, to produce a
lead-bismuth slag and an alloy of gold and silver. The lead-bismuth
slug is reduced by carbon to form u lead-bismuth alloy which is refined
by chlorination as described above.
229
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The only significant waste which may be land disposed is the
calcium and magnesium chloride slag which may contain trace quantities of
heavy metals. The quantity and exact disposition of this slag has not been
ascertained because of no plant visits.
230
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17.0 PRIMARY SMELTING AND REFINING OF COBALT
17.1 INDUSTRY CHARACTERIZATION
There is very little coha.'t metal produced in the United States.
U.S. production is reported by The Bureau of Mines (Ref. 1) as 146 MT in
1970 and 140 MT in 1971. There was not Jomestic smelter production of
cobalt for 1972 or 1973 (Ref..), "i.ic^e j.s only one U.S. smelter'with the
capabilities to produce cobalt metal. This is the American Metal Climax,
Port Nickel, La. refinery. For litose reasons the industry has not been
investigated in detail ilurin,; the OIKJJ. t <.f this study.
A Clow d i .i^r.iin .nul iirifl descriptions of pot eat iul l\ h>i^.ai\luu.s
ivslduals which may he land disposed i^ presented here. A flow diagram
for production of cobalt metal is shown 'in Figure 21. Iron pyrite concentrate,
containing cobalt value and produced by flotation of ore, is roasted to
convert sulfide ore minerals to oxides, sulfatcs, and similar compounds.
While details of dust control in roaster off gases are not available, the use
of these off gases in a sulfuric acid plant suggests that entrained dusts
are removed and either recycled or disposed of. These dusts contain Cu,
Co, Ni, and Zn and are therefore a pcicntia.1 hazard if land disposed.
Following roasting, the calcine is subjected to a chloridizing
roast to produce soluble chloride salts. Emissions contain Cu, Co, Ni and
Zn dust which pose a potential hazard if disposed of on land after collection.
The chloridized calcine is leached with water, filtered and a filter
cake washed to yield a chloride solution and a filter cake of iron oxide
which serves as feed for the steel industry. The chloride solution is
treated by scrap iron cementation (to remove copper), controlled neutraliza-
tion (to precipitate iron values), and finally filtration to isolate a
cobalt chloride solution. The cement copper is sent to copper refinery
operations for recovery and the iron hydroxide is cither sintered to provide
steel industry feed or' disposed of to landfill or pond impoundments. Tin1
i ron hydroxide' I'.'ikc i out ;i i tr. r|i lor I ill"* of tin, l!t>, Ni. tnul .'n tiiul must be
i (in-, i iK-rnl pot out i .1 1 ly ha :• tii ilou'> if luiul dli -.
The cobalt chloride solution is treated with sodium carbonate and
chlorine gas to produce an insoluble cobalt hydrous oxide which is subse-
quently recovered by filtration. The filtrate, containing nickel, zinc,
chloride, and other species is apparently discharged as waste to tailings
disposal lagoons and thus must be considered potentially hazardous. The
hydrous cobalt oxide is next dried and then reduced, using a charcoal rcduct-
ant, to produce cobalt metal.
231
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CONCENTRATE
ROASTER
CALCINE
NaCI
FtS,
CHLORODIZING
ROASTER
CHLORODIZEO
CALCINE
GASES I H2SO4
| PLANT I
EMISSIONS _ AIR POLLUTION
p*. CONTROL
H,0
LEACHING
CHLORIDE
SALTS
FILTER
CHLORIDE
SOLUTION
FILTER CAKE
(F»203)
SCRAP IRON
CEMENTATION
COBALT
SOLUTION
CEMENT
COPPER
TO STEEL
PLANTS
TO COPPER
REFINERY
CtO
FILTER
COBALT
SOLUTION
FILTER CAKE
3^ TO LAND FILL
OR STEEL PLANT
AGITATOR
COBALT
PULP
FILTER
COBALT
CAKE
WASTE
WATER
-•»• TO LAGOON
CHARCOAL
KILN
COBALT METAL
Figure 21 PRIMARY COBALT SMELTING AND REFINING
232
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18.0 PRIMARY SMELTING AND REFINING OF ZIRCONIUM AND HAFNIUM
Only small amounts of zirconium and hafnium sponge metal are
presently produced in the United States. The only U.S. producers of these
metals in 1974 were the Wah Chang, Albany, Oregon, Division of Teledyne
Corporation and the Amax Specialty Metais Division of Amax, Inc., Akron,
New York (Ref. 9). The New York plant closed in 1975. Statistical data on
production of zirconium spcrge ^rd T>«ia: and hafnium sponge is withheld to
avoid disclosure. Production or hafnuii crystal bar which is made from
hafnium sponge was reported as only j,6 NT :i 1972 (Ref. 1). U.S. consumption
of zirconium metal in 1973 was 2/22 Ml (Ref, 2).
Zirconium and hafnium are used in nuclear reactors and production
of these metals is expected to increase if there is expansion of nuclear
;>ower in the United States.
In order to obtain estimates of the types and quantities of solid
and hazardous wastes from the smelting and refining of zirconium and hafnium
metal it will be necessary to visit and sample the Oregon plant. This was
not done under the current program. For this reason, and because of minor
production, detailed presentations or solid and hazardous waste generation,
treatment and disposal technology :>nn associated costs are not given in this
report. Production processes are discussed and possible sources of poten-
tially hazardous wastes identified.
Production of Zirconium and Hafnium. The elements zirconium and
hafnium always occur together in minerals which are associated in sand
deposits with minerals containing titanium and rare-earth elements. The
earth's crust has been estimated to contain approximately 50 ppm of
zirconium minerals. Zircon (ZrSiO.) and baddeleyite (ZrO,), are marketed
commercially, but zircon is by far the more important source (Ref. 7 ).
Two types of zirconium are produced - commercial grade, and
reactor grade to be used in nuclear reactors. Commercial-grade zirconium
is produced by blending zircon with coke which is then reacted or fused
in an electric arc furnace. The carbon and the nitrogen from air react
with zirconium to yield zirconium carbonitride containing about 85 percent
zirconiun. The carbonitride is chlorinated to crude zirconium tetrachloride
which is reduced to zirconium metal with magnesium metal under an inert
atmoshpere. The reduction stage is called the Kroll process. The resulting
mixture of zirconium metal and magnesium chloride is vacuum-distilled to
remove the magnesium chloride. The residual zirconium sponge contains
about 2 percent hafnium but is stilJ suitable for most nonnuclear applica-
tions. Direct chlorination of zircon concentrates to produce the mixed
chlorides also is reportedly utilized commercially.
To produce reactor-grade or hafnium-free zirconium, crude
zirconium tetrachloride is dissolved in acid containing ammonium thiocyanate
complexing agent and put through solvent extraction, which strips hafnium
from the solution. The zirconium component is precipitated and calcined
233
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to oxide which is rechlorinated to high-purity zirconium tetrachloride
and reduced by the Kroll process.
The resulting zirconium sponge is crushed, compacted into con-
sumable electrodes, and vacuum melted in an inert atmosphere into ingot.
The ingot is used as consumable electrode to produce second-melt ingot,
which is machined and readied for fabrication.
The solvent used to remove hafnium from zirconium is methyl
isobutyl ketone (MIBK). The solvent is tighter than and is not miscible
with the aqueous solution. Hafnium thiocyanate dissolves preferentially
in the solvent as it passes upward through the column. The hafnium
rhiocyanate is stripped from the organic phase with sulfuric acid, at
which time it is converted to the sulfate. Hafnium is precipitated as
the hydroxide and is then roasted to the oxide containing about 99
percent hafnium oxide.
Hafnium oxide, mixed with carbon, surcrose, and a small amount
of water, is briquetted, dried, and then chlorinated in a vertical-shaft
ohiorinator. The resulting hafnium chloride vapors are purified by
injection into a molten mixture of potassium chloride (KC1) and sodium
chloride (NaCl), containing a small quantity of metallic sodium. High-
purity hafnium chloride is volatilized from the salt bath and is reacted
with magnesium in a Kroll-type reaction at about 800 C. When the
furnace has cooled, the resulting mixture of magnesium and sodium chlorides
and any excess magnesium is volatilized, and the residual hafnium sponge
is removed from the crucible.
Further refining of the sponge metal may be accomplished by
the iodide process, by electrorefining, or by electron beam melting.
Sources of potentially hazardous wastes include dilute filter
solutions containing ammonium thiocyanate (NH^CNO) and methyl isobutyl
ketone (MIBK).
234
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SIXTION 111
SiiCONDMY NON-H.ltKOUS SMlimNC AND Rlil-JNINC
'I'his section presents the results of investigations and analyses
of on-land disposal or storage of process and pollution control residuals
from the United States seconder mriferr.ms smelting and refining industry.
Characteristics of each indu..i, */ ....,».• eluding plant locations,
production capacities and smelting an I raining processes have been
identified and described.
Land disposed or stored /t.sj.uuais Deluding slags, dusts, and
sludges have been identified and characterized physically and chemically.
State, regional and national estimates have been made of the total
quantities of land disposed or st»re.J '•£;.> idua Is and potentially hazardous
constituents thereof.
Current methods employed by the secondary metals industry for
the disposal or storage of process and pollution control residuals on
land are described. Principal methods include lagoon storage of sludges
-ind open dumping of slags. Method., c. resiujal treatment and disposal
considered suitable for adequate he..itu a.id environmental protection
have been described. Finally, the costs incurred by typical plants in
each secondary smelting and refining category for current and environ-
mentally sound residual disposal or storage on land, have been estimated.
Secondary metals are categorized to include all metals and
metal alloys recovered from scrap and waste origins. Final products are
produced with chemical composition specifications derived according to
individual materials application requirements. Individual alloy products
are indistinguishable in quality from comparable grade materials made by
primary producers. The feasibility of producing specific grades metal
alloys at secondary smelters is, therefore, governed mainly by cost
competition with primary producers. Some of the techniques employed at
secondary smelters to meet alloy specifications often include the addition
of primary metals for dilution or alloying purposes. On the other hand,
much of the waste products generated at the secondary smelter often
contain metal values exceeding those present in the corresponding ore
concentrates and, thus presents a raw materials source for primary
smelting.
The difficulty in characterizing solid waste generated from
secondary metals smelting by SIC categorization lies in the fact that
secondary metals are produced mainly according to the nature of scrap
supply sources rather than specific SIC metal classifications. The
extent of waste produced from secondary smelting of each metal category
will differ greatly according to the metallurgical processing required.
The amount of waste requiring direct land disposal by the smelter will also
depend on whether there are demands for such waste products for further
235
-------
,,KUi.. <'•• lA'ory or alternate applications. Thus, an assessment of prevalent
operatjons within the individual secondary metals SIC classifications is
necessary so that pertinent subcategorization of hazardous waste sectors
within the SIC 3341 secondary metals smelting industry can be derived.
SIC .43412 -~ Secondary Smelting and Refining of Copper
As shown by the 1971 and 1972 production data summarized in tne
u.j Jiiroau of Mines-Minerals Yearbook (Reference 1), unalloyed copper
proa^non at secondary smelters accounts for only about 15 percent of
.such a product from scrap sources. Further, the refined copper produced
^>v secondary smelters is accounted for almost entirely by recovery from
new scraps. In the case of secondary copper recovered as alloy products,
brass mills account for about 65 percent of the total production. Brass
:fiij t products, however, are generated mainly from new scraps by melting of
appropriately formulated alloy heats. Waste dross generated from the flux
cover used in the melting process is shipped to primary or secondary smelters
for further metal recovery. Land-destined wastes within the SIC 33412
category are, therefore, attributed mainly to the production of alloy pro-
ducts from old scraps at secondary smelters. In addition, it should be
j:otco tliut significant quantities of tin, lead, zinc, and nickel are recovered
in the production of secondary copper alloys. About 47 percent of the
secondary tin recovered is in the remelting of brass and bronze scraps.
alC 33413--Secondary Smelting and Refining of Lead
Secondary lead recovery is one of the oldest known recycling
operations. Today, the volume of secondary lead recovery represents nearly
50 percent of the total lead production in this country. About 29 percent
jf The lead produced from secondary sources is in the form of refined
pure lead, the remainder is recovered in a variety of different lead alloys.
Antimonial lead used in storage battery plates represents by far the single
most important product of the secondary lead smelting industry, accounting
For 57 percent of the total secondary lead recovery. Another category of
lead alloys, consisting of varying mixtures of lead, tin. and antimony,
are known as white metals. Alloy scraps of these types are generally
grouped together for making up specified heats in the production of various
tvr-': HI'.-till, babbitt, and solder alloys. These white metal alloys account
for about 14 percent of the secondary lead market. Thus, for the purpose
of a.sst; ,sin*.' land-destined wastes within the SIC 33413 category, three
aitierent types of processing operations, corresponding to soft lead,
antimoniai lead, and white metal productions, must be considered. Second-
ary antimony recovered with lead alloys represents not only the entire
secondary antimony production but also about 63 percent of the total antimony
product ion in this country.
SIC 33413 -- Secondary Smelting and Refining of Zinc
By far the largest single smelting operation for secondary zinc
recovery is in the remelting of copper base alloys, accounting for 46 percent
236
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of the secondary zinc recycle and f>3 percent of metal recovery. About 13
percent of the scrap zinc recycle is an the form of metal value recovery as
zinc chemicals. Zinc recovery by remelting the various zinc-containing
alloy scraps accounts for about 54 percent of the secondary zinc smelting
operation. Scrap distil]at;on for recovery of pure zinc in the form of
either slab zinc or zinc dust accounts for the remaining 33 percent of the
secondary zinc production nwrk,1:. The secondary zinc smelting industry
is beset with air pollution (.on.,,,1 t, x/.emii, but very little land-destined
wastes are generated. Dust residues collected from air pollution control
devices generally have very h>,h zinc contents and are used in zinc chemical
productions principally ziiu okic "."• „;• accounts for the fact that a rather
high percent of the secondary ziiu- tccovery market is in the production of
zinc chemicals. in view of the fact that little solid wastes are disposed
within the SIC 33414 industry sector, a detailed evaluation of hazardous
waste generation in this industry category is not provided in this study.
SIC 33415 -- Secondary Smelting and Refining of Antimony,
Cadmium, Magnesium, Nickel, Tin and Titanium
As noted early in the assessment of copper- and lead-based alloy
smelting operations covered respei.i;i.
-------
Secondary magnesium is recovered only to a minor extent in about
six plants in the United States. Secondary smelting of magnesium is, in
addition, limited only to remelting, dilution and alloying operations. Very
little land-destined wastes are generated. Hazardous constituents are not
found in these wastes.
On the basis of the above considerations, quantitative) data on
h.i '...]',,. s wasre generation and disposal in the SIC 33416 category were,
t.(u-: -> IV -.2, not derived as a separate part of this study.
SIC 33417 -- Secondary Smelting and Refining of Aluminum
Secondary aluminum recycle is one of the most important segments
of t 'ue secondary nonferrous metals smelt4 ng industry category. Although a
large variety of aluminum alloys are produced by secondary aluminum smelters,
most of the alloys contain only minor amounts of alloying elements. Thus,
the secondary aluminum smelter is in direct competition with the primary
producer far more closely in product lines in comparison with other nonferrous
metals smelting categories. Aluminum is an easily oxidizable metal. Only
the, rtv.oval of magnesium and oxide impurities can be accomplished in
scc.mdary aluminum smelting. A majority of secondary aluminum smelters are
engaged only in remelting and demagging (i.e. removal of magnesium) of high
grade aluminum scraps. A number of smelters do have capabilities for
aluminum recovery from dross waste generated in remelting and foundry opera-
tions. Waste generation and disposal problems faced by high grade scrap vs.
dross smelters are far different. Separate assessments for each of these
s ure, therefore provided in this study.
On the basis of prevalent secondary nonferrous metals smelting
practices within the recycling industry, detailed hazardous waste assessiaents
and data compilations are provided in the following categories:
Secondary Copper Smelting (SIC 33412, 33414, 33416)
Secondary Soft Lead Smelting (SIC 33413)
Secondary Antimonial Lead Smelting (SIC 33413, 33416)
Secondary White Metal Smelting (SIC 33413, 33416)
Secondary Scrap Aluminum Smelting (SIC 33417)
Secondary Aluminum Dross Smelting (SIC 33417)
238
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1.0 SECONDARY SMELTING AND RLF1NINC OF COPPER (SIC 33412)
1.1 INDUSTRY CHARACTERIZATION
Recovery of copper from scrap is not conducted entirely at
secondary copper smelters. Of 1.1P million MT of copper recovered in
1<>72 28% was from primary producers, 44% from brass mills and 23% from
secondary smelters (Ref. 1) r*i^ rprnai ung 5% was reclaimed at chemcial
plants, foundries and manuf.i:turrti,, ^lid wastes from primary smelters
is handled elsewhere in this report. Brass mills engage in scrap melting
only and as a result produce nagiigibie waste.
There arc an estimated 40 secondary copper smelters in the
United States. Plant capacities vary widely from 500 MT per year to
about 80,000 MT per year. The typical pyrometallurgical plant produces
10,000 MT of copper alloy metals per year and operates 250 to 300 days/year.
This represents the median value taken from the plant size distribution
of pyrometallurgical plants. Plants employing electrolytic refining
typically produce 40,000 MT of copper per year although the smallest
electrolytic plant produces only 540 MT copper per year and the largest
approximately 80,000 MT per year. T-.ble 100 gives the estimated state
by state capacities for secondary copper smelters. Production capacity
is missing for 4 plants. According to the 1972 Minerals Yearbook
255,431 MT of copper was produced at secondary copper smelters in 1971
and 266,224 MT in 1972 (Ref. 1). Twenty-two percent of the 1.36 million
tons of scrap copper or J(>7,<143 MT was recovered at secondary copper
•.molic-rs in l')73 (lU'f. 2). The average soiling price of copper in 1973
wa.s .')!).:> conts/lb. Thus tho national sales of copper from secondary
sniolLer.-, may bo Obtimuted at, $350,083,000.
1.2 WASTE CHARACTERIZATION
This section contains descriptions of production technology at
secondary copper smelters and the resultant by products or wastes which
are either recycled directly, shipped to other smelters for further
metallic recovery or disposed of on land. Estimates are given for the
quantities of wastes and potentially hazardous constituents which are
disposed of on land either in lagoons or open dumps.
1.2.1 Process Descriptions
Pyrometallurgical. A generalized secondary copper and brass
smelting process waste flow diagram is shown in Figure 22. The typical
secondary copper producer will engage in the production of copper and
brass and bronze alloys by pyrometallurgical processes. Electrolytically
refined copper is more likely to be produced at the primary copper
refiner although -1 secondary plants produce electrolytic copper. The
major source of solid wastes in the secondary copper smelter is in the
blast furnace recovery of copper.
239
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It is important to note tha<; when high grade scrap is smelted
the welted scrap is sent directly to the reverbatory furnace rather than
to the blast furnace. In this case the slag is not disposed but recycled
for metal recovery in the blast furnace.
Blast furnace processing in the secondary copper smelting
pi.jtit is mainly intended for recycle of drosses, skimmings and slag
•va.fo accumulated from high grade scrap melting and smelting operations.
'!),„ bla.st furnace may be operated periodically only when sufficient
waste byproducts are accumulated for at least one week continuous furnace
operation. Only about 20 secondary copper smelting plants in the United
States have blast furnace scrap processing capabilities.
During cupola and blast furnace processing, low grade scrap
mixtures are melted with suitable flux chemicals, such as limestone,
borax and waste glass. Coke is charged for reduction of copper compounds
to the metal. The blast air is blown in through tuyeres near the bottom
of the furnace, liach heat of scrap and flux is maintained for about a
day. The molten metal and slag are tapped intermittently. A basic
difference between cupolas and blast furnaces lies in the fact that the
formei are not intended for copper compound reduction and so less fuel
and lower blast temperatures are used. The intermediate metal product
recovered is known as black copper. With blast furnace charges from
typical scrap sources, average composition of the metal is as follows:
Cu, 83.7$; Si), 3.7?»; I'b, 3.8%; Sb, 0.2%; Fe, 2.5%; Ni, 0.8%; S, 0.5%,
In, •!.8",. Black cc-ppor is further refined in converter and anode furnaces
whcrt' the impurity motul constituents are converted to oxides and removed
with tlid added flux us slug, The slag Is, in turn, recycled to the
lil.i-.i furauco \ar further copper recovery.
The "typical" plant may be assumed to have an annual total
copper alloy production volume of 10,000 metric tons (MT). Only about
1,500 MT black copper is produced annually during the overall smelting
process cycle. This black copper is consumed in-plant for brass and
bronze alloy production. The typical blast furnace daily production
capacity may be assumed to be 15 MT.
Electrolytic Refining. The principal difference between
secondaxy copper plants using strictly pyrometallurgical processes for
i-ofipii.r ,md those using the electrolytic refining is that the copper
i i-om the reverbatory furnace is further refined by first dissolving in
sulfuric and then electrolytically redepositing pure copper on the
cathodes of electrolytic cells. This process is shown as Figure 23.
The typical electrolytic .secondary copper plant is larger than the
plant.s which strictly employ pyrometallurgical processes <»nd will pro-
iiuec about .10.000 MT of copper per year. One electrolytic plant produces
about 80,000 MT copper per year.
242
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ANODE COPPER
(FROM SMELTER)
I
SLIMES
PROCESSED FOR
PRECIOUS METAL-<-
AND/OR NICKEL
RECOVERY
ELECTROLYTIC
CELLS
SPENT
ELECTROLYTE\TREAT
99.9% PURE COPPER
(CATHODE COPPER)
SLUDGE TO LAGOON
0.4 Kg/MT COPPER
Figure 23 SECONDARY COPPER ELECTROLYTIC REFINING
243
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Description of Waste Streams. The typical blast furnace dust
will have the following chemical composition: Zn, 58-61%, Pb, 2-8%; Sn,
5-15%; Cu, 0.5%; Sb, 0.1%; and Cl, 0.1-0.5%. Because of the high zinc
content, the dust can be recycled for zinc compound manufacture largely
'i;> paint additives. This waste byproduct is thus sold to chemical and
paint manufacturers and does not require direct disposal by the smelter.
Pyrometallurgical. The t>pical blast furnace slag will have
the following composition: FeO, 15-20%; CaO, 7-11%; SiO?, 30-34%;
'-l/V 4'6%» Cu- J'6%; Sn> °>n> Pb» c-s%' Zn> 1%- Analyses of samples
obtained on this study are given ir. Appendix A. At the assumed "typical"
plant, about 3,500 MT blast furnace sJag is produced annually. During
the period of blast furnace operation, this slag waste is produced at
the rate of 35 MT per day. This waste is disposed by either on-site or
off-site by open dumping. The slag material is normally tapped and cast
in crucibles at about 100 kg sizes. This material is present mainly in
a silicate type chemical structure with high density, hardness and
impact strength. Although this slag would not be expected to easily
leach solubility tests as described in Appendix B showed significant
solubility of zinc, cadmium, copper, and lead. For this reason blast
furnace slag from secondary copper smelters must be considered poten-
tially hazardous at this time.
Electrolytic. Lime treatment of spent electrolyte generates
a predominantly lime sludge containing significant concentrations of
nickel, zinc, copper, chromium and cadmium. This sludge is generated at
a rate of approximately 1.3 Kg wet sludge/MT copper or approximately 0.4
Kg. dry weight per MT copper. Appendix A contains chemical analyses of
residuals from secondary copper smelters.
1.2.3 Waste Quantities. Table 101 summarizes the generation factors
for the land disposed residuals from the secondary copper smelting and
refining category (i.e. blast furnace slag, electrolyte sludge). This
table also gives typical concentrations of potentially hazardous consti-
tuents. Table 102 gives the yearly quantity of wastes and hazardous
constituents thereof for a typical pyrometallurgical plant producing
10,000 MT of brass and bronze ingot or anode copper per year. Table 102
^ives these values for a typical secondary electrolytic copper plant
producing 40,000 MT of electrolytic (i.e. cathode) copper.
Using the waste generation factors and concentration factors
contained in Table 101 and and information on the geographic distribution
of smelter types, capacities and pollution control methods estimates
have been made for the quantities of land disposed or stored wastes and
hazardous constituents thereof from secondary copper smelting and refining.
These estimates are given for 1974, 1977 and 1983 on a state-by-state,
regional and national level in Table 103.
244
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Principal products of secondary copper smelters are copper-
base alloys in the form of brass and bronze ingots. Production figures
for secondary copper smelters are therefore most appropriately represented
by the annual statistics for brass ana bronze ingot deliveries as compiled
by the American Bureau of Metal Statistic (ABMS). Brass and bronze
produced from secondary copper recycle and reused in captive brass mills
are not included among these delivery figures. Over the past eleven-
year period from 1963 to 1971s, i.nnu il de.'ivery volumes showed only
slight variations with the a\ assumed that the secondary
copper smelting product volumes ii> the years i977 and 1983 will not be
significantly different from i.in"ivr< ,. >-l >: on figures. Significant
changes in secondary copper smelting process technology are not expected
for the short term. Hazardous waste data for the years 1977 and 1983
are therefore projected to be essentially the same as those for 1974.
States producing the largest quanfity of potentially hazardous slag
waste are Georgia, Illinois and t'ermsyivanja. Sludges are generated
only from electrolytic refiners and are of .significance only in Georgia
and Illinois
1.3 TREATMENT AND DISPOSAL TECHNOLOGY
Current Waste Treatment Ana Disposal Practices
Currently slags from pyrometallurgical and electrolytic plants
are open dumped. Electrolytic plant wastewater sludges is settled in
unlined lagoons. As previously discussed, these wastes are considered
potentially hazardous at the present time. Levels of treatment and
technology and environmental adequacy thereof for these wastes are
described in the following sections.
1.3.2 Present Treatment and Disposal Technology (Leve1 I)
Pyrometallurgical. At the present time blast furnace slags
generated at a rate of 350 kg/MT of copper production are open dumped.
This practice is inadequate if leached heavy metals including zinc,
copper and lead percolate through permeable soil to groundwater.
Electrolytic. Spent electrolyte wastewater treatment sludge
generated at a rate of 0.4 kg/MT copper (1.2 kg/MT wet weight) is pre-
sently disposed of in unlined lagoons. This practice is inadequate if
heavy metals including zinc, copper and lead percolate through permeable
soils to groundwater. Electrolytic plants also generate blast furnace
slag in preparing copper for electrolytic refining which is of the same
nature as that from pyrometallurgical plants.
249
-------
J.3.3 Best Technology Currently Employed (Level II)
Pyrometallurgicftl. Open dumping of blast furnace slag is the
only disposal practice and is therefore Level II technology.
Electrolytic. The use of unlined lagoons for spent electrolyte
wastewater sludge disposal and open- dumping of blast furnace slag is
Level II technology,
1.3.4 Technology to Provide Adequate Health and Environmental
Protection (Level III)
Pyrometallurgical. Solubility tests as described in Appendix B
indicated that zinc, copper, lead and other heavy metals may leach from
blast furnace slag. In this event, in permeable soils, the ground would
need sealing with bentonite or other sealant and collection ditches in-
stalled to collect runoff.
Electrolytic. Discarded blast furnace slag would have to be
handled in the same way as slag from strictly pyrometallurgical plants
(.i.e. surface sealing and collection of runoff). Lined lagoons would be
required to prevent percolation of leached heavy metals through per-
meable soils to groundwater.
Features of Levels I, II and III treatment and disposal tech-
nology are summarized in Tables 104 and 105.
1.4 COST ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental
protection were described. The costs of implementing this technology
for typical plants is considered in this section. The typical secondary
copper pyrometallurgical processing plant has an annual production of
10,000 MT of copper alloy per year. The blast furnace is operated
about 100 days per year. About 1500 MT of "black copper" is produced as
an intermediate product.
The typical electrolytic plant produces 40,000 MT of 99.9%
pure electrolytic (i.e. cathode) copper per year and operates 360 days.
1.4.1 Cost of Present Treatment and Disposal Technology (Level I)
Pyrometallurgical Refining. Slag produced totals 3,500
MT/year or 35 MT/operating day. Equipment and personnel requirements
are 200 hours each of truck and front loader time, and 40 hours of
bulldozer time/year.
The density of the slag is 2000 kg/m . An area of .4 ha is
needed to store 20 years of waste assuming a height of 10 m. The dump
site is located on semi-industrial land on the outskirts of an urban
area.
250
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There are no other disposable wastes. Costs of Level I treatment
and disposal technology are given in Table 106.
Electrolytic Refining. The wastes generated consist of
56.7 MT of slag and 0.11 m of sludge, daily. The slag is deposited at
an on-sitc dump. Annual equipment and personnel requirements for disposing
of the slug are 1,000 hours of front loader and truck time, and 200
hours of bulldozer time at -i-e ;!r • 'JJT; , The slag dump, sized to
contain 20 years of waste, extends over a 1.6 ha area.
The sludge is piped to a holding lond designed to contain 20
years of waste. It characterv~t: ":
Volume 1,320 m3
Bottom Width 15.2 m
Top Width ..M. I m
Bottom Length .SO.4 m
Top Length 38.4 m
Total Depth 2 m
Depth of Excavation 1.1 m
Circumference 135 m.
Dike Volume 581 m''
Dike Surface 1,269 m^
Total Width 32.8 m
Total Length 48.0 m
Required Area 0.16 ha
Level 1 capital and annual costs are presented in Table 107.
1.4.2 Cost of Best Technology Currently Employed (Level II)
The technology and costs of Level II are the same as Level I.
1.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III)
Pyrometallurgical Refining. The ground is sealed at the slag
dump; collection ditches, pump and piping are installed. Costs incurred
are:
Capital cost
Land scaling $ S.OOO
Collection UUohos i)o()
Ptunp 1} piping 7,
-------
TABLE 106
COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
SECONDARY COPPER PLANT (PYROMETALLURGICAL REFINING)
Cajipment
Truck (10%)
frontloader (10%)
Bulldozer (2%)
Survey
Land
Total
Capital Cost
Slag
$ 2,500
2,000
320
250
1,580
$ 6,650
Annual Cost
Land
\-f,ortizat.Ion
Construction
Equipment
Operating Personnel
Repair and Maintenance
Construction
Equipment
Energy
Fuel
Electricity
Taxes
Insurance
Total
$ 160
30
770
4,805
240
585
60
40
65
$6,755
256
-------
SECONDARY COPPER PLANT (ELECTROLYTIC REFINING)
Capital Cost
Sludge Slag
Lagoon
Site Preparation
Survey $ 100
Test Drilling 490
Sample Testing 250
Report Preparation 1,200
Construction
Excavation § Forming 775
Compacting 1,075
Fine Grading 570
Soil Poisoning 165
Transverse Drain Fields 150
Land 635
Equipment
Truck (50%) $12,500
Frontloader (50%) 10,000
Bulldozer (10%) 1,600
Dump
Survey 1,190
Land 7,515
Total $5,410 $52,805
Annual Cost
Land $ 65 $ 750
Amortization
Construction 555 140
Equipment - 3,830
Operating Personnel - 24,030
Repair and Maintenance
Construction 80
Equipment - 1,205
Energy
Fuel - 2,930
Electricity - 295
Taxes 15 190
Insurance 55 330
Total $ 770 $33,700
257
-------
Annual cost
Construction Amortization $ 1,040
Equipment Amortization 1,260
Construction Repair 5
Maintenance 270
Equipment Repair §
Maintenance 395
Energy
Electricity 30
Insurance 170
Total $ 3,165
Electrolytic Refining. The ground is sealed at the slag dump;
collection ditches, pump and piping are installed. The sludge is piped
to a lined lagoon. Level III capital and annual costs are shown in
Table 108.
Costs of Level I, II and III treatment and disposal technologies
are summarized in Table 109 for secondary copper from pyrometallurgical
refining and in Table 110 for secondary copper from electrolytic refining.
Costs are given in dollars per metric ton of dry and wet waste and
dollars per metric ton of copper product. Costs for each of the types
of wastes (i.e. slag from pyrometallurgical refining and sludge and slag
from electrolytic refining) at each level of technology are also expressed
as percentages of metal selling price.
Total industry costs for Levels I, II and III technologies
have been estimated in 1973 dollars and also are in Tables 109 and 110
for pyrometallurgical and electrolytic categories respectively. The
industry annualized cost for Levels I and II disposal of slag from
pyrometallurgical plants is estimated as $210,000 or 0.05% of estimated
liJ-'j i,ales. The cost of Level III technology (i.e. adequate for environmental
protection) is estimated as $300,000 or 0.08% of estimated 1973 sales
value.
The industry annualized cost for Level I and II treatment and
dl.i j.«- i of slag and sludge from electrolytic refiners is estimated as
$120,000 or 0.06% of estimated national sales. The cost of Level III
technology is estimated as $170,000 or 0.08% of estimated 1973 sales.
258
-------
"-ASLE 108
COST OF LEVIil. 111 I'RKATMENT AND DISPOSAL TECHNOLOGY
SECONDARY COITliR PLANT (l-LliCTROLYTIC RHHNINC)
Capital Cost
onstruction
Lagoon Liner
Land Sealing
Collection Ditches
equipment
Pump and Piping
Sludge
$5,360
Total
$5,360
Slag
$38,000
2,090
11,410
$51,500
Annual Cost
Land
f ;nortization
Construction
Equipment
Operating Personnel
Repair and Maintenance
Construction
Equipment
Energy
Fuel
Electricity
Taxes
Insurance
$ 620
160
Total
$ 4,650
1,815
1,205
570
95
515
$ 8,850
259
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2,0 SECONDARY SMELTING AND REFINING OF LEAD (SIC 33413)
M INDUSTRY CHARACTERIZATION
There are approximately 82 secondary lead smelters in the United
State-, Production capacities of individual plants varies from about 2000 KTT
of lead per year to about 40,000 MT. The geographical distribution of
i;ecc;rj.'iiy lead smelters by state and EPA region is given in Table 111.
Capacities of 30 secondary lead smelters were not available. The 1972
minerals yearbook gives the 1972 production of secondary lead as approximately
553,000 MT inclusive of about 7,000 MT recovered at primary smelters
(Ref. 1). The production of secondary lead has remained fairly constant
ar about 540,000 MT/year since 1969. The 1973 recovery of lead at secondary
smelters was 593,568 MT valued at $213,166,000 (Ref. 2) which represents
an estimate of national sales.
Scrap battery waste comprises by far the most important supply
source for recovery of secondary lead. The recovered lead is most frequently
returned to use in battery manufacture. Battery plate grids are fabricated
of hard lead containing about 5% antimony. The lead oxide paste used in the
plates must be manufactured from antimony-free soft lead. Both soft and
hard lead can be produced from scrap battery waste. Soft lead is recovered
by reverberatory smelting, and the recovered metal is converted to lead
oxide either by the Barton process or by a milling process. The reverbera-
tory slag is then smelted in blast furnace for antimonial lead recovery
using scrap iron as reducing agent. Alternatively, scrap battery waste may
be smelted directly in the blast furnace and only hard lead is produced.
l.i WASTL CHARACTERIZATION
This section contains descriptions of production technology at
secondary lead smelters and the resultant byproducts or wastes which are
either recycled directly, shipped to other smelters for further metallic
recovery or disposed of on land. Estimates are given for the quantities
of wastes and potentially hazardous constituents thereof which are dis-
posed of on land either in lagoons or open dumps.
2.2,1 Process Description. Secondary lead smelting operations
differ somewhat according to the type of products produced. The three
major typical operations product are the following:
Hard lead (antimonial lead) - contain approximately 5%
antimony
Soft lead - pure lead
White Metal - lead-tin alloys (approximately 20% tin,
80% lead)
262
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111
GEOGRAPHIC DISTRIBUTION of SECONDARY LEAD PRODUCTION CAPACITY
STATE
ALA.
CALIF.
COLO.
DEL.
FLA.
GA.
ILL.
IND.
KY,
LA.
MO.
MASS.
MICH.
MINN.
MISS.
»
-------
I
Hascu uh the median value of plant size distribution the "typical" secondary
lead smeJcar consuming scrap battery waste may be assumed to have an annual
production volume of 20,000 metric tons total lead products. Any or all of
the above products may be produced at individual plants. The distribution of
produc ts between soft lead, antimonial lead and lead oxide may vary widely.
There are about 10 plants capable of producing both soft and antimonial
lead as shown in Figure 24 and another 10 plants which produce only anti-
monial lead as shown in Figure 25. The "typical" plant is assumed to produce
either 20,000 MT antimonial lead annually, or 10,000 MT antimonial lead
jitd 10,000 MT soft lead and lead oxide annually. Lead recovery from dis-
carded storage battery plates represents well over 50% of secondary lead
recovered. Individual plants may have either reverbatory furnaces for soft
lead production or cupola or blast furnaces for antimonial lead production
or a combination thereof for both soft and antimonial lead production.
Both the revcrberatory and the blast furnaces are continuously operated.
Assuming a normal ten percent shutdown, the daily total lead production
volume would therefore amount to about 60 MT.
Secondary white metal smelting consists in the recovery of lead-
tin allgys from solder, babbit, and type metal scrap waste. The metal
1ecovered from this scrap source consists of tin-lead alloys averaging
about 20% tin and 80% lead, and it is known as white metal. For high
grade scrap, white metal is recovered simply by remelting in a pot furnace.
ror low grade drosses and skimmings, the metal is recovered by smelting
in a reverberatory furnace with limestone, silica and iron scale flux.
Coal fines are charged as fuel and reducing agent.
White metal is widely recycled in relatively small quantities
by pot furnace melting. Reverberatory smelting of drosses and skimming
front which metal recovery is performed in only a few secondary smelting
plants. All of the plants in this category are of nearly the same size
and the "typical" secondary white metal reverberatory smelting may be
issumed to have an annual production capacity of about 9,000 metric tons.
Kevcrberatory white metal smelting is operated continuosly. Daily pro-
duction volume is about 27 MT.
2.2.2 Description of Waste Streams.
Soft and Antimonial Lead. Figure 24 illustrates processes for plants
',cKich predi'i'c both soft and antimonial lead. In making soft lead,
h>product -slag is sent to a blast or cupola furnace as input for antimonial
lead [i.e., hard lead) production. The emissions from the reverberatory
furnace are collected in a baghouse and immediately recycled. At the
present time, therefore, there is no land disposed waste except for one
plant in Pennsylvania which uses a lime SO- scrubber system resulting in
a sulfate sludge which is land disposed.
264
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SCRAP LEAP SCRAP IRON FLUXES
DUST RECYCLE-
r
SLAG RECYCLE •
CUPOLA FURNACE
ANTIMONIAL LEAD
1000kg
(IMT)
r SLAG TO DUMP1
225 kg
NUMERICAL VALUES IN kg/MT LEAD PRODUCT
Figure 25 SECONDARY ANTIMONIAL LEAD SMELTING
266
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SO- Sludge. It is expected that sulfur dioxide emissions
venerated from lead sulfate present IP the battery plate paste and
residual sulfuric acid in the ,>c..ap will be controlled by 1977 because
~i more stringent air pollution /'e.tJLations. Lime scrubbing is most
.ommonly used for S0_ emission control. For the "typical" plant where
10,000 MT is produced annual.!-/ -.bni•; i,500 MT lime scrubber sludge
'hickened to about 30% solid cor^env. vou.ld be thus generated. This will
.-mount to about 45 kg dry wtr; : - <-* "TP T'jd?e per metric ton of lead
roduced. Analyses of an SO, ; <: i,t ; *.r ;• •><.:. T- as given in Appendix A
;ave high concentrations of ca-.tr- . • '•••-* , lead (53,000 ppm) and
n her toxic metals. In solubi^; v • •?•<•; de?Tibed in Appendix B cadmium
<,n ppm Cd; 2.5 ppm Pb.
:>r these reasons SO- scrubber sJu('<'<: '« considered potentially hazardous.
Blast Furnace Slag. In H "Or MT blast furnace slag is gen-
erated for 10,000 MT antimonia] !e*.d produced and disposed of in open
Jumps. This amounts to 450 kg of sl.vg per MT of combined soft and
antimonial lead production. The hard siliceous blast furnace slag is
expected to have the following major chemical composition: FeO, 34.8%;
CaO, 15.2%; Si02> 28.4%; and Pb, i,c" Trac netals include zinc,
. 'pptr and antimony. In soluhM-1 . .,5 •"<= >~:bed in Appendix B no
. .'tals solubilized to an extent of a.\ least 1 ppm. This slag is therefore
not considered potentially hazardous.
AntimoniaJ Lead. With direct blast furnace smelting of scrap battery
waste to produce antimonial lead (soe Figure 25), the sulfur constituent
in the scrap waste is scavenged by iron resulting in the formation of
both a matte and a slag together with the molten metal. About 25% of
this matte and slag mixture is recycled as flux for subsequent smelting.
For the annual production of 20,000 MT antimonial lead, the disposable
slag waste amounts to 1,800 MT annually and the matte waste, 2,700 MT
annually. The quantities of disposed matte and slag are 135 and 90
kg/MT of lead product respectively. Typical chemical composition of the
slag is as follows: FeO, 41.8%; Si02> 29.8%; CaO, 11,75%; and Pb, 0.6%.
Typical chemical composition of the matte is as follows: Fe, 60.7%; Pb,
4.5%; and S, 15%. The former is typically a mixed silicate and the
latter consists mainly of a mixed sulfide. The slag and the matte are
tapped together and cast into a solid mass before land disposal. The
slag from antimonJal lead production as described previously did not
solubilize trace metals to greater than J ppm and is not considered
potentially hazardous.
White Metal. As shown in Figure 26 reverberatory furnace flue dust is
controlled by baghouses. Flue dust composition will vary according to
the nature of the scrap being smelted. An example of analysis is as
follows: Sn, 27.7%; Sb, 0.5%; Cu, 0.06%; Pb, 17.5%; S, 6.0%; Zn,
13.3%; and Cl, 4.8%. The flue dust, because of the rather high zinc
content, is not directly recyclable to the smelting furnace. It is sub-
jected to leaching with dilute sulfuric acid for zinc removal, and the
267
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DROSSES AND
SKIMMINGS
SLURRIEO DUST
'RECYCLE
ZINC
REMOVAL
BY ACID
LEACHING
LEACHING \
SOLUTION )
TO LAGOON /
IRON
SCALE
REVERBERATORY
FURNACE
REFINED METAL
1000kg
(IMT)
FLUXES
r
SLAG RECYCLE-
SLAG TO
DUMP
166kg
NUMERICAL VALUES IN kg/MT LEAD PRODUCT
Figure 26 SECONDARY WHITE METAL SMELTING
268
-------
residue is recycled. The leaching solution in neutralized and stored in
a lagoon. For the "typical" plant producing 9,000 MT white metal, about
50 MT fluedust is generated and treated annually. This amounts to
5.5 kg/NTT lead produced.
If it is assumed that only the zinc and chloride contents of
the- : »ue dust is leached and disposed, the amount of waste disposed by
Ufci'.uing (Figure 25) would amount: tc, 9 MT for the typical plant, or 1
uroduct. It may be noted t^.i,; ,.., ..... ,£], only 4 plants are engaged
: •- ••!>. reverberatory furnace smelting oi j.ccudary white metal. Flue
, u . s shipped to one plant in New .ie,ac> tut zinc leaching, and 54 MT
-;r jmlfate is recovered annually ftou . h:-. -encr solution. This represents
:-. ' - % recovery of zinc value in the fiue dust from these 4 plants.
The reverberatory slag is generated at the rate of about
4,JCiO 4T per year (166 kg/NTT of product) for the "typical" plant with
9,000 MT white metal annual production volume. The slag composition
will vary somewhat according to the quality of scrap being smelted. As
an example scrap with a high tin copper content will yield a slag with
significant copper concentration. A typical set of sample analysis data
• s 'ollows: FeO, 15%; CaO, 15%; S?0 ?5%; A^CL, 5%; Sn, 2%; and
• ,%. This slag waste is tappoj -,i,.i ".us* iu a large mass, and it is
in a hard silicate form. On-site dumping is the most common current
disposal method.
Chemical analyses of samples collected from secondary lead
smelters are given in Appendix A.
2.2.3 Waste Quantities
Soft Lead and Antimonial Lead. Table 112 summarizes the
generation factors for the land disposed residuals from secondary lead
smelters producing soft lead and antimonial lead (i.e. blast furnace
slag and S09 scrubber sludge). This table also gives typical concentrations
of pOLfci.ciatiy hazardous constituents. T^bJ« 113 gives the yearly
quantity of wastes and hazardous constituents thereof fcr a typical
plant producing 10,000 MT of soft lead and 10,000 MT of antimonial lead
per year.
Antimonial Lead. Table 112 also summarizes the generation
factors for the land disposed residuals (i.e. blast furnace or cupola
furnace slag) from secondary lead smelters producing only antimonial
lead (i.e. hard lead). Table 113 gives the yearly quantity of slag and
hazardous constituents thereof for a typical plant producing 20,000 MT
Of antimonial lead per year.
White Metal. Table 112 summarizes the generation factors for
the land disposed residuals from secondary lead smelters producing only
white metal. Table 113 gives the yearly quantity of reverbatory furnace
slag from a plant producing 9,000 MT of white metal per year.
269
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Using the slag and SO- sludge generation factors for each of
tin- plant type;; previously described (i.e. soft lead and antimonial
lead, .'int. imonjal lend, white metal) and the estimated production capacities
of i ho various state:; i'or each of these types of metals estimates were
made I m' the total quantities of land disposed sJags ana sludges. These
estimates arc given in Table 114 for 1974, 1977, and 1983.
For the purpose of hazardous waste disposdl assessment, the
secondary lead smelting sector may be subcategorized into soft lead,
antimonial lead, and white metal smelters. Annual production volumes in
each of these categories have remained essentially constant for the most
recent five years from 1969 to 1973, for which statistical data are
available from the U.S. Bureau of Mines (Ref. 1). Antimonial lead
production, which is closely related to storage battery recycle, had
exhibited a steady increase in the annuaJ production rate prior to 1969
and again showed an appreciable gain in 1973. In view of the drop in
automotive vehicle production experienced in recent years, a further
j'.rowth in antimonial lead production would appear unlikely at least for
the- next low years. Waste projections associated with the secondary
lead blast furnace slag in both 1977 and 1983 may therefore be assumed
to ijf unchanged from current figures as presented in Table 114.
SuJ fur dioxide coiitrol for secondary lead smelters was practiced
at oply one location in Pennsylvania in 1974. This produced 3,000 NTT of
lime sludge with potentially hazardous constituents. As the S00 air
emission standards are being implemented, wet lime scrubbers would be
required for all soft load smelters with installation by 1977. This
accounts for the large projected increases in sludge in 1977 as compared
i.> 19'!. 1 urther changes by 19K3 is not expected. Thus, the air emission
«.O.IM\>! scrubber sludge waste projections for the secondary lead industries
in both 1977 and 1983 arc given in Table 114.
States with the largest quantities of total and hazardous land
disposed wastes are California, Indiana, New York, Pennsylvania and
Texas.
TRIiAIMhNT AND DISPOSAL TLCHNOI.OCY
uurvcnt .Waste '!"reiLt:J:!!fJ^ an^ Disposal Practices
At the present time, all discard slags from secondary lead
smelt inj1 are disposed by open dumping on land.
All of the discard slags are disposed of as hard large chunks.
With the exception of the antimonial lead matte which is a mixed sulfide,
all discard slags are primarily mixed silicates. All slags contain
significant concentrations of potentially hazardous constituents, including
lead, antimony and copper. However, as previously described, leaching
of these metals was not observed to a significant extent in solubility
tests. Open dumping is therefore considered environmentally adequate in
272
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view of available data. Leaching solution from acid leaching of baghouse
dust and scrubber sludge are currently disposed of in unlined lagoons.
They are considered potentially hazardous and are discussed in further
detail in the next section.
2.3.2 Present Treatment and Disposal Technology (Leve11)
Leaching Solution. At the present time leaching solution
from acid leaching of baghouse dust from white metal smelting is kept in
unlined lagoons (see Figure 25) and is therefore Level I practice.
Sulubi1izution of heavy metals including cadmium and lead can be expected
and the use of unlined lagoons is therefore considered inadequate.
Scrubber Sludge. As discussed previously, significant quantities
of lime sludge are generated if SO scrubbers are installed on secondary
lead .smelters. Sludges are normally disposed of in unlined lagoons.
This sludge contains significant concentration of cadmium and lead which
can solubilizc. Tims the use of unlined lagoons is not environmentally
adequate.
2.3,-"- Best Technology Currently Employed (Level II)
Leaching Solution. The use of unlined lagoons is the only
disposal mode employed and is therefore Level II technology. This
practice is environmentally inadequate as discussed previously.
Scrubber Sludge. The uso of unlined lagoons is the only
il i sno.-.a I mode employed uiul is therefore Level II technology. This
l>rai-i i (-c i;, onv j ronmeutu 1 ly inadequate as discussed previously.
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.!.-1.l Coat of I'rpapia Treatment iwd Ulttpoaal Teclmolojiy (Level I)
_ - -- „ ---- -— . -___- -- „. ------- „. „„-• i, . i. - r J — ~- — — - --- . . , .II-.MI^..,!.!. . -- . — •*•* >
Aiitimoinul^ and Soft Lead Smelting. The typical plant producing
both hard and soft lead characteristically produces 10,000 MT of soft lead
and 10,000 MT of hard lead per year and operates 333 days/yr.
Scrubber Sludge. As previously discussed it is expected that
the control of SO- emissions from secondary lead reverbatory furnaces by
1983 will result in a lime sludge. The typical plant producing 10,000 MT
of hard lead and 10,000 MT of soft lead per year will generate 1500 MT of
sludge with a 30 percent solid content which is hauled-by tank truck to an
on-site lagoon. The sludge density is about 1250 kg/m . Thus 1200 m of
sludge are disposed each year.
The lagoon is designed to hold 20 years of waste. Its design
characteristics are:
Volume 30,000 m"5
Bottom width 66 m
Top width 78 m
Bottom length 132 m
Top length 144 m
Total depth 3 m
Depth of excavation .8 m
Circumference 457 m,
Dike volume 7,443 m
Dike surface 6,605 m
Total width 93 m
Total length 159 m
Required area 1.5 ha
Four test holes arc drilled and 8 soil samples analysed as part of
site preparation.
One tank truck hour/day is assigned to haul the daily sludge
produced, 4.1 m , to the lagoon.
Capital and annualized costs are given in Table 116.
Antimonial Lead. The plant has an annual capacity of 20,000 MT
operating 330 days/year. Blast furnace smelting results in 1,800 MT of
slag waste and 2,700 MT of matte waste annually. The resultant daily waste
production is 5.45 MT and 8.1 MT, respectively. There are no costs attrib-
utable to hazardous waste disposal since the slag is not considered hazardous.
White Metal. Annual production is 9000 MT based on 333 operating
days. Waste generated consists of 1500 MT of slag and 50 MT of flue dust.
The dust is shipped for reclamation; waste slag is not hazardous.
279
-------
Table llf>
COST OP UiVlii, I TIOiATMliNT AND DISPOSAL IV.CHNOLOCY
SECONDARY ANTIMONJAL AND JJOI-'T LI'AD
Capital Cost
Lagoon
Site Preparation
Survey $ 940
Test Dri1 ling 980
Sample Testing 500
Report Preparation 1,500
Construction
lixcavation (, Forming 9,900
Compacting 14,325
Vine Grading 2,970
Soil Poisoning 565
Transverse Drain Fields 1,285
Land 5,935
Equipment
Tank Truck (12.5%) 5,000
Total $43,900
Annual Cost
Land $ 595
Amortization
Construction 3,825
Equipment 795
Operating Per5onncl 4,045
Repair and Maintenance
Construction 870
Equipment 250
linergy
Fuel 535
Electricity 55
Taxes 150
Insurance 440
Total $11,560
280
-------
J.4.2 Cost of Best Technology Currently Employed (Level II)
Antimonial and Soft Lead Smelting. The technology and costs of
Level II are the same as Level I for disposal of scrubber sludge.
Antimonial Lead . There are no hazardous wastes and therefore
no attributable costs.
White Metal. The technology and costs of Level II are the same
as Level I.
2•4•3 Cost of Technology to Provide Adequate Health and Environmental
Protection (Level III).
Antimonial and Soft Lead Smelting. This entails lining the
lagoon used for scrubber sludge disposal. Associated costs are:
Capital Cost
Lagoon liner
Total
Annual Cost
Construction Amortization $ 5,965
Construction Maintenance § Repair 1,545
Insurance 515
Total $ 8,025
Costs of Level I, II and III treatment and disposal technologies
are summarized in Table 117 for sludge from secondary antimonial (i.e.
"hard") lead and soft lead smelting and refining. Costs are given in
dollars per metric ton of lead product. Costs for each level of technology
are also expressed as a percentage of metal selling price.
Total industry costs for levels I, II, and III treatment and
disposal technologies have been estimated in 1973 dollars and are presented
in Table 117. The annualized industry cost for Levels I and II treatment
and disposal technology is estimated as $100,000 which is 0.16% of
estimated 1973 sales value. The cost of Level III technology (i.e.
necessary for environmental protection) is estimated as $170,000 or
0.27% of estimated 1973 sales value.
281
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3.0 SECONDARY SMELTING AND REFINING OF ALUMINUM (SIC 33417)
.^.1 INDUSTRY CIIAKAC'ir.KlZATlON
The secondary aluminum smelting and refining industry is character-
ized by a large number of relatively small operations in urban areas through-
out the United States. Many of these plants are engaged in only high grade
scrap remelting with little attendant wastes. Plant capacities vary from
1000 MT per year to 55,000 MT per year with about 20,000 MT capacity for the
typical capacity of a low grade scrap recovery operation.
Table 118 gives the estimated state, regional and national dis-
tribution of secondary aluminum smelters. Production capacities of individual
plants is held proprietary by the industry. The Aluminum Recycling Association,
Washington, D.C., provided information on EPA regional weekly production
capacities. Based on 50 weeks per year operation, estimates of EPA regional
and national production capacity are given in Table 119.
National production of aluminum from secondary smelters for
1971 and 1972 is given as 740,623 and 857,964 metric tons, respectively
(Reference 1). The recovery of aluminum from secondary smelters in 1973
was 943,314 MT at an average value of 29 cents/lb (Ref. 2). Thus the value
of national sales are estimated as $601,834,000.
3.2 WASTE CHARACTERIZATION
This section contains descriptions of production technology
;it sccond.iry aluminum smelters and the resultant byproducts of wastes which
are either recycled directly, shipped to other smelters for further metallic
recovery or disposed of on land. Estimates are given for the quantities
of wastes and potentially hazardous constituents thereof which are disposed
of on land either in lagoons or open dumps.
3.2.1 Process Description. A generalized secondary aluminum smelting
process flow diagram is shown in Figure 27.
Scrap melting. High grade aluminum scrap can be easily recycled
by remelting in pot or rotary furnaces. The smelting of low grade scraps
and foundry drosses is performed with reverberator/ or rotary furnaces.
Fluxing agents and alloying agents are charged with the scrap. Magnesium
is removed by A1F_ or Cl_ demagging. Degassing is susually achieved
simultaneously with the demagging operation. Upon completion of the
smelting operation, the fluxing agent is skimmed and the purified metal melt
is poured and cast into ingots. Common salt and potash mixtures are normally
employed as fluxing agents.
283
-------
TABLE 118
GliOCRAI'HICAL DISTRIBUTION OF
SECONDARY ALUMINUM SMELTERS
STATE
Alabama
Arkansas
California
Connecticut
Delaware
Florida
Illinois
Indiana
Kansas
Kentucky
Maryland
Massachusetts
Michigan
Missouri
Nebraska
New Jersey
New York
Ohio
NO. OF
PLANTS
6
3
16
4
1
1
15
7
1
1
2
1
3
2
1
3
9
18
284
-------
TABLE 118 (cont.)
STATE
Pennsylvania
South Carolina
Tennessee
Texas
Virginia
Washington
Wisconsin
NO. OF
PLANTS
6
2
1
2
1
2
1
EPA REGION
1
11
111
IV
V
VI
VII
VIII
IX
X
NATION
NO. OF
PLANTS
5
12
10
10
45
5
4
0
16
2
109
285
-------
TABU; 119
REGIONAL AND NATIONAL PRODUCTION CAPACITY
TOR SECONDARY ALUMINUM *
I'.l'A REGION i'KODUCTION CAPACITY
(METRIC TONS/YEAR)
1 5 II 82,220
III 29,030
IV 62,600
V 520,500
VI l\ VII 58,510
IX 72,580
NATIONAL 825,440
Callul.itud from Data Supplied By Aluminum Recycling Association,
March, 1975. Data is Combined For Some Regions to Avoid Disclosure
<>f I'roprietary Information.
286
-------
DEMAGGING
CHEMICALS
FLUXES
SCRAP
WET
..SCRUBBER
REVERBERATORY OR
ROTARY FURNACE
SLUDGE TO
.LAGOON 75 kg
DROSS
DROSS
PROCESSING
J 100 kg
REVERBERATORY OR
ROTARY FURNACE
REFINED ALUMINUM
1000kg
(IMT)
-------
Many of the secondary aluminum smelters in operation in the
United States are engaged in little more than high grade scrap remelting,
giving rise to little need for smelting waste disposal. For the purpose
of assessing solid waste disposal in aluminum scrap smelting, the "typical"
plant -nay be assumed to engage in low grade scrap recycle with fluxing
and debugging capabilities. Based on the median value of plant size
distribution the plant production volume is assumed to be about 20,000
MT per year or about 60 MT per day.
Dross Smelting. Secondary aluminum smelting from low grade
drosses containing 10 to 30% aluminum metal values requires pre-processing
of the dross to enrich the aluminum metal content to about 75%. Either
wet or dry processes may be employed. In the dry process, the dross
material is crushed and comminuted in impact ion or ball mills. The
fine-, are removed, leaving enriched metal granules for smelting in
reverberator/ or rotary furnaces. Alternatively, the dross raw material
may be processed by water washing in rotating drum, where the water-
soluble fluxing salts are removed. Wet processing is more commonly
applied in relatively low volume dross processing plants.
Tlu; typical aluminum dross smelting plant is assumed to produce
!(),()()() fir .1 lumi nuni mo till uimu.illy from raw dross inator ial . This is
derived 1 rom typical product ions of plant-, engaged in aluminum dross
smelting. In practice dross smelting is performed in relatively large
plants wnere the total metal production amounts to about 50,000 MT per
year. For waste disposal assessment, dross smelting may be considered
separately as taking place in a hypothetical plant. The daily production
volume is taken as 30 MT.
' - ••"' Description of Waste Streams
Scrap melting. In either reverberatory or rotary furnace
aluminum scrap smelting, the byproduct slag or dross waste will consist
mainly of a salt mixture of sodium chloride (Nad), potassium chloride
(KC1) and magnesium chloride (MgCl9) with about 15% aluminum value.
This slag is further processed in plants with dross smelting capabilities.
For the typical plant with 2(),l)0() MT annual production volume, about
2,00(1 MT of the dross is generated annually (100 kg/MT of aluminum
metal). 'I his byproduct, dross is currently recycled. It ma\ present a
future disposal problem if druss smelting for further metal recovery
!>ei J.IK-S non-profitable or impractical. Byproduct dross is stored in
covered areas so there is no danger of salt leaching.
Another smelting activity related waste arises from air pollution
control residues. Air emission control is provided mainly by wet-
scrubbing with lime treatment. For the typical plant with 20,000 MT
aluminum alloy production, it is estimated that about 5,000 MT sludge
thickened to 30% solid content is generated annually. In solubility
tests described in Appendix B fluoride was found to leach from emission
control dust in significant concentration. Similar leaching is expected
from emission control sludge. For this reason emission control sludge
(or dusts) are considered potentially hazardous.
288
-------
Dross Smelting. About 10-15% of total secondary aluminum
metal is produced from dross smelting. The initial raw dross is enriched
from 25% metal content to 75% metal content. In wet processing, the
flux salt remaining from dross enrichment is removed by dissolution in
water. In dry processing, the flux salt is recovered as Al-0 which is
recycled as steel melting flux cover agent. The major source of solid
waste arises from both the impure constituents in the enriched dross and
the salt flux used in the smelting process. The waste generated is in
the form of a high salt slag consisting mainly of NaCl and KC1 with
about 6-8% residual aluminum value. For the typical plant production of
10,000 NTT aluminum from drosses, about 14,000 NfT high salt slag is
generated (1400 kg/WT aluminum). Relatively small quantities of flue
dust are also generated as furnace emissions, and these are combined
with the slag waste. This waste is considered potentially hazardous
because of the high soluble salt content (NaCl, KC1) which can easily be
leached and contaminate ground water.
3.2.3 Waste Quantities. Table 120 summarizes the generation factors
for the land disposed residuals (i.e. scrubber sludge, high salt slag)
from secondary aluminum smelting operations. This table also gives
typical concentrations of potentially hazardous constituents. Table 121
gives the yearly quantity of wastes and hazardous constituents thereof
for a typical plant producing 20,000 MT of aluminum alloy metal from
scrap metal and 10,000 MT of aluminum metal from aluminum drosses.
Appendix A gives analyses of samples collected from secondary aluminum
smelters.
Using the sludge and high salt slag generation factors given
in Table 120 and concentration factors given in Table 120 along with
the capacity of each state for secondary aluminum production estimates
have been made for total and hazardous waste generation. These estimates
are given for 1974, 1977, and 1983 in Table 122.
Projections of secondary aluminum production volumes for the
years 1977 and 1983 cannot be derived by simple extrapolation of past
production data. The annual consumption of aluminum has been increasing
exponentially at the rate of about 6% per year. Primary aluminum production
capacity has kept pace more as a step function. Thus, the secondary
smelter production, in satisfying the total market demand, had to contend
with occasional primary smelting overcapacities and metal stockpiles.
As a result, secondary aluminum annual production volume growth has
followed a rather regular step function pattern of making a 20% increase
after about every three years at a steady production rate. It appears
that, even though the growth in aluminum consumption and primary smelting
will level off somewhat, secondary aluminum smelting will continue to
grow at the exponential rate of 7% per year. The projected 1977 secondary
aluminum production volume would be about 20% higher than that in 1974,
and the 1983 volume would be about 70% higher. Both the smelter process
technology and air emission control technology are not expected to
change significantly within the short term. Therefore, both the air
289
-------
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291
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Table 122-b
ESTIMATED STATE, REGIONAL, AND NATIONAL WASTE FROM
SECONDARY ALUMINUM SMELTING
TOTAL SCRUBBER SLUDGE, 1977 (METRIC TONS. DRY WEIGHTS}*
STATE
ALABAMA
ARKANSAS
CALIFORNIA
CONNECTICUT
ILLINOIS
INDIANA
KANSAS
MARYLAND
MICHIGAN
NEW JERSEY
NEW YORK
OHIO
IT NNSVI VANIA
S CAROLINA
VLXAS
WASHINGTON
WISCONSIN
EPA REG. I
U
m
isr
sr
yr
-so.
rx
x
NATIONAL TOTAL
TOTAL
DISPOSED
3,720
3,000
12,000
3,000
15,000
15,000
1,440
1,440
5,280
3.000
7,440
37.hfiO
.1,000
3.000
3,000
1,440
3.000
3,000
10,440
4,440
6,720
75,840
6,000
1,440
12,000
1,440
121,320
TOTAL
POTENTIALLY
HAZARDOUS
3,720
3,000
12,000
3,000
15.000
15,000
1,440
1,440
5.280
3.000
7.440
37.660
3,000
3,000
3.000
1,440
3.000
3.000
10,440
4,440
6,720
75,840
6,000
1,440
12,000
1,440
121,320
TOTAL
HAZARDOUS
CONSTITUENTS
29.3
23.5
94.7
23.5
118.0
118.0
11.4
11.4
43.0
23.5
68.8
29S.8
23.6
23.6
23.6
11.4
23.5
23.5
82.3
34.9
52.8
598.3
47.0
11.4
94.7
11.4
966.3
DISPOSAL
METHOD
LAGOON
HAZARDOUS
CONSTITUENTS
Cu
4.7
3.7
15.0
3.7
18.7
18.7
1.8
1.8
7.9
3.7
9.4
46.9
3.7
3.7
3.7
1.8
3.7
3.7
13.1
5.5
8.4
95.9
7.4
1.8
15.0
1.8
152.6
Pb
0.5
0.4
1.7
0.4
2.0
2.0
0.2
0.2
0.7
0.4
1.1
6.3
0.4
0.4
0.4
0.2
0.4
0.4
1.5
0.6
0.9
10.4
0.8
0.2
1.7
0.2
18.7
Zn
24.1
19.4
78.0
19.4
97.2
97.2
9.4
9.4
34.3
19.4
48.4
243.6
10.4
19.4
13.4
9.4
19.4
19.4
67.8
28.8
43.5
491.7
38.8
9.4
78.0
9.4
786.8
•MULTIPLY BY 3.0 TO CONVERT TO APPROXIMATE WET WEIGHTS.
MULTIPLY BY 1.1 TO CONVERT TO SHORT TONS.
SOURCE: CALSPAN CORPORATION
292
-------
Table 122-c
ESTIMATED STATE, REGIONAL, AND NATIONAL WASTE
FROM SECONDARY ALUMINUM SMELTING
TOTAL SCRUBBER SLUDGE, 1983 (METRIC TONS, DRY WEIGHTS)*
STATE
ALABAMA
ARKANSAS
CALIFORNIA
CONNECTICUT
ILLINOIS
INDIANA
KANSAS
MARYLAND
MICHIGAN
NEW JERSEY
NEW YORK
OHIO
PENNSYLVANIA
S. CAROLINA
TEXAS
WASHINGTON
WISCONSIN
EPA REG. I
n
m
Iff
sr
VI
OT
TJC
X
NATIONAL TOTAL
TOTAL
DISPOSED
5,270
4,250
17,000
4,250
21,250
21.250
2,040
2,040
7,480
4,250
10,540
53.210
4,250
4750
4.250
2,040
4,250
4,250
14,790
6^90
9.520
107,440
8,500
2,040
17.000
2,040
171,870
TOTAL
POTENTIALLY
HAZARDOUS
5,270
4,250
17.000
4,250
21.250
21.250
2.040
2,040
7,480
4,250
10,540
53710
4,250
4.250
4.250
2,040
4750
4,250
14.730
6,290
9,520
107,440
8.500
2,040
17,000
2,040
171.870
TOTAL
HAZARDOUS
CONSTITUENTS
41.5
33.3
134.1
33.3
167.1
167.1
16.2
16.2
60.9
33.3
83.3
419.1
33.3
33.3
33.3
16.2
33.3
33.3
116.6
49.5
74.8
847.6
66.6
167
134.1
16.2
1,354,8
DISPOSAL
METHOD
LAGOON
1
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HAZARDOUS
CONSTITUENTS
Cu
6.6
5.3
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5.3
26.5
26.5
1.6
2.6
11.2
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66.5
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7.9
11.9
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2.6
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216.8
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2.9
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110.5
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137.7
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13.3
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345.1
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275
27.5
13.3
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27.5
96.0
40.8
61.7
696.6
55.0
13.3
110.5
13.3
1,114.7
•MULTIPLY BY 3.0 TO CONVERT TO APPROXIMATE WET WEIGHTS.
MULTIPLY BY 1.1 TO CONVERT TO SHORT TONS.
SOURCE: CALSPAN CORPORATION
293
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emission control scrubber sludge figures for scrap smelting and the high
salt slag data for dross smelting for 1974 were multiplied hy a factor
of 1.2 to derive the corresponding 1977 projections and by 1.7 for 1983
projections.
The state of Ohio has the greatest quantity of potentially
hazardous sludge generated followed by California, Illinois and Indiana.
Ohio also generates the greatest quantity of potentially hazardous high
salt slag followed by Indiana, California and Pennsylvania. High salt
slag is generated in only four states whereas sludge is generated in 17
states. This is because only 4 states have dross processing facilities
which generate high salt slag.
3.3 TREATMENT AND DISPOSAL TECHNOLOGY
3.3.1 Current Waste Treatment and Disposal Practices
Currently wet scrubber sludges from secondary aluminum smelters
is usually sent to unlined lagoons. The high salt slag remaining from
dross smelting is currently open dumped either on site or off site. It
is estimated that on a national basis about 50% of the slag is disposed
of on site and 50% off site by contract disposal service. Both of these
wastes are considered potentially hazardous as discussed previously.
Levels of technology for their treatment and disposal are discussed in
the following sections.
3.3.2 Present Treatment and Disposal Technology (Level I)
Scrubber Sludge. At the present time scrubber sludge is
usually disposed of in unlined lagoons. This practice is environmentally
inadequate if fluorides leach through permeable soils to groundwater.
High salt slag. At the present time high salt slag residue
from dross processing is disposed of in open dumps.
Potassium, sodium and chloride which are present at high
concentrations in high salt slag are relatively non-toxic when compared
to heavy metals. However, the high concentration of these constituents
and their high solubility presents a potential hazard to ground water
quality. Open dumping of high salt slag in permeable soil areas is
therefore environmentally unacceptable.
3.3.3 Best Technology Currently Employed (Level II)
Scrubber Sludge. At least one secondary smelter uses lined
lagoons for scrubber sludge disposal. The use of lined lagoons to
preclude leaching of fluorides and other constituents is environmentally
acceptable and comprises Level II technology.
297
-------
High Salt Slag. Level II technology is the same as Level I
(i.e. open dumping) and is not environmentally sound.
3.3.4 Technology to Provide Adequate Health and Environmental
Protection (Level III)
Scrubber Sludge. Level III technology is the use of lined
lagoons which is environmentally adequate.
High Salt Slag. Level III technology is ground sealing of high
salt sJag disposal areas with bentonite or other soil sealants to prevent
movement of soluble salts through permeable soils.
Tables 123 and 124 summarize features of Levels I, II and III
treatment and disposal technology including extent of use, factors affecting
degree of hazard, and environmental adequacy.
3.4 COS'I ANALYSES
In the last section various treatment and disposal technologies
currently employed or considered for adequate health and environmental
protection were described. The costs of implementing this technology for
typical plants is considered in this section. Costs are given separately
for reverbatory smelter producing 20,000 MT of aluminum metal per year and
the dro:;:; sineJting facility producing 10,000 MT of aluminum metal.
3.4.1 Cost of Present Treatment and Disposal Technology (Level I)
Reverbatory Smelting. About 5,000 MT of scrubber water sludge
with a 30 percent solid content is generated annually which is disposed in
an on-site lagoon.
The estimated sludge density is 1250 kg/m . About 4,000 m of
sludge are deposited annually in the lagoon. The lagoon is designed to
retain 80,000 m of sludge which represents 20 years of operation. The
lagoon characteristics necessary to accommodate this volume of sludge are:
Volume 100,000 m
Bottom width 95 m
Top width 115 m
Bottom length 189 m
'Iup length 209 m
Total depth 5 m
Depth of excavation 1.35 m
Circumference 659 m.
Dike volume 24,772 nu
Dike surface 14,524 m
Total width 135 m
Total length 230 m
Required area 3.1 ha
298
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Eight test holes each 10 m deep, are drilled and 2 samples are
analyzed per hole during site preparation. The lagoon is situated on semi-
industrial land. Costs of Level I treatment and disposal technology are
given in Table 125.
Dross Smelting. Annual production of 10,000 MT results in 14,000 MT
of slag. Current practice is disposal by contractors at off-site landfills.
The slag density is 2000 kg/m . Contractor costs consist,of
loading, hauling and grading at the dump site and amount to $3.57/m . Costs
are summarized below for Level I disposal technology:
Capital Cost N.A. --
Annual Cost
Contractor $24,990
Total $24,990
_!/ Contractor Disposal
3.4.2 Cost of Best Technology Currently Employed (Level II)
Reverbatory Smelting. At least one secondary aluminum smelter
uses a lined lagoon for containment of scrubber water sludge. This practice
will preclude any leaching. The added capital cost of the lagoon lining for
the hypothetical plant is estimated as $106,525. Annual costs are summarized
below:
Annual Cost
Construction Amortization $12,370
Construction Repair § Maintenance 3,195
Insurance 1,065
Total $16,630
Dross Smelting. Disposal of high salt slag in landfills is both
Level I and II practice. Thus there are no added costs for Level II.
3.4.3 Cost of Technology to Provide Adequate Health and Environmental
Protection.
Reverbatory Smelting. The use of lined lagoons for scrubber water
sludge is also Level III practice and there are no added costs.
Dross Smelting . The high salt slag wastes are disposed by an
outside contractor. It is assumed that the contractor must secure his land-
303
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TABLE 125 COST OF LEVEL I TREATMENT AND DISPOSAL TECHNOLOGY
SECONDARY ALUMINUM PLANT - REVERBATORY SMELTING
CAPITAL COST
Sludge
Lagoon
Site Preparation
Survey $ 1,940
Test Drilling 1,960
Sample Testing 1,000
Report Preparation 1,800
Construction
Excavation $ Forming 32,945
Compacting 45,830
Fine Grading 6,535
Soil Poisoning 815
Transverse Drain Fields 3,105
Land 12,260
TotaJ $108,190
ANNUAL COST
Land $ 1,225
Amortization
Construction 11,140
Equipment
Operating Personnel
Repair and Maintenance
Construction 2,675
Equipment
Energy
Fuel
Electricity 30
Taxes 305
Insurance 1,080
Total $16,455
304
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fill in the same manner as required for on-site dumping; namely seal the
ground,, construct collection ditches and install a pump and piping. The
estimated cost to the plant is $10,670 per year.
Annual Cost
Construction Amortization $ 3,460
Equipment Amortization 1,740
Construction Repair § Maintenance 895
Equipment Repair $ Maintenance 545
Energy
Electricity 80
Insurance 405
Adm. S Profit 3,565
Total $10,670
Costs of Levels I, II and III treatment and disposal technologies
are summarized in Table 126 for reverbatory smelting and Table 127 for
dross smelting. Costs are given in dollars per metric ton of dry and
wet waste and dollars per metric ton of aluminum product. Costs for
each type of waste (i.e. sludge, high salt slag) at each technology
level are also expressed as percentages of metal selling prices.
Total industry costs for Levels I, II and III treatment and
disposal technologies have been estimated in 1973 dollars and are also
presented in Tables 126 and 127 for reverbatory smelting and dross
smelting respectively. The industry annualized cost for Level I sludge
disposal from reverbatory smelting is estimated as $620,000 which represents
0.13% of 1973 estimated sales. The industry costs for Levels II and III
technology are estimated as $1,250,000 and consists of lining sludge
disposal lagoons for adequate environmental protection.
Total annualized industry costs for Level I and II treatment
and disposal of high salt slag from dross smelting is estimated as
$470,000 which represents 0.39% of estimated 1973 national sales. The
annualized industry cost for Level III treatment and disposal technology
(i.e. adequate for environmental protection) is $670,000 or 0.56% of
estimated 1973 sales.
305
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307
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SECTION IV
LIST OF REFERENCES
1. Minerals Yearbook 1972, Vol. I, Metals, Minerals and Fuels, Prepared by
Staff of the Bureau of Mines, U.S. Govt. Printing Office, Washington,
D.C., 1974.
2. Minerals Yearbook, 1975, Vol. I, Metals, Minerals and Fuels, Prepared by
Staff of the Bureau of Mines, U.S. Govt. Printing Office, Washington,
D.C., 1975.
3. "The Outlook for Mercury in the United States", V. Anthony Cammarvta, Jr.,
Proc. of the First Intl. Mercury Congress, Barcelona, Spain, V. 1, May
6-10, 1974.
4. Beneficiation of Titanium Chlorination Residues, C.C. Merrill, M.M. Wong,
and D.D. Blue, U.S. Bureau of Mines Report of Investigations 7221, 1969.
5. "The Electrolytic Tin Refining Plant at Texas City, Texas", Thomas S.
Mackey, J. of Metals. June 1969.
6. Trace Elements in Biochemistry, H.J.M. Bowen, Academic Press, 1966.
7. Mineral Facts and Problems, Bureau of Mines Bulletin 650 by Staff,
Bureau of Mines, U.S. Govt. Printing Office, Washington, D.C., 1970.
8. Water Pollution Control in the Primary Nonferrous Metals Industry -
Vol. I., Copper, Zinc and Lead Industries, liPA-R2-73-247a, September,
1973.
9. Engineering and Mining Journal, March 1975.
10. Cadmium Metal Industry Supplement B-l, Supplement to Development
Document For Effluent Limitations Guidelines and Standards of
Performance For the Miscellaneous Non-Ferrous Smelting and Refining
Point Source Category. Calspan Report No. ND-5782-M-11 for the
USEPA, Effluent Guidelines Division, March, 1976.
SW-145C.2
309
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