ENVIRONMENTAL ASSESSMENT OF THE
DOMESTIC PRIMARY COPPER, LEAD
AND ZINC INDUSTRIES
VOLUME I
PEDCo ENVIRONMENTAL
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PEDCo-ENVIRONMENTAL
SUITE 13 • ATKINSON SQUARE
CINCINNATI. OHIO 45246
513 / 77 1-433O
ENVIRONMENTAL ASSESSMENT OF THE
DOMESTIC PRIMARY COPPER, LEAD
AND ZINC INDUSTRIES
VOLUME I
Prepared by
PEDCo-Environmental Specialists, Inc.
Suite 13, Atkinson Square
Cincinnati, Ohio 45246
Contract No. 68-02-1321
Task No. 38
EPA Project Officer
Margaret J. Stasikowski
Prepared for
U.S. ENVIRONMENTAL PROTECTION AGENCY
Industrial Environmental Research Laboratory
5555 Ridge Avenue
Cincinnati, Ohio 45268
November, 1976
BRANCH OFFICES
Suite 110, Crown Center Suite 107-B Professional Village
Kansas City, Mo. 64108 Chapel Hill, N.C. 27514
-------
DISCLAIMER
This report has been reviewed by the Industrial Environmental Research
Laboratory, U.S. Environmental Protection Agency, and approved for
publication. Approval does not signify that the contents necessarily
reflect the views and policies of the U.S. Environmental Protection
Agency, nor does mention of trade names or commercial products con-
stitute endorsement or recommendation for use.
n
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ABSTRACT
This report presents the results of a multi-media (air, water, and
.solid waste) study of the environmental impacts from the U.S. primary
copper, lead and zinc industries. The open literature was surveyed to
identify and describe all processes employed by these industries and to
characterize pollutant effluents and environmental effects from those
processes. Various pollution control systems are described and eval-
uated for domestic application, and alternate production processes are
reviewed.
Principal environmental impacts from the copper industry are air
emissions of S02 and trace metals from smelting operations. Insuffi-
cient markets for by-product sulfur compounds are a disincentive for SO?
control. Water pollution problems include acid drainage and trace metal
contamination of wastewaters from ore beneficiation processes. The fact
that many mines and smelters are located in remote arid regions tends to
mitigate the severity of these pollution problems. Newer hydrometallur-
gical processes may produce leach residues that will require further
efforts to control.
Lead is produced by six pyrometallurgical smelters in this country.
There are large lead deposits in the State of Missouri so that a large
percentage of future lead supplies are expected to originate in that
state. Mining and concentrating operations produce metal-laden waste-
waters that are effectively controlled in the Missouri area by biotic
degradation and natural precipitation at high pH. Principal smelter
emissions include S02 and particulates from sintering and blast furnace
operations. Electrostatic precipitators, baghouses, and sulfuric acid
plants are used for pollution control.
Zinc smelters are mainly located in populated areas and have large-
ly been forced to clean up their emissions in recent years. Economic
prospects have forced several smelters to close in recent years. S02 is
controlled at all smelters by means of acid plants. Particulates are
controlled by electrostatic precipitators and baghouses. There is some
concern that volatile metals may pass through these control devices in
the vapor phase.
Several programs for further research and development are indi-
cated.
m
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CONTENTS
Page
ABSTRACT
FIGURES
TABLES
ACKNOWLEDGMENT
1.0 INTRODUCTION
2.0 COPPER INDUSTRY
Industry Description
Industry Analysis, Copper
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
Process
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
1,
2,
3,
4,
5,
6,
7,
8,
9,
10,
11,
12,
13,
14,
15,
16,
17,
18,
19,
20,
21,
22,
23,
24,
25,
26,
27,
28,
29,
30,
31,
32,
33,
34,
35,
36,
37,
38,
Mining
Concentrating
Multiple-Hearth Roasting
Fluidization Roasting
Dryi ng
Reverberatory Smelting
Electric Smelting
Flash Smelting
9, Peirce-Smith Converting
Hoboken Converting
Noranda
Slag Treatment
Contact Sulfuric Acid Plant
DMA SOo Absorption
Elemental Sulfur Production
Arsenic Recovery
Fire Refining and Anode Casting
Electrolytic Refining
Electrolyte Purification
Melting and Casting Cathode Copper
Slime Acid Leach
CuSO, Precipitation
Slimes Roasting
Slime Water Leach
Dore Furnace
Scrubber
Soda Slag Leach
Selenium and Tellurium Recovery
Dore Metal Separation
Heap and Vat Leaching
Cementation
Solvent Extraction
Electrowinning
Sulfation Roasting
Sponge Iron Plant
CLEAR Reduction
CLEAR Regeneration - Purge
CLEAR Oxidation
Process No. 39, Cymet Leaching
1
12
12
27
30
35
42
50
53
55
64
67
70
77
79
81
83
90
93
95
98
106
110
114
117
119
121
123
125
127
129
131
133
135
138
140
142
144
146
148
150
152
154
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CONTENTS (Continued)
3.0
Process No. 40,
Process No. 41,
Process No. 42,
Process No. 43,
Process No. 44,
Process No. 45,
LEAD INDUSTRY
Industry Description
Industry Analysis
Cymet Crystallization
Cymet Reduction
Cymet Solvent Regeneration
Arbiter Leaching
Arbiter Precipitation
Arbiter Decomposition
4.0
Process No. 1
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
ZINC INDUSTRY
Industry Description
Industry Segment Analysis
Mining
Concentrating
Sintering
Acid Plant
Blast Furnace
Slag Fuming Furnace
Dressing
Dross Reverberatory Furnace
9, Cadmium Recovery
10, Reverberatory Softening
Kettle Softening
Harris Softening
Antimony Recovery
Parkes Desilverizing
Retorting
Cupelling
Vacuum Dezincing
Chlorine Dezincing
Harris Dezincing
Debismuthizing
Bismuth Refining
Final Refining and Casting
2,
3,
4,
5,
6,
7,
8,
11
12
13
14
15
16
17
18
19
20
21
22
Process No
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
1, Mining
2, Concentrating
3, Multiple-Hearth Roasting
4, Suspension Roasting
5, Fluidized-Bed Roasting
6, Sintering
7, Horizontal Retorting
8, Vertical Retorting
9, Electric Retorting
10, Oxidizing Furnace
11, Leaching
12, Purifying
13, Electrolysis
14, Melting and Casting
Page
156
157
158
159
161
163
165
165
169
172
176
184
193
195
203
208
212
214
215
218
220
222
224
226
228
230
232
233
235
237
239
241
241
253
256
259
268
272
275
279
285
293
299
303
307
311
313
318
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CONTENTS (Continued)
Page
Process No. 15, Cadmium Leaching 320
Process No. 16, Cadmium Precipitation 323
Process No. 17, Cadmium Purification and Casting 325
5.0 AIR MANAGEMENT 328
Emission Characteristics 328
Emission Control Systems 330
Fugitive Emissions 333
6.0 WATER MANAGEMENT 335
Sources of Secondary Water Pollution 335
Concentrator Effluent 337
Sulfide Weathering 338
Waste Characteristics 340
Waste Handling and Treatment 341
Advanced Treatment 352
7.0 EMERGING TECHNOLOGY 359
8.0 RECOMMENDED RESEARCH AND DEVELOPMENT PROGRAMS 382
APPENDICES
A. HEALTH EFFECTS OF PRIMARY COPPER, LEAD AND ZINC SMELTING 398
1. SELECT REVIEW OF PUBLISHED TOXICOLOGY AND EPIDEMIOLOGY 399
2. AN EPIDEMIOLOGICAL ANALYSIS OF DISEASE SPECIFIC MORTALITY 447
ASSOCIATED WITH PRIMARY Cu, Pb and Zn SMELTING
3. CONCLUSIONS AND AN ASSESSMENT OF THE PUBLIC HEALTH 480
HAZARD OF PRIMARY Cu, Pb and Zn SMELTING
B. ECOLOGICAL EFFECTS 482
VI1
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LIST OF FIGURES
No. Page
1-1 Yearly Copper, Lead, and Zinc Commodity Prices, 1962-1976 5
1-2 Primary U.S. Nonferrous Smelting and Refining Locations 9
2-1 Copper Industry Flow Sheet 28
3-1 Lead Industry Flow Sheet 170
4-1 Zinc Industry Flow Sheet 254
viii
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LIST OF TABLES
No. Page
1 Copper Industry Summary of Process Waste Streams 4
2 Lead Industry Summary of Process Waste Streams 11
3 Zinc Industry Summary of Process Waste Streams 16
1-1 By-Product Elements of The Industries - 1975 3
1-2 By-Products - Primary Non-Ferrous Smelter and 6
Refineries
1-3 Mine Production of Copper, Lead, and Zinc by States 10
(1975)
2-1 Copper Minerals Important in U.S. Production 13
2-2 Typical Analysis of Copper Ore Used at White Pine 15
Copper Company, Michigan
2-3 Consumption of Refined Copper in 1975 16
2-4 Salient Statistics of the Primary Copper Industry in 18
the United States in 1974
2-5 Largest By-Product Sulfuric Acid Producers - 1974 19
2-6 U.S. Primary Copper Producers (Conventional Smelting/ 20
Refining Operations)
2-7 Twenty-Five Leading Copper Mines in the United States 21
in 1974
2-8 1976 Survey of Mine and Plant Expansions in the 23
United States
2-9 Annual Generation of Hazardous Pollutants from Primary 24
Copper Industry - 1968 (metric tons)
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LIST OF TABLES (continued)
No. Page
2-10 Raw Waste Load in Water Pumped from Selected Copper 31
Mines
2-11 Analysis of Copper Concentrate 37
2-12 Typical Flotation Collectors 38
2-13 Typical Size Profile of Multiple-Hearth Copper Roaster 44
Effluents
2-14 Concentration and Weight Analysis of Particulate 45
Effluents from a Multiple-Hearth Copper Roaster
2-15 Typical Levels of Volatile Metals in Domestic Copper 46
Ore Concentrations
2-16 Composition of Reverberatory Furnace Exhaust Gases 58
2-17 Effluents from Slag Granulation (mg/1) 59
2-18 General Range of Reverberatory Furnace Slag Com- 60
position
2-19 Material Balance on Converters - Smelters in Arizona 71
(percent)
2-20 Composition of Converter Dust 72
2-21 Particle Size Distribution in Converter Dust 73
2-22 Particulate Emissions Analysis at Stack Outlet for 73
Reverberatory Furnace and Converter
2-23 Converter Off-Gas Composition 74
2-24 Estimated Maximum Impurity Limits for Metallurgical 85
Off-Gases Used to Manufacture Sulfur Acid (Approxi-
mate limit, (mg/Nm?)
2-25 Raw Waste Characterization: Acid Plant Slowdown 86
2-26 Acid Plant Slowdown Control and Treatment Practices 88
2-27 Analysis of Arsenic Plant Washdown Water 97
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LIST OF TABLES (continued)
No. Page
2-28 General Range Analysis of Anode Copper 99
2-29 Water Requirements for Copper Refineries 101
2-30 Waste Effluents from Anode Cooling Water 102
2-31 Contact Cooling Water Control and Treatment Practices 103
2-32 General Range Analysis of Electrolyte, Refined Copper 108
and Anode Slime
2-33 Waste Effluents from NiSO, Barometric Condenser 112
2-34 Analysis of Water Used to Cool Refinery Shapes 116
(Concentrations in mg/1)
2-35 Dore Metal Analysis 125
3-1 Lead Minerals, By Name, and Composition 166
3-2 Domestic Primary Lead Producers 168
3-3 Analysis of a Missouri Mine Water 174
3-4 . Typical Southeastern Missouri Lead Concentrate 178
Analyses (Percent by weight)
3-5 Western Lead Concentrate Analyses 179
3-6 Flotation Chemicals 180
3-7 Lead Mill Wastewater Analysis 181
3-8 Sinter Analysis 185
3-9 Sinter Machine Feed 185
3-10 Grain Loading and Weight Analysis of Input Feed and 188
Effluents-Updraft Lead Sintering Machine
3-11 Typical Size Profile of Effluents, Updraft Lead 188
Sintering Machine
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LIST OF TABLES (continued)
No. Page
3-12 Analysis of Sinter Machine Exhaust Gases (Missouri 189
Lead Operating Company)
3-13 Atmospheric Control Systems on Primary Lead Sintering 190
Machines
3-14 Wastewater Treatment at Primary Lead Acid Plants 194
3-15 Scrubber Wastewater Treatment at Primary Lead Plants 194
3-16 Lead Bullion Composition 196
3-17 Typical Blast Furnace Slag Analysis 197
3-18 Typical Blast Furnace Charge 198
3-19 Atmospheric Control Systems on Primary Local Blast 200
Furnaces
3-20 Waste Effluents from Slag Granulation 204
3-21 Primary Lead Slag Granulation Wastewater Treatment 206
3-22 Lead Bullion Analysis Basic: As drossed 209
3-23 Typical Compositions of Softened Lead Bullion and 216
Slag (Amounts in weight percent)
3-24 Typical Retort Analysis 226
4-1 Mining, Production, and Consumption of Zinc and 242
Cadmium (metric tons)
4-2 Twenty-Five Leading Zinc Mines in the United States 243
4-3 Common Ores Mines for Their Zinc Content 246
4-4 U.S. Slab Zinc Consumption - (1975) 248
4-5 Grades of Commercial Zinc 248
4-6 Primary Zinc Processing Plants in the United States 250
4-7 Range of Compositions of Zinc Concentrates 261
Xll
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LIST OF TABLES (continued)
No. Page
4-8 Typical Flotation Reagents Used for Zinc Concentra- 262
tions
4-9 Ranges of Constituents of Wastewaters and Raw Waste 264
Loads for Five Selected Mills
4-10 Product Sinter Composition (percent) 280
5-1 Elemental Analysis of Particulate Size Fractions 329
1-1 A Classification of the Effects of Metals 400
1-2 Target Organs of Metals 402
q
1-3 Acceptable Average Concentrations (i^g/m ) of 403
Occumpational Expsoure Based on 8-Hours Exposures
1-4 Tolerance Levels for Metals in Drinking Water and 404
Results of Sampling of Community Water Supplies
(969) in 1969
1-5 Metal Carcinogenesis in Experimental Animals 405
1-6 Effects of Metals on Reproduction 406
1-7 Body Burden and Human Daily Intake and Content in 407
the Earth's Crust of Selected Elements
1-8 Some Toxic Levels of Cadmium Exposure 411
1-9 Effects of Inorganic Lead Salts in Relation to 414
Absorption
1-10 Distribution of Sequelae Following Various Modes of 416
Onset in 425 Patients with Plumbism
2-1 Causes of Death used to Generate Mortality Profiles 449
for Counties Containing Cu, Pb, or Zn Smelters
2-2 Cancer Mortality Rates from 1950-1969 for Counties 450
in the ARea of the Bunker Hill Lead Smelter
2-3A Liver and Biliary Passages Cancer Mortality Rankings 452
for Counties in the Area of the Bunker Hill Lead
Smelter
xm
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LIST OF TABLES (continued)
No. Page
2-3B Trachea, Lund and Bronchus Cancer Mortality Rankings 453
for Counties in the Area of the Bunker Hill Lead
Smelter
2-3C Kidney Cancer Mortaility Rankings for Counties in the 454
Area of the Bunker Hill Lead Smelter
2-3D Bladder Cancer Mortality Rankings for Counties in the 455
Area of the Bunker Hill Lead Smelter
2-3E Thyroid Cancer Mortaility Rankings for Counties in the 456
Area of the Bunker Hill Lead Smelter
2-3F Mortaility Rankings for all Cancers Combined for 457
Counties in the Area of the Bunker Hill Lead Smelter
2-4 Lead Smelter Cancer Mortality Data Summary 458
2-5 Zinc Smelter Cancer Mortality Data Summary 459
2-6 Copper Smelter Cancer Mortality Data Summary 460
2-7 Lead Smelter Non-Cancer Mortality Data Summary 461
2-8 Zinc Smelter Non-Cancer Mortality Data Summary 462
2-9 Copper Smelter Non-Cancer Mortality Data Summary 463
2-10 Observed Mortality Associations for Counties 466
Containing Primary Lead Smelters
2-11 Observed Mortality Association for Counties 467
Containing Primary Zinc Smelters
2-12 Observed Mortality Associations for Counties 468
Containing Primary Copper Smelters
2-13 Copper, Lead or Zinc Smelter-Containing Counties with 472
Significantly Elevated Cancer Mortality Rates
2-14 Definition of Ranges of Element Concentrations in Cu 474
and Zn Smelter Ore Concentrates
2-15 Quantities of Nine Elements in Zn Smelters Ore 475
Concentrates
2-16 Quantities of Nine Elements in Cu Smelter Ore 476
Concentrates
xiv
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ACKNOWLEDGMENT
This report was prepared by PEDCo-Environmental Specialists, Inc.,
under the direction of Mr. Timothy W. Devitt. Principal authors were
Dr. Gerald A. Isaacs, Mr. Thomas K. Corwin, Mr. Hal M. Drake, Mr.
Douglas J. Morel!, and Mr. Jeffrey A. Smith.
Project officer for the U.S. Environmental Protection Agency was
Ms. Margaret J. Stasikowski.
The authors appreciate the efforts and cooperation of everyone who
participated in the preparation of this report.
xv
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XVI
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1.0 INTRODUCTION
This report presents the results of an environmental assessment of
the primary copper, lead, and zinc industries in the United States. The
report is multi-media in scope, covering the fields of air, water, and
solid waste. An open literature survey identified all of the processes
employed in the primary production of these three metals having a signi-
ficant environmental impact. The report covers all phases of production,
from the mining and concentrating of ores to the smelting, refining, and
casting of copper, lead, zinc, and related products. Engineering studies
of the available data were undertaken to describe all source of pollution
from existing and potential production processes used in the manufacture
of these metals. Potentially hazardous process outputs were singled out
for special attention, and all toxic materials released to the environ-
ment were identified. Two appendices are included. The first assesses
the toxic potential of the residual metals emitted from the smelters and
refineries, and the second details the ecological impacts of the primary
production processes. As a result of certain limitations on the amount
of information available, some important environmental questions have
been left unanswered. However, the report constitutes a comprehensive
multi-media environmental overview which identifies these information
gaps as well as research and development efforts needed for more effec-
tive pollution control.
As defined for this report, the domestic primary copper, lead, and
zinc industries consist of the facilities in this country that extend
from the mining of the ores of these three metals through the production
of the purified metals as marketable castings. Included are concen-
trating plants that partially separate the ore minerals, smelters that
produce the metals, and refineries that purify them to within accepted
quality specifications. These industries do not include operations that
fabricate the metals into commercial products, or that blend them to
manufacture alloys.
Although each of the three industries makes a distinct product,
there is such a strong interrelation between them that for many years
they have been considered a single economic unit. Their products are
marketed through many of the same channels, and with the same procedures.
Since many of the ores contain recoverable quantities of more than one
of the metals, there is regular exchange of material, and several
companies produce two, or all three, of the metals. They share similar
production techniques, and produce similar waste materials.
-------
The principal products of the industries are the three metals,
copper, lead, and zinc. Copper has been one of the most useful metals
since the beginning of recorded history. Its applications now include
direct fabrication into articles that range from plumbing to cooking
pots, and copper is absolutely necessary for the continuation of the
electrical and electronic industries as we know them. Lead is used in
solders and type metal, and it is the principal material used for stor-
age batteries and for radiation shielding. Zinc is most used as a base
for die casting alloys, as galvanizing for protection of steel, and as
the alloy with copper known as brass.
The value of the products of these industries in 1973 was $2.9
billion, or about 0.2 percent of the gross national product. This
figure includes the value of the many by-products of the industry.
Eighteen other chemical elements are isolated from copper, lead, and
zinc ores, although in many cases the final processing is completed at
plants other than the smelters. Some of these elements, such as cadmium,
selenium, and tellurium, have no other industrial source. A major
source of domestic gold and silver is as a by-product of this industry.
Table 1-1 lists these by-product elements and pertinent data.
In terms of domestic consumption, copper is the largest of the
three primary metals. In 1975, 1.4 million metric tons of copper were
used, compared to 0.8 million metric tons of zinc and 0.6 million metric
tons of lead. Although consumption for the last two years has been
substantially below previous years, the demand for copper is expected to
increase. Forecasts for the year 2000 range from 3 to 5 times the
present consumption, as high as 8 to 13 million tons per year. Markets
for lead and zinc, however, have both suffered declines that are expected
to continue, at least for the short term. Other materials have entered
strongly into their former markets for paints, plumbing, and small
castings. A major use of lead is in additives for motor fuel, whose use
is being phased out. Plastic coatings are replacing some of the corro-
sion-prevention applications that these metals once filled.
These industries are a major source of employment in some sections
of the country. Total direct employment is approximately 50,000 workers.
These industries are the largest single employer in four states, and in
several communities they are virtually the only ones. A metropolitan
area, Salt Lake County, Utah, estimates that 14.6 percent of the work
force is directly or indirectly accountable to these industries.
Thirteen different companies operate smelters in the United States.
In addition, many other companies, large and small, are engaged primarily
in mining and concentrating operations. No single company dominates any
one of the industries. The products are sold as commodities, meeting
international specifications for quality, at prices that must meet
international supply and demand. There is very strong international
competition.
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Table 1-1. BY-PRODUCT ELEMENTS OF THE INDUSTRIES - 1975
By-product
Source
Approximate quantity
Remarks
Selenium
Tellurium
Nickel
Arsenic
Molybdenum
Rhenium
Silver
Gold
Platinum
Palladium
Antimony
Cadmium
Indium
Gallium
Thallium
Mercury
Germanium
Bismuth
Copper electrolytic slimes
Copper slimes and lead concentrates
Copper smelter
Copper smelter flue dusts
Copper flotation mills
Copper flotation mills
Copper slimes and lead concentrates
Copper slimes and lead concentrates
Copper slimes and lead concentrates
Copper slimes and lead concentrates
Lead smelters
Zinc smelters
Zinc smelter residues
Zinc smelter residues
Zinc smelter residues
Zinc smelter residues
Zinc smelter residues
Copper and lead ores
325 metric tons
120 metric tons
815 metric tons
Not disclosed
18,000 metric tons
4,500 metric tons
1,350 metric tons
10 metric tons
Not disclosed
Not disclosed
8,500 metric tons
2,270 metric tons
Not disclosed
Not disclosed
Not disclosed
Very small
Very small
Not disclosed
Only industrial source
Partial disclosure
From imported concentrates
From imported concentrates
From molybdenum concentrates
Important economic by-product
Important economic by-product
Important economic by-product
Important economic by-product
Common constituent of lead ores
Only industrial source
From cadmium concentrates
From cadmium concentrates
From cadmium concentrates
Only occasional ores recoverable
No data has been disclosed
Limited market
-------
Commodity price trends for the past 10 years are shown in Figure
1-1. They show an increase in price typical of other materials, but
they also show the rather unstable price picture that has affected this
industry in the last few years. This graph also shows that copper is
considerably more costly than the other two metals. Copper is not a
common element in the earth's crust, as indicated by the grade of the
ore being mined. Most lead ores are 3 to 8 percent lead, sometimes much
higher. Zinc ores average 8.8 percent zinc. Domestic copper ores,
however, rarely reach 4 percent copper. They average less than 1
percent, and ores as low as 0.4 percent are being processed.
The United States is almost self-sufficient in its smelter produc-
tion of copper and slightly less so for lead; it is not self-sufficient
for zinc. Comparison of 1975 consumption figures with production of
primary and secondary metals show that we smelted and refined 100 percent
of the copper we used and 78 percent of the primary lead, but only about
half of the zinc. The U.S. Bureau of Mines (USBM) has classified zinc
as a shortage commodity.
In addition to copper, lead, and zinc, a variety of other elements
and chemicals are produced at primary non-ferrous smelters and refiners.
Table 1-2 lists these by-products. One of the most important by-pro-
ducts of these industries is sulfuric acid. For one company that is
favorably located near major markets for this product, sulfuric acid is
a co-product, if not the primary one. Most other smelters, however, are
located far from potential users; transportation costs make it less
expensive for the user to buy sulfur for on-site manufacture of the
acid. Location of markets is a major factor in the international
competition in these industries. Most other countries have nothing
comparable to the extensive deposits of sulfur that are found along the
U.S. Gulf Coast. In other countries, smelter sulfur has no major
competition, and their acid finds a profitable market.
The map in Figure 1-2 shows the location of the 28 copper, lead,
and zinc smelters in the United States, and of the seven copper re-
fineries. Copper is refined in a separate plant, whereas each lead and
zinc smelter includes its own facilities for purifying the product.
This map shows that plants are located in fifteen states. Most of the
copper smelters are located in or near copper mining districts, and five
states account for 85 percent of the blister copper. Four states pro-
duce all the lead, using concentrates partly from local mines. Only one
zinc smelter is now located in an active mining district.
Mine production of the ores is shown in Table 1-3. Arizona ores
produce more than half the copper, and Missouri ores produce about 80
percent of the lead. Zinc is frequently found in both copper and lead
ores, and much zinc is mined as a by-product or co-product with other
metals.
-------
80
70
60
50
5 40
o
•—*
CO
30
20
10
'62 '63 '64 .'65 '66 '67 '68 '69 '70 '71 '72 '73 '74 '75 '76
YEAR
Figure 1-1. Yearly copper, lead, and zinc commodity prices, 1962-1976.
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Table 1-2. BY-PRODUCTS - PRIMARY NON-FERROUS SMELTERS AND REFINERIES
Company
Location
Primary
product
By-products
AMAX, Inc.
The Anaconda Co.
ASARCO, Inc.
The Bunker Hill Co.
Sauget, Illinois
Boss, Missouri
Anaconda, Montana
Great Falls, Montana
Tacoma, Washington
Hayden, Arizona
El Paso, Texas
Amarillo, Texas
East Helena, Montana
Omaha, Nebraska
Glover, Missouri
Corpus Christi, Texas
Columbus, Ohio
Kellogg, Idaho
Zinc
Lead
Copper
Copper
Copper
Copper
Copper
Copper
Lead
Lead
Lead
Zinc
Zinc oxide
Lead, zinc
Sulfuric acid
Sulfuric acid
Sulfuric acid, beryllium
Copper sulfate
Arsenic, arsenic trioxide, sulfur dioxide,
sulfuric acid
Sulfuric acid
Sulfuric acid
Zinc fume
Bismuth
Sulfuric acid, zinc sulfate, cadmium
Sulfuric acid, zinc oxide, cadmium, copper
matte, copper residue, refined gold dore",
refined silver, silver concentrates
-------
Table 1-2 (continued). BY-PRODUCTS - PRIMARY NON-FERROUS SMELTERS AND REFINERIES
Company
Location
Primary
product
By-products
Cities Service
Co., Inc.
Inspiration
Consolidated
Copper Co.
Kennecott Copper
Corp.
Magma Copper Co.
National Zinc Co.
Phelps-Dodge Corp.
Copperhill, Tennessee
Miami, Arizona
Garfield, Utah
Hurley, New Mexico
Hayden, Arizona
McGill, Nevada
San Manuel, Arizona
Bartlesville, Oklahoma
Morenci, Arizona
Douglas, Arizona
Ajo, Arizona
Copper
Copper
Copper
Copper
Copper
Copper
Copper
Zinc
Copper
Copper
Copper
Copper carbonate, copper sulfate, iron sulfate
(ferric), sodium hydrosulfite, sulfuric acid,
sulfur trioxide, surface-active agents,
p-toluene sulfonic acid, emulsifiable liquid
copper fungicide, chelated plant and animal
nutrients
Sulfuric acid
Ammonium perrhenate, molybdenum disulfide,
molybdenum trioxide, nickel sulfate, sulfuric
acid
Sulfuric acid
Sulfuric acid
Sulfuric acid
Ammonia, carbon dioxide, sulfuric acid, zinc
oxide, cadmium, metal powders, spiegeleisen,
zinc alloys
Sulfuric acid
Sulfuric acid
-------
Table 1-2 (continued). BY-PRODUCTS - PRIMARY NON-FERROUS SMELTERS AND REFINERIES
Company
Location
Primary
product
By-products
St. Joe Minerals
Corp.
White Pine Copper
Co.
Hidalgo, New Mexico
El Paso, Texas
Herculaneum, Missouri
Monaca, Pennsylvania
White Pine, Michigan
Copper
Copper
Lead
Zinc
Copper
Sulfuric acid, elemental sulfur
Copper sulfate
Sulfuric acid
Sulfuric acid, zinc oxide, cadmium
00
-------
S V!
• COPPER SMELTER
D COPPER REFINERY
0 ZINC SMELTER
A LEAD SMELTER
Figure 1-2. Primary U.S. nonferrous smelting and refining locations,
-------
Table 1-3. MINE PRODUCTION OF COPPER, LEAD, AND ZINC BY STATES (1975)
Arizona
California
Colorado
Idaho
Illinois
Maine
Michigan
Missouri
Montana
Nevada
New Jersey
New Mexico
New York
Pennsylvania
Tennessee
Utah
Virginia
Washington
West Virginia
Wisconsin
Other states
Totals
Metric tons as 100 % of the metals
. Copper
724,143
134
3,247
2,932
1,836
66,489
12,679
92,271
71,610
132,970
9,187
162,100
430
1,280,028
Lead
457
27
24,780
45,763
963
332
466,971
169
2,727
1,772
2,746
11,545
2,371
1,692
737
563,052
Zinc
7,973
193
44,308
36,824
4,599
7,544
68,439
103
4,990
28,858
10,014
69,409
19,568
75,601
18,276
14,057
9,407
7,^49
563,052
Source: U.S. Bureau of Mines.
10
-------
Many of the smelters and refineries in the United States, espe-
cially in the copper industry, are old plants, some having begun opera-
tions before the turn of the century. Although most of these older
plants have since been modified, they are generally, by the standards of
some other industries, inefficient and obsolete. The owners of these
plants find it very difficult to justify the cost of rebuilding or
modernizing them. Some plants are directly associated with large mines,
and the owners find that as the mines become deeper and the best ores
become exhausted, production costs increase. Other plants are "custom"
operations that are operated as service industries, and process concen-
trates for a fee and at a very low profit. A number of plants have been
closed in recent years, two of them since 1974. There are, however, new
smelters now operating and new operations are being considered in
Wisconsin and Washington state. A lively search for new, more efficient,
processes is under way.
11
-------
2.0 COPPER INDUSTRY
INDUSTRY DESCRIPTION
Many recent technical articles emphasize the changes taking place
in the primary copper industry. There is much speculation as to the
direction these changee will or should take. Whether the trend is
toward improved pyrometallurgical processing or toward adoption of
hydrometallurgy, most experts agree that some basic changes are immi-
nent.
At five installations, newer technology is now operating. One new
smelter now uses a continuous flash smelting process. One installation
is producing copper with a roast-leach-electrowinning technique. Three
advanced hydrometallurgical processes are approaching semicommercial
production. The following description of the industry segment does not
concentrate on these installations, since they do not now account for a
sizable percentage of the copper being produced.
Most copper production is now being accomplished with the "con-
ventional" pyrometallurgical methods that center on the energy-ineffi-
cient reverberate ry furnace. Matte from the reverberatory furnace is
converted to blister copper, and the blister copper is reduced, cast into
anodes, and refined in electrolytic cells. These operations occur in
twenty locations, all but three of which were operating before World War
II. In twelve of these locations, copper has been produced since before
World War I. Although new equipment was provided during the intervening
years, in most of the plants, new technology was not. Most domestic
copper is being made now by the same procedures used 50 years ago.
Raw Materials
The principal raw materials for copper production are the domestic
ores, which consist of copper minerals embedded in a gangue rock.
Throughout the world, the copper in the minerals is most often chemi-
cally combined with sulfur, frequently also with iron or arsenic, and
sometimes with other elements. Table 2-1 shows five of these sulfide
minerals; the first three listed are most abundant in the ores of this
country.
When sulfide minerals are exposed to air and water, they lose their
sulfur content. Natural weathering may create deposits of oxidized
copper minerals. The table shows four of these, the highly colored
12
-------
Table 2-1. COPPER MINERALS IMPORTANT IN U.S. PRODUCTION
1
Mineral
Sulfide Ores
Chalcopyrite
Chalocite
Bornite
Covellite
Enargite
Oxide Ores
Malachite
Azurite
Cuprite
Chrysocolla
Native copper
Composition
CuFeS2
Cu2S
Cu5FeS4
CuS
Cu3As5S4
CuC03-Cu(OH)2
2 CuC03-Cu(OH)2
Cu20
CuSO.-2H90
*3 L.
Cu
Copper content,
percent
35
80
63
66
48
57
55
89
36
100
Occurrence3
SW, NW, NC
SW, NW, NC
SW, NW
SW, NW
NW
SW, NW
SW, NW
SW
SW
NC, SW
NW - Montana and surrounding area.
NC - Michigan and surrounding area.
SW - Arizona and surrounding area.
13
-------
azurlte and malachite being most abundant in domestic ores. Weathering
may also form small amounts of metallic native copper.
The ore deposits of northern Michigan are a unique occurrence of
primary origin; and the native copper is mixed with sulfide minerals.
Table 2-2 gives an analysis of the ore from this deposit. Except for
the elemental copper content, this analysis is similar to that of most
domestic ores, since it shows iron present in much higher concentration
than copper in a gangue rock of silica and alumina minerals.
The first step in the processing of an ore is to form a copper
concentrate, which consists of the copper minerals separated from most
of the gangue. These concentrates are an article of commerce, and
represent another raw material of this industry. Ores mined primarily
for other metals may be the origin of concentrates rich in copper, which
are sold to copper producers. This may constitute 5 to 10 percent of
all the primary copper that is mined in this country. Also, concen-
trates are regularly imported from other countries; domestic smelters
frequently process concentrates from Canada, South America, Australia,
and the Philippines.
The industry also imports copper from other countries at several
intermediate stages of processing. These imports include partially
smelted matte and crude anode or blister copper. Although they do not
represent a large fraction of the copper consumed in this country, most
copper imports are in the form of these intermediate products. A small
amount of high-grade ore that has received no on-site processing is also
imported.
The industry consumes other materials in various processing steps,
but not in large quantities. Mining and concentrating entail use of
explosives and small amounts of organic chemicals, and smelting requires
limestone and silica rock as fluxing materials.
Products
Commerce recognizes a number of different grades of copper, classi-
fied into two main groups. Relatively impure grades are directly pro-
duced in a copper smelter. These are sold for use in alloys or for
other special purposes, or they may be exported to be refined elsewhere.
More than 90 percent of the copper produced, herein, is refined in this
country into one of the electrolytic grades. Table 2-3 shows the
distribution of consumption of electrolytic copper in 1973. More than
two-thirds of it is used directly to manufacture wire and tubing. More
than half the electrolytic copper is cast at the refinery into wirebars
for direct use on wire and tubing forming machines.
14
-------
Table 2-2. TYPICAL ANALYSIS OF COPPER ORE USED AT WHITE PINE
COPPER COMPANY, MICHIGAN
Element
Cu
Ag
Au
A1203
Si02
CaO
Fe
MgO
Ni
S
Pb
As
Mo
Bi
Mn
Zn
Na
K
Co
Se
Percentage
1.0
0.0006
Trace
15.0
61.5
7.4
6.6
3.7
0.005
0.35
0.001
0.0005
0.002
0.0001
0.05
0.001
1.5
1.0
0.003
. 0.0005
15
-------
Table 2-3. CONSUMPTION OF REFINED COPPER
IN 19752
Consumer
Quantity,
metric tons
Wire mills
Brass mills
Secondary smelters
Chemical plants,
foundries, and
miscellaneous plants
961,864
398,097
1,445
32,659£
Estimated.
16
-------
Several by-product elements are isolated from the copper ores. In
1973, the copper industry segment produced all the arsenic and selenium,
manufactured in this country, almost all the platinum and palladium, and
almost half the gold, silver, and molybdenum. Except for molybdenum,
all of these were produced as purified metals or compounds. Molybdenum
was sold as concentrate, and most copper producers also reclaim zinc and
lead as a concentrate. Several copper smelters recover tellurium, and
one company, situated near steel mills, makes a high-grade iron sinter
from the iron pyrite in its ore. This same company manufactures sul-
furic acid as a major product, and most others have facilities to manu-
facturer it as a by-product. Three companies produce copper sulfate,
and two manufacture chemicals of a specialized nature in the same plant
with their copper operations.
Table 2-4 provides the basic 1974 statistics of this industry, and
Table 2-5 lists major sulfuric acid production of the industry.
Companies
Eight companies own and operate the sixteen conventional smelters
and seven electrolytic refineries in the United States. Table 2-6 lists
these companies, with applicable data. The three largest domestic
producers are Kennecott Copper Corporation, Phelps-Dodge Corporation,
and ASARCO, Inc.
Most of these companies own or control domestic mines that supply
at least part of their own needs. Table 2-7 lists the twenty-five
largest copper mines operating in 1974. Twenty of these were directly
owned by one of the producing companies. At least three other large
companies own mines or leaching operations intended primarily for pro-
duction of copper, as do several smaller companies. The 25 listed mines
produced more than 95 percent of the domestic copper in 1974. The
remaining five percent of domestic copper was produced from two or
three smaller mines, or as by-products of other mining industries.
Several of the larger U.S. copper companies have invested sub-
stantially in copper mines and production facilities in Mexico, Peru,
Chile, Canada, Republic of South Africa, and Zambia. Most of them
continue also to expand their U.S. mining interests. Table 2-8 provides
a survey of the expansions projected in 1976.
An estimated 40,000 persons are employed by the primary copper
industry. Most are engaged in mining and concentrating. In 1974,
industry employment was reported as 33,942 persons, of which 14,861 were
in open-pit mines, 9,545 in underground mines, and 4,536 in ore concen-
trating mills. The copper industry is the largest employer in Arizona,
Montana, Nevada, and Utah.
17
-------
Table 2-4. SALIENT STATISTICS OF THE PRIMARY COPPER INDUSTRY
IN THE UNITED STATES IN 1974
Ore produced, thousand metric tons
Open pit mine production, percent9
Underground mine production, percent"
Average yield of copper, percent
Primary copper produced, thousand metric tons
Mines, from domestic ores
Smelters, from domestic ores
Refineries, from domestic ores
Refineries, from foreign ore, matte, etc.
Exports, thousand metric tons
Unmanufactured
Refined
Imports, thousand metric tons
Unmanufactured
Refined
266,205.62
89
11
0.49
1,448.77
1,389.86
1,289.02
212.06
179.59
115.19
552.48
284.86
Copper metal production is 81 percent.
produced from dump leaching.
Copper metal production is 19 percent.
produced from in-piace leaching.
This includes copper metal
This includes copper metal
18
-------
Table 2-5. LARGEST BY-PRODUCT SULFURIC ACID PRODUCERS - 1974
Producer
American Smelting and Refining Co.
El Paso, Texas
Haydfcn, Arizona
Tacoma, Washington
The Anaconda Co.
Anaconda, Montana
Cities Service Co. , Inc.
Copperhill, Tennessee
The Inspiration Consolidated Copper Co.
Inspiration, Arizona
Kennecott Copper Corp.
Chino Mines Div., Hurley, New Mexico
Ray Mines, Hayden, Arizona
Utah Copper Div., Salt Lake City, Utah
Magma Copper Co., San Manuel, Arizona
Phelps Dodge Corp.
Ajo, Arizona
Morenci , Arizona
Hidalgo, New Mexico
Capacity
metric ton/year
154,000
299,000
60,000
313,000
l,143,000a
227,000
181,000
250,000
544,000
599,000
181,000
245,000
120,000
Input material contains pyrites and smelter gases.
Ref. Stanford Research Institute. 1974 Dictionary of Chemical Producers
United States of America. California 1974.
19
-------
Table 2-6. U.S. PRIMARY COPPER PRODUCERS
3,4
(Conventional smelting/refining operations)
NJ
o
Company
The Anaconda Company
ASARCO, Inc.
Cities Service Co., Inc.
Inspiration Consolidated
Copper Co.
Kennecott Copper Corp.
Magma Copper Company
Phelps -Dodge Corp.
White Pine Copper Company
Location
Anaconda, Montana
Great Falls, Montana
Tacoma, Washington
Hayden, Arizona
El Paso, Texas
Amarillo, Texas
Copperhi 1 1 , Tennessee
Miami , Arizona
Garfield, Utah
Hurley, New Mexico
Hayden, Arizona
McGill, Nevada
San Manuel, Arizona
Morenci , Arizona
Douglas, Arizona
Ajo, Arizona
El Paso, Texas
Hidalgo, New Mexico
White. Pine, Michigan
Description
Smelter
Refinery
Smelter/refinery
Smelter
Smelter
Refinery
Smel ter
Smelter/refinery
Smelter/refinery
Smelter
Smelter
Smelter
Smelter/refinery
Smelter
Smelter
Smelter
Refinery
Smelter
Smelter
Capacity
metric ton/yr
165,000
165,000
90,000
165,000
90,000
135,000
250,000
90,000
72,000
45,000
180,000
175,000
125,000
70,000
405,000
230,000
77,000
Number of
employees
1550
750
1000
450
900
2000
200
6000
1200
840
3700
2450
550
1300
800
3000
Date
plant built
1902
1893
1900
1912
1912
1976
M920
1911
1900
1939
1908
1954
1942
1903
1916
1930
1976
1952
-------
Table 2-7. TWENTY-FIVE LEADING COPPER MINES IN THE UNITED STATES IN 19744
Rank
1
2
3
4
5
6
7
8
9
10
11
12
13
Mine
Utah Copper
San Manuel
Morencl
Butte oper-
ations
Tyrone
Ray Pit
Pima
White Pine
Slerrita
Chi no
Twin Buttes
New Cornelia
Inspiration
County
and state
Salt Lake, Utah
Plnal, Arizona
Greenlee, Ariz-.
Silver Bow, Mont.
Grant, N. Mexico
Plnal, Arizona
P1ma, Arizona
Ontanagon, Mich.
Pima, Arizona
Grant, N. Mexico
Pima, Arizona
Pima, Arizona
Glla, Arizona
Company
Kennecott Copper Corp.
Magma Copper Co.
Phelps Dodge Corp.
The Anaconda Co.
Phelps Dodge Corp.
Kennecott Copper Corp.
Cyprus P1ma Co.
White Pine Copper Co.
Duval Sierrlta Corp.
Kennecott Copper Corp.
The Anaconda Co.
Phelps Dodge Co.
Inspiration Consolidated
Opera-
tion
A
A
B
C
C
D
B
A
E
A
C
B
F
Type of
mining
Open pit
Underground
Open pit
Open pit
under-
ground
Open pit
Open pit
Open pit
Underground
Open pit
Open pit
Open pit
Open pit
Open pit
Ore
grade,
percent
0.635
0.7
0.82
0.7
0.8?'
0.929
0.49
1.0
0.861
0.61
0.71
Annual
ore tonnage,
metric tons
29,000,000
16,000,000
14,000,000
aver 10,000,000
aver 10,000,000
9,500,000
16.500.000
7,000,000
6,000.000
>ver 10,000,000
7.400.000
7.000.000
Products
Copper, molybdenum,
gold, silver, sul-
furic acid, selenium,
ammonium perrhate,
platinum, palladium
Cathode copper, molyb-
denum, electrolytic
slimes, sulfuric acid
Copper
Copper, silver, gold
Copper, gold, silver
Copper
Copper, molybdenum,
silver
Copper, silver
Copper, molybdenum,
silver
Copper, molybdenite
Copper cone.
Copper
Copper, silver, gold,
selenium, rhenium
Mineralization
Chal copy rite
Copper sulphide
Copper sulphide
ChalcopyHte
Porphyry copper
Copper sulphide
Chalcocite. chryso-
colla, malachite.
•write
-------
Table 2-7 (continued). TWENTY-FIVE LEADING COPPER MINES IN THE UNITED STATES IN 1974
Rank
14
15
16
1.7
18
19
20
21
22
23
24
25
Mine
Mission
Nevada Mines
Yerington
Silver Bell
Copper Cities
Mineral Park
Superior Olv.
Copper Queen
Continental
Bagdad
Esperanza
Battle Moun-
tain
Property
County
and state
Plma, Arizona
White Pine, Nev.
Lyon, Nevada
P1ma, Arizona
Gil a, Arizona
Mohave, Arizona
Plnal, Arizona
Cochise, Arizona
Grant, N. Mexico
Yavapal , Ariz.
Pima, Arizona
Lander, Nev.
Company
American Smelting and
Refining Co.
Kennecott Copper Corp.
The Anaconda Co.
American Smelting and
Refining Co.
Cities Service Co.
Ouval Corp.
Hagna Copper Co.
Phelps Dodge Corp.
UV Industries, Inc.
Cyprus Mines Corp. •
Duval Corp.
i Duval Corp.
Opera-
tion
C
D
C
G
G
G
B
B
B
G
G
B
Type of
mining
Open pit
\
Open pit
Open pit
Open pit
Open pit
Open pit
Underground
Open pit
under-
ground
Open pit
under-
ground
Open pit
Open pit
Open pit
Ore
grade,
percent
0.601
0.5
0.4149
4.5
0.6
4.06
0.7
Ore tonnage,
netrlc tons
1,000,000 to
10,000,000
6,000,000
1.000,000 to
10,000,000
over 10,000,000
3.000.000
1,000,000 to
10,000,000
625,000
1,000,000 to
10,000.000
450.000 to
1,000.000
1,900.000
1,000,000 to
10,000.000
1 ,000,000 to
10,000,000
Products
Copper, silver, moly-
bdenum
Copper
Copper cone.
Copper, molybdenum,
silver
Copper
Copper, molybdenum,
silver
Copper cone.
Copper, gold
Silver
Copper, zinc
Copper, molybdenum,
silver
Copper, molybdenum,
silver
Copper, gold, silver
concentrate
Mineralization
Chalcopyrlte,
molybdenite
Chalcocite. chaleo
pyrite
Chalcopyrite
Copper, molyb-
denum, silver
Chalcopyrite,
bornite, chal-
cocite
Copper sulphide
and oxide
Porphyry
Copper, gold.
silver
A - Mine, concentrator, smelter, refinery
B - Mine, concentrator, smelter
C - Mine, concentrator
0 - Mine, concentrator, smelter, ore leach plant/precipitate plant
E - Hlne, concentrator, hydrometallurglcal recovery
F - Mine, concentrator, smelter, refining, leaching-electro mining plant
G.- Nine, concentrator, leach plant
-------
Table 2-8. 1976 SURVEY OF MINE AND PLANT EXPANSIONS IN THE UNITED STATES
Company
Anaconda
Anamax
Brenmac Mines
Cities Service
Continental Copper
Cyprus Bagdad Copper
Ouval
Magma Copper
Noranda Exploration
Phelps Dodge
Location
Tooele, Utah,
Twin Buttes, Ariz. ,
Sultan, Washington,
Miami East, Ariz. ,
Marble Peak, Ariz.,
Bagdad, Arizona,
SierHta. Ariz.,
San Manuel , Ariz. ,
Rhlnelander, Wise.,
Safford, Arizona,
Project
UGmlco
ml/Co
ml/Co
UGmlsm
ex
ex
Pi
ml
OPml
UGmlco
Planned
10M
120M
9.9H
2M
2M
70M
22. 5M
75M
350H
Now
63M
20K
62. 5M
Units
tpd cone.
tpy Cu
tpy Cu
tpd Cu ore
tpd ore .
tpy Cu
tpy Cu
tpd ore
tpy ore
Start
1979
1978
1977
1978
1976
Class
A
C
C
A
AB
A
C
A
C
Notes
Carr Fork Mine: also Mo, Ag, Au
Weak markets render timetable questionable
Proved OR: 18.59MM t 0.35% Cu, 0.04% molyb-
denite
OR 50MH t 1.95% Cu ore. May be some small
production 1n 1976.
OR 12MM t 2.2% Cu. Operation expects a 20-
year lifespan.
OR 300MM t 0.49% Cu. Smelter on site post-
poned.
Chloride leach process recovers electrolytic
copper.
Cu/Zn mine. Dependent on environmental
clearance.
OR 400MM tons of 0.72% Cu. Indefinitely
postponed.
NJ
U)
Abbreviations:
UG - underground tpd - tons per day, tpy - tons per year
•1 - nine so - smelter
CO - concentrator ex - complex
pi - plant
-------
Table 2-9. ANNUAL GENERATION OF HAZARDOUS POLLUTANTS FROM PRIMARY COPPER INDUSTRY - 1968
(metric tons)
to
it*.
Source
Mining
Roasting
Reverberatory furnace
Converters
Material handling
Total
Arsenic
741
329
946
206
2222
Percent
of total3
10.07
4.48
12.87
2.80
30.22
Cadmium,
Neg.
189
77
222
49
537
Percent
of totar
7.59
3.12
8.95
1.96
21.62
Copper
156
2386
1023
3068
681
7314
Percent
of total
1.41
21.54
9.23
27.70
6.15
66.03
Fluorides
164
71
211
47
493
Percent
of totala
0.13
O.OS
0.16
0.04
0.38
Lead.
284b
104
44
134
30
,596
Percent
of total
3.72
1.37
0.58
1.76
0.39
7.82
Selenium
14
6
18
S
43
Percent
of total
1.99
0.94
2.57
0.59
6.09
fPercent of total of this pollutant.
Combined total from copper, zinc and lead mining.
-------
Environmental Impacts
Smelting is the most important source of environmental problems in
the copper industry.
Emissions of air pollutants from copper processing are of concern,
particulary with respect to those that are potentially hazardous to
human health. Among the known trace elements of concern are arsenic,
emitted as arsenic trioxide, and cadmium, known to be a threat to human
health. The copper industry is a primary source of arsenic emissions,
producing about 30 percent of total arsenic emissions in the United
States.
Copper smelters are a major source of sulfur dioxide, emitting 80
percent of the total amount of S02 emitted from the copper, lead, and
zinc industries. The industry is implementing control methods to re-
cover some of the S02 as a salable product. Fifteen percent of the S02
generated by the industry occurs in fugitive dust.
Most of the water used by the industry is recycled. The mine
wastewater may contain acid and dissolved metals. Mill tailings may
also contain heavy metals. Smelter and refining wastes often contribute
a heavy load of dissolved metals to the tailings pond. These wastes can
affect the quality of the decant water as well as effluent volumes. The
slag from the industry, which is dumped, contains many elements. Table
2-9 presents approximate quantities of hazardous pollutants from the
primary copper industry. Primary copper smelting and fire refining
produce approximately 3 metric tons of wastes containing slag, sludge,
and dust (including acid plant sludge) per metric ton of product. The
electrolytic refineries produce about 2.4 kg of slag per metric ton of
product.
In 1974 the primary copper industry produced about 6.1 million
metric tons of solid waste containing about 1.56 percent hazardous
constituents. Of this total, the emissions from the plants with fire
refining facilities were 4.9 percent sludge, 94.5 percent slag, and 0.6
percent dust.
Bibliography
1. Mining Informational Services of the McGraw-Hill Mining Publica-
tions. 1975 E/MJ International Directory of Mining and Mineral
Processing Operations. McGraw-Hill Inc., 1975.
2. U.S. Department of Interior, Bureau of Mines. Mineral Industry
Surveys. Copper in 1975. March 26, 1976.
25
-------
3. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA 450/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Protection
Agency. Research Triangle Park, North Carolina. October 1974.
4. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Copper
Segment of the Nonferrous Metals Manufacturing Point Source
Category. EPA 440/1-75/032b. November 1974.
26
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INDUSTRY ANALYSIS. COPPER
The environmental impacts of many industries, including the primary
copper industry, have received wide attention and have been the subject
of many industrial and governmental studies. Emissions of S02 aftd their
impacts on the atmosphere are considered especially important.
This industry analysis examines each individual production opera-
tion, called here a process, to examine in detail its purpose and its
actual or potential effect on the environment. Each process is examined
in the following aspects:
1. Function
2. Input materials
3. Operating conditions
4. Utilities
5. Waste streams
6. Control technology
7. EPA classification code
8. References
The only processes included in this section are those that are
either operating in the United States, are under construction, or have
been reported as being under active consideration for inclusion in a
U.S. facility. Figure 2-1 is a flowsheet showing these processes, their
interrelationships, and their major waste streams.
27
-------
Figure 2-1. Copper industry flow sheet.
28
-------
Figure 2-1 (continued). Copper industry flow sheet,
29
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PRIMARY COPPER PRODUCTION PROCESS NO. 1
Mining
1. Function - Rock containing enough copper to justify its recovery is
removed from the ground and transported to a concentrator plant. Mining
methods are determined by the size, depth, and configuration of the ore
body, as these are adaptable for underground or open pit mining. The
capability for high productivity in large-scale open pit operations has
made possible the development of large deposits of relatively low-grade
porphyry ores; in 1973, 83 percent of all the ore mined in the United
States came from open pits.'
In an open-pit mine, holes for placement of explosives are drilled
behind the face of a near-vertical bank. Other explosives are placed in
secondary drill holes. The explosives reduce the rock to sizes that can
be handled by power shovels or other mechanical equipment. Shovels load
the ore into trucks or railroad cars or onto belt conveyors for trans-
portation to the concentrator plant.
2. Input Materials.- Important copper ore minerals are listed in Table
2-1. Chalcopyrite, bornite, and enargite are considered the primary
minerals that were formed deep underground by igneous processes. The
other sulfide minerals were formed by the leaching action of underground
water in the absence of oxygen. When oxygen was present, the sulfur was
oxidized, and minerals such as chrysocolla, azurite, and malachite were
formed. Native copper is sometimes found in these oxidized deposits.
Porphyry deposits are now the major source of the world's copper.
Porphyry is the term applied to the type of deposit in which the copper
minerals are uniformly distributed in thin veins through a rock com-
posed of other minerals. The copper content is between 0.6 and 2 per-
cent.2 The copper ores of southwestern United States are from porphyry
deposits.
Explosives used in copper mining are almost always a mixture of
ammonium nitrate and fuel oil (AN-FO). Some mines add sodium nitrate to
make the explosive slightly more powerful.3 No values are available for
the consumption of these explosives.
3. Operating Conditions - Most copper ores are mined in the arid
regions of the West or Southwest, where open-pit operations continue
through both hot and cold weather. Other mining operations in Michigan
and Tennessee are underground mines where more constant temperatures
prevail.
30
-------
Table 2-10. RAW WASTE LOAD IN WATER PUMPED FROM SELECTED COPPER MINES'
Parameter
Flow
PH
IDS
TSS
Oil & grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Nn
Cd
N1
Ho
Sr
Hg
Pb
Un
Concentration
(mq/i)
3,815.3m3/day
7.37a
29,250
69
<1.0
<4.5
819
2.19
'-•-• 0.87
<0.04
<0.077
0.60
<0.07
2.8
<0.5
<0.1
2.22
<0.02
<0.05
<0.5
119
<0.0001
<0.1
derqround etna
Raw waste load per unit ore mined
koyiOOO metric tons
17.28 m3/100b metric tons
7.37a
5,053.9
11.9
<0.173
<0.778
141.5
1 0.378
0.150
<0.007
<0.013
0.104
<0.012
0.484
<0.086
<0.017
0.384
<0.003
<0.009
<0.086
20.6
<0. 00002
<0.017
Open-pit Mine
Concentration
409 nfVday
6.96a
1,350
2 .
7
10
4
0.07
.1.05
<0.06
0.096
<0.2
<0.01
0.1
<0.5
<0.1
0.9
<0.03
<0.05
<0.2
0.8
<0.0001
<0.5
Raw waste Toad per unltoire mined
kg/1000 wtrlc tons
75 iB3/1000 netrlc tons
6.96a
101
0.2
0.5
0.75
0.3
0.005
0.08
<0.005
0.007
<0.02
<0.0008
0.008
<0.04
<0.008
0.07
<0.002
<0.004
<0.02
0.06
<0. 000008
<0.04
aV»lue In pH units
-------
4. Utilities - In most mines, electrically operated power equipment is
used for drilling, loading, and hauling. In 1973, 880 kilocalories of
energy was consumed by the mining process per kilogram of copper pro-
duced. 4
A small amount of water is needed for equipment cooling, drill
lubrication, dust control spraying, equipment washing, and sanitary
facilities. Occasionally the water sent to the mine is reused water
from the concentrator plant or tailings pond.
5. Waste Streams - Mining operations generate fairly large amounts of
dust from drilling, blasting, loading, and transporting operations. One
estimate of 110 grams of fugitive dust per metric ton of ore mined is
given as the average for several types of nonferrous mining.5 The dust
composition is dependent on the character of the ore being mined, and
there is a large variation in particle size.
Large amounts of solid wastes are generated in a mining operation.
Overburden stripped to uncover an ore body, shaft and tunnel spoil, and
low-grade ore (less than 0.4% copper) found within the mine are disposed
of near the mine. The amount varies widely, from as little as 0.004
metric ton per metric ton of ore up to 15 metric tons per metric ton of
ore mined. Average quantities in 1973 were reported as 2.65 metric tons
per metric ton for open-pit mines, and 0.13 metric ton per metric ton
for underground mines.6 These spoils contain small and varying amounts
of copper minerals, sometimes minerals of other metals, and large amounts
of the native rock of the region.
Wastewater from copper mining comes from seepage or runoff from the
mine or spoil dumps, and from the water sent into the mine for utility
uses. The amount of wastewater from open-pit copper mines ranges from
zero to 0.3 cubic meter of water per metric ton of ore mined. From
underground mines, the amount ranges from 0.008 to 4.0 cubic meter per
metric ton of ore.7 Chemical characteristics are typical of those from
any sulfide mine, as described in a later section of this report. Table
2-10 gives analyses of waters from two copper mines.
6. Control Technology - The only control provided for fugitive dust is
the manual use of water sprays, to be used when needed. Most open pit
copper mines are very large, with sufficient natural ventilation that
dust conditions are not unbearable.
Mining companies attempt to locate spoil dumps where natural
seepage will not contaminate a stream or underground aquifer. Otherwise
there is little control of these solid wastes in the copper industry.
There is no control of blowing dust and no attempt to reclaim or natura-
lize the dump areas.8
32
-------
Mine water wastes and seepage from the spoil dumps are major po-
tential sources of water pollution from the primary copper industry.
Although treatment of these wastes is discussed in detail in the section
covering all sulfide mines, two characteristics of copper mines differ
from those in the lead and zinc industries. First, their location
greatly simplifies control of water discharges. All of the large open-
pit mines are in regions of deficient rainfall, and some are in desert
areas. Natural evaporation within the pit greatly reduces the volume of
wastes that must be pumped out, and seepage from spoil dumps rarely
enters a stream. The water that accumulates is in many cases disposed
of merely by pumping it onto a nearby flat area, where it either seeps
into the alkaline soil or evaporates. It is likely that most of the
dissolved metals are converted to insoluble compounds by this process,
and officials of several mining companies state that they have shown
that none of these waters has entered underground supplies. In the
vicinity of some of these mines, copper, zinc, selenium, and arsenic are
detected in analysis of water from springs and wells, in concentrations
usually less than 0.1 milligrams per liter; it is not clear whether this
amount exceeds the natural concentrations in a highly mineralized
region.1 These alkali waters naturally have total dissolved solids that
can be several thousand milligrams per liter.
Near some copper mines, also, the practice of leaching low-grade
ore is being practiced. This practice is described in Process No. 30,
along with its effect in the control of mining wastewaters.
7. EPA Source Classification Code - None
8. References -
1. Mineral Facts and Problems, Washington, D.C. U.S. Department
of the Interior, Bureau of Mines, 1970.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
4. Energy Consumption in Domestic Primary Copper Production, U.S.
Bureau of Mines.
5. Davis, W.E. National Inventory of Sources and Emissions:
Copper, Selenium, and Zinc. U.S. Environmental Protection
Agency, (NTIS), Research Triangle Park, North Carolina.
PB-210 679, PB-210 478, and PB-210 677. May 1972.
33
-------
6. Minerals Yearbook. Washington, D.C. U.S. Department of the
Interior, Bureau of Mines, 1973.
7. Development Document for Interim Final and Proposed Effluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source
Category Volumes I and II. Environmental Protection Agency,
Washington, D.C., EPA/1-75/032-6. February 1975.
8. Dayton, S. The Quiet Revolution in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975. p.
87-89.
34
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PRIMARY COPPER PRODUCTION PROCESS NO. 2
Concentrating
1. Function - Sulfide ore from the mine is separated by the concentra-
tion process into two or more fractions. The fractions rich in valuable
minerals are called concentrates, and the waste rock, low in metals
content, is called the gangue. With this process, ore that usually
contains less than 1 percent copper, is concentrated into a fraction
analyzing from 20 to 30 percent copper. At least 85 percent of the
ore copper content is recovered in the concentrate.
Concentrating consists of milling the ore, crushing and grinding it
to a fine powder, and then separating the minerals by froth flotation.
In milling, the ore is sent through crushers and then through fine
grinders. Between each stage it is classified (screened), and the final
milled product is a mixture of particles between 65 and 200 mesh.
In the last stages of milling, water is added along with chemicals
to condition the ore for froth flotation.
Flotation is a continuous process that uses compressed air and
various flotation chemicals to separate the ore into its fractions. By
proper selection of additives, certain minerals are caused to float to
the surface and are removed in a foam of air bubbles, while others sink
and are carried out with the slurry. The ore passes through many flota-
tion stages to accomplish this separation. The chemicals that are added
are classified as "frothers", which create the foam; "collectors", which
cause certain minerals to float; and "depressants", which cause certain
minerals to sink.
In the flotation of copper ores, the frothers most often used are
pine oils or long-chain alcohols. Lime is usually added in the final
stages of grinding, both to adjust the pH of the slurry to an optimum
level and to act as a depressant for iron pyrite. Various xanthates or
dithiophosphates act as collectors for the valuable sulfide minerals,
and the copper and other recoverable minerals come off with the frothj
The gangue does not float and is discarded as "tailings".
After this initial separation, the valuable minerals are sent
through further flotation stages that further separate them by selective
or differential flotation. By use of proper collectors or depressants,
the concentrates may be upgraded to remove more iron pyrite. In some
cases, other fractions high in lead and zinc, or molybdenum and rhenium,
may be produced. These are usually sold to processors in the industries
handling those metals. The copper ores of the west are a prime source
for molybdenum; and to separate this fraction, the concentrate must be
steam stripped to remove the collector originally added.2
35
-------
Occasionally, a concentrator will batch-treat a copper concentrate
with cyanide to dissolve its silver and gold content. After separating
the leached solution from the concentrate, zinc metal is added to re-
precipitate the precious metals.
Concentrates are dewatered by clarification and filtration. They
may be partially dried to simplify handling and shipment, or may be more
completely dried for direct feed to a smelting furnace (see Process No.
5). Ten of the sixteen conventional smelters in this country have
concentrator plants on-site or nearby.3
Table 2-11 shows typical composition of copper concentrates;
compositions vary with the character of the ore and the amount of pro-
cessing employed.
2. Input Materials - Only sulfide ores of copper can be successfully
separated by the flotation process. Oxidized ores are treated by hydro-
metallurgical processes (see Process No. 30).
Lime is used for pH adjustment and as a pyrite depressant. Quan-
tities added vary between 0.9 and 18.0 kilograms per metric ton of ore
processed.5 The pine oil frothers are usually consumed at a rate of
about 0.09 kilogram per metric ton of ore.5 No data are available on
the quantities of long-chain alcohols required if they are substituted.
Table 2-12 lists some of the chemicals that are used as collectors
in the flotation process. The use and quantity of any one of these
materials depends on the mineral assemblage particular to each ore type.
Alkyl-based organic molecules are more commonly used than aryl com-
pounds.
Miscellaneous compounds such as cyanides, zinc dust, various filter
aids, and inorganic salts are occasionally used in small quantities.
3. Operating Conditions - Most portions of this process are carried
out at ambient temperatures in closed buildings. At few points tempera-
tures may approach 100°C (i.e., steam stripping for molybdenite concen-
tration). Occasionally circulating streams are heated slightly to
retain efficiency during cold weather.
4. Utilities - In 1973, usage of water at 21 copper concentrators
ranged from about 100 to 500 cubic meters of water per metric ton of
concentrate produced, the amount depending on the complexity of the
process employed.3 in the same year, concentrators consumed about 5.7
x 109 kg-eal (6630 million kWh) of electricity, which is about 344,000
kg-cal (400 kWh) per metric ton of primary refined copperJ>6 The
greater part of this electricity was used to operate the crushing and
grinding equipment, with a smaller amount for production of compressed
air.
36
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Table 2-11. ANALYSIS OF COPPER CONCENTRATE
Element
Cu
S
Pb
Fe
Zn
Ag
Au
Pt etc.
Pd
As
Sb
Bi
Se
Te
Re
Ni
Co
Cd
In
Ge
Sn
Cl
F
Al
Si
Ca
Mg
Mo
Mn
Flotation reagents
Composition, %
20 - 50
30 - 38
tr. - 0.67
20 - 30
0.2 - 4.0
0.13
31.5a
tr.
tr.
tr. - 4.0
tr. - 0.36
tr. - 0.05
tr. - 0.03
tr.
tr.
tr. - 0.1
tr. - 0.02
tr. - 0.01
tr.
tr.
tr.
0.05
0.05
Varies
Varies
Varies
tr.
tr.
tr.
tr.
Value for Au in grams/metric ton.
tr. = trace
37
-------
Table 2-12. TYPICAL FLOTATION COLLECTORS
1
Type
Formula
Xanthate
Dithiophosphate
Dithiocarbamate
Thiol (mercaptan)
Thiocarbanilide
Fatty acid soaps
Arenesulfonate or
alkylarenesulfonate
Alkyl sulfate
Primary amine
Quaternary ammonium salt
Alkylpyridinium salt
ROCSSNa
(RO)2PSSNa
R2NCSSNa.
RSH
(CgH5NH)2CS
RCOONa
RS03NA
ROS03Na
RNH3C1
RN(CH3)3C1
RCCH/IN-HC1
0 H
R is the abbreviation for an alkyl group
such as CH3(CH2)n. Although alkyl com-
pounds are common, alkyl aryl compounds
may also be used, as in alkylarenesul-
fonates.
38
-------
5. Haste Streams - The handling and milling of dry ore is the prin-
cipal source of air pollutants in this process. Items of equipment are
always enclosed, but transitions between pieces of equipment are diffi-
cult to seal tightly. Ore classifiers are not always completely sealed.
Dust quantity is reported as about 1 kilogram per metric ton of ore.2
By far the largest source of solid waste in the industry is the
production of tailings by the concentrator plant. More than 95 percent
of all the ore brought from the mine is discharged from this process as
tailings; this quantity totals approximately 158 million metric tons of
waste material each year from the industry.!>°>' Tailings are composed
primarily of the common rock-forming minerals, but they also contain
around 15 percent of the heavy metals originally found in the ore, and
usually, at present, much of the iron pyrite. Production of higher-
grade concentrates to minimize air pollution has increased the propor-
tion of pyrites in the tailings. In this solid waste, the minerals have
been pulverized and intimately mixed, and are therefore subject to
weathering much more rapidly than rock masses of similar composition.
They form a soil that is usually highly acidic and that contains no
plant nutrients.
This process also contributes the largest amount of wastewater in
the industry. The ore flotation water is used to sluice the tailings
into a pond. Although part of the water is recycled to the plant, the
remainder is discarded. Character of the wastewater is discussed in
more detail in the section of this report that examines effluent from
the concentrators of all sulfide mining industries. Excluding the
amount lost by evaporation in the tailings pond, the volume of waste-
water from this process will equal the water consumption, ranging from
100 to 500 cubic meters per metric ton of concentrate.2
Reported analyses indicate that the water from the concentrator may
contain up to 3500 milligrams per liter of dissolved solids, from 0.01
to 0.1 milligrams per liter of cyanides, and the following ranges of
metallic elements:2
Element
Arsenic
Antimony
Cadmium
Copper
Cobalt
Iron
Manganese
Mercury
Molybdenum
Nickel
Concentration, mg/1
0.07 approximately
0.2 to 1.0
0.02 to 0.05
0.08 to very high
0.04 to 1.68
0.1 to 2.0
0.05 to 4.8
0.001 to 0.05
0.2 to 20
0.05 to 3
39
-------
Element
Lead
Selenium
Silver
Strontium
Zinc
Concentration, mg/1
0.01 to 3
0.003 to 0.02
0.1 approximately
0.03 to 2.5
0.05 to 8.50
The water may also contain thiosulfates and thionates, and the
materials, both inorganic and organic, used as flotation additives.
6. Control Technology - Dust from the milling operations is generally
reduced by drawing air through the equipment and collecting the dust
with cyclone separators. This is both a dust control and an integral
part of the process, since it allows these small particles to bypass one
or more crushing and grinding operations. Fugitive dust is usually
uncontrolled, unless the amount being lost justifies economically the
installation of equipment for its recovery.
Control of wastewater is discussed in more detail in following
sections. This waste, although it is the major one, is rarely con-
trolled independently, since waters from many sources find their way
into the tailings pond. Occasionally this source is kept separate; the
pond itself represents one stage of treatment.
Disposal of the tailings is a major problem in this industry. As
the tailing pond becomes filled with solids, the pond is abandoned or
the tailings may be dredged or mechanically moved into a pile and the
pond reused. There is no universal solution for disposal of such vast
quantities of solid materials; each concentrator plant requires separate
study. Problems of secondary pollution of water are discussed in the
section on water managment from all sulfide concentrating plants.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Hallowell, J.B. et al. Water Pollution Control in the Pri-
mary Nonferrous Metals Industry - Volume I. Copper, Zinc,
and Lead Industries. Environmental Protection Agency,
Washington, D.C., EPA-R2-73-274a. September 1973.
3. Development Document for Interim Final and Proposed Effluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source Cate-
gory Volumes I and II. Environmental Protection Agency,
Washington, D.C., EPA/1-75/032-6. February 1975.
40
-------
4. Little, A.D. Economic Impact of New Source Performance
Standards on the Primary Copper Industry: An Assessment.
U.S. Environmental Protection Agency, Washington, D.C.
C-76072-20. October 1974.
5. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/0'32-a. February 1975.
6. Dayton, J. The Quiet Revolution in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975.
7. Minerals Yearbook. Washington, D.C. U.S. Department of the
Interior, Bureau of Mines, 1973.
41
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PRIMARY COPPER PRODUCTION PROCESS NO. 3
Multipie-Hearth Roasti ng
1. Function - Roasting is frequently the first of the pyrometal1ur-
gical processes applied to the copper ore concentrate at the copper
smelter. The purpose of roasting is to reduce the sulfur content so
that subsequent processes operate efficiently. Roasting also removes
the water from the concentrate, volatilizes some of the arsenic and
antimony, and preheats the ore before it is charged to the reverberatory
furnace.
Roasting is accomplished either with a multiple-hearth or fluidi-
zation process (See Process No. 4). Another alternate is the Noranda
process (Process No. 11), which combines most of the functions of
roasting, smelting, and converting. In the multiple-hearth roaster,
concentrate is introduced at the top of a cylindrical vessel fitted with
a series of round horizontal trays, or hearths. The ore is raked across
each hearth in turn until it is discharged from the bottom of the
cylinder. Air is admitted into the roaster, along with a fuel if
necessary to maintain adequately high temperature.
Most of the chemical reactions that occur in the roaster are with
the pyrite in the concentrate rather than with the copper minerals.
Copper has a higher affinity for sulfur, whereas iron combines pre-
ferentially with oxygen. Admitting a limited amount of air, therefore,
causes the pyrite to oxidize, producing iron oxide and sulfur dioxide
gas.'
The heat of the roasting process generally vaporizes much of the
arsenic and some of the antimony in the ore, and these "fumes" leave the
roaster with the SOg gas.
Multiple-hearth roasting is currently in use at four domestic
copper smelters. The roasters are built to handle from 140 to 725 tons
of concentrate per day.2 There appears to be a trend away from their
use, except in "custom" smelters, since with higher-grade concentrates
the cost of operation frequently outweighs the benefits realized.'
Where roasting is still profitable, fluidization roasting (Process No.
4) offers several advantages. Custom smelters usually require roasters
to provide operating flexibility.
2. Input Materials - Copper concentrate, as received from the copper
concentrator, is the only input. Composition is shown in Table 2-11
(Process No. 2).
42
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3. Operating Conditions - Multiple-hearth roasters generally operate
at temperatures from 760°C on the bottom hearth down to around 200°C on
the top hearth.3 The roaster operates at approximately atmospheric
pressure.
4. Utilities - With the concentrates now being used, some fuel in the
form of oil or natural gas is always required. If concentrates are
especially high in sulfur content (24% or more), sufficient heat is
released by the burning sulfur and supplemental fuel is required only to
preheat the roaster at start-up (autogenous roasting). A reported
energy requirement is 280,000 kg-cal (1,100,000 Btu) per metric ton of
copper produced.^
Electricity is used to drive the roaster rakes and for auxiliary
materials handling equipment. Approximately 1360 kg-cal/min (95 kilo-
watts) are required for each operating hour for a plant of 100,000
metric tons of copper per year capacityJ3
5. Waste Streams - Gases leaving most present-day multiple-hearth
roasters are too weak in $03 gas for direct use by a sulfuric acid plant
unless the inlet air can be enriched in oxygen or unless air infiltra-
tion is reduced. The S02 concentration is variable, being dependent on
the required degree of sulfur removal and the amount of fuel necessary
to maintain roasting temperatures. With lower-grade concentrates, 20 to
50 percent of the total S02 generated by smelter facilities once came
from the roaster. Since fuel gas was not required, the effluent con-
tained from 5 to 10 percent S02.^ Published data on current operations
are inconsistent and contradictory; a recent report gives SQ2Qconcen-
trations in effluent from roasters at 0.5 to 2 percent.2,6,7,8,9,10,11
Gas from roasting is a steady stream, however, and if sufficiently
concentrated in S02 is otherwise suitable for sulfuric acid manufacture.
Emissions of particulate matter from most multiple-hearth roasters
are little affected by operational changes. About 75 kilograms of
particulates are produced per metric ton of copper produced.8,9,10,12,13
Fifteen percent is present in sizes below 10 microns.14 Table 2-13
presents typical particle-size profiles. Although composition is de-
pendent on the ore, particulates may be expected to contain the more
volatile elements, such as arsenic, antimony, selenium, zinc, mercury,
bismuth, rhenium, and lead. These will leave the roaster as vaporized
materials. Some copper and iron will be physically carried over by the
gas stream. Table 2-14 gives a weight analysis of particulate and fume
emissions from a multiple-hearth roaster. Table 2-15 lists the typical
levels of volatile metals found in copper ore concentrates. These
metals apparently appear in these dusts as sulfates, sulfides, oxides,
chlorides, and fluorides, but it is not known which of the metals is
combined with each negative radical.
43
-------
Table 2-13. TYPICAL SIZE PROFILE OF MULTIPLE-HEARTH
COPPER ROASTER EFFLUENTS15
Size, microns
Percent by weight
Entrained particles, carried from the
roaster as solids.
230 - 218
149 - 230
100 - 149
74 - 100
44 - 74
28 - 44
20 - 28
10 - 20
< 10
4.6
4.0
5.3
7.4
10.6
12.8
6.8
8.0
10.5
Sublimed particles, condensed from
metallic vapor.
0.5 - 10
30.0
44
-------
Table 2-14. CONCENTRATION AND WEIGHT ANALYSIS OF PARTICULATE
EFFLUENTS FROM A MULTIPLE-HEARTH COPPER ROASTER15
Concentration
Entrained particulates
1.4 g/Nm3
Sublimed particulates
0.6 g/Nm3
Effluent
Chemical
Cu
Fe
S
As
Sb
Pb
Zn
Sn
Cd
Ni
Mn
Se
Si02> CaO
CaS04
Op (oxides)
inerts
AS2°3
Sb2o3
inerts
% Weight
23.8 - 34.5
21.2 - 30.7
1.7 - 2.5
tr.
tr.
tr.
tr.
tr.
tr.
tr.
tr.
tr.
10 - 15
13 - 19
0.8
tr. - 17
tr. - 13
tr.
45
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Table 2-15. TYPICAL LEVELS OF VOLATILE METALS IN
DOMESTIC COPPER ORE CONCENTRATIONS5
Lead
Zinc
Arsenic
Cadmi urn
Beryllium
Vanadium
Antimony
Tin
Selenium
Concentration
level
<5000 ppm
5000 ppm-<2%
>2%
1%
<1000 ppm
<10 ppm
<100 ppm
<200 ppm
>200- 500 ppm
>5000 ppm
<1000 ppm
Percent of
concentrates
surveyed
96
2
2
67
1
88
10
2
100
100
100
97
3
1/2
100
46
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Any organic materials that enter the roaster with the copper con-
centrates are vaporized or decomposed in the roaster.
The multiple-hearth roaster produces no liquid wastes.
6. Control Technology - At present, none of the operating multiple-
hearth roasters is equipped with controls for S02 emissions. The only
suitable controls are the various scrubbing systems, as outlined for
reverberatory furnace gases (Process No. 6).
Proposals have been made that enrichment of the air to a roaster
with oxygen would increase its S02 content to a level where it could be
used as feed to a sulfuric acid plant. This proposal has been rejected
on several grounds by smelter operators. One company that recovers
arsenic from flue dusts maintains that oxygen enrichment causes the
arsenic to oxidize to AS204 and prevents its removal by current tech-
niques. 16 Most multiple-hearth roasters also cannot mechanically
withstand the high temperatures caused by oxygen enrichment.
A variety of devices are used in various combinations for control
of particulate emissions. Larger particulates are occasionally separat-
ed with cyclone collectors or with "balloon flues". The latter are
oversized ducts in which flue gas velocity is reduced enough that the
particles settle by gravity. Removal efficiency is 30 to 60 percent.'6
Cyclones can remove 80 to 85 percent of the solids, but require an
addition of energy to compensate for pressure drop. Smaller particles
in the gas stream are separated by hot gas electrostatic precipitators,
or the gas may be cooled with water sprays before entering an ESP unit
or a baghouse. As described further in the Air Management Section of
this report, control is not complete, especially in regard to sublimed
particles.
The collected dusts are returned to the metallurgical processing,
usually to the smelting furnace. One smelter extracts arsenic trioxide
before returning the dusts to the furnace. Excessive accumulation of
impurities causes dusts to be discarded but the means of disposal has
not been reported.
7. EPA Source Classification Code - 3-03-005-02
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Compilation and Analysis of Design and Operation Parameters
for Emission Control Studies. Pacific Environmental Services,
Inc. (Individual draft reports).
47
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3. Control of Sulfur Oxide Emissions in Copper, Lead, and
Zinc Smelting. Bureau of Mines Information Circular 9527,
1971.
4. Fejer, M.E. and Larson, D.H. Study of Industrial Uses of
Energy Relative to Environmental Effects. U.S. Environmental
Protection Agency. Research Triangle Park, North Carolina.
July 1974.
5. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
6. Copper Smelters. In: Compilation of Air Pollutant Emission
Factors, Second Edition, Environmental Protection Agency,
Research Triangle Park, North Carolina. AP-42. April 1973.
7. Donovan, J.R. and P.O. Stuber. Sulfuric Acid Production from
Ore Roaster Gases. Journal of Metals. November 1967.
8. Exhaust Gases from Combustion and Industrial Processes,
Engineering Science, Inc., Washington, D.C., October 2, 1971.
Distributed by National Technical Information Center.
9. High, M.D. and M.E. Lukey. Exhaust Gases from Combustion and
Industrial Processes. U.S. Environmental Protection Agency,
(NTIS), Durham, North Carolina. PB-204 861. October 1971.
10. Measurement of Sulfur Dioxide, Particulate and Trace Elements
in Copper Smelter Converter and Roaster/Reverberatory Gas
Streams. Environmental Protection Agency, Washington, D.C.
EPA 650/2-74-111. October 1974.
11. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW.
June 1969.
12. Vandegrift, A.E. (Dr.), L.J. Shannon (Dr.), P.G. Gorman,
E.W. Lawless (Dr.), E.E. Salless, and M. Reichel. Particulate
Pollutant System Study - Mass Emissions, Volumes 1, 2, and 3.
U.S. Environmental Protection Agency, (NTIS), Durham, North
Carolina. PB-203 128, PB-203 522, and PB-203 521. May 1971.
13. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation, Park Ridge, New Jersey. 1972.
48
-------
14. Goldeberg, A.J. (Dr.). A Survey of Emissions and Controls
for Hazardous and Other Pollutants. Environmental Protection
Agency, Washington, D.C. Publication Number EPA^R4-73-021.
February 1973.
15. Duncan, L.J. and E.L. Keitz. Hazardous Particulate Pollution
from Typical Operations in the Primary Nonferrous Smelting
Industry. Presented at the 67th Annual Meeting of the Air
Pollution Control Association. Denver, Colorado. June 9-13,
1974.
16. Personal Communication, A.D. Little Company.
49
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PRIMARY COPPER PRODUCTION PROCESS NO. 4
FlUidization Roasting
!• Function - The function of fluidization roasting is the same as for
multiple-hearth roasting (Process No. 3): to reduce the sulfur content
of the concentrate and to oxidize some of the iron, so that the matte
produced in the smelting process can be treated most efficiently in the
copper converter (Process No's. 9 and 10), Roasting also volatilizes
some impurities and preheats the reverberatory feed. The Noranda
process (Process No. 11) provides most of the functions of roasting,
smelting, and converting.
The fluidization roaster is the pyrometallurgical application of
the fluidized-bed principle that has revolutionized so many operations
in other industries over the past 30 years. It is based on the dis-
covery that if particles of a solid are added to a gas stream that is
moving vertically upward at just the right velocity, the solid takes on
many of the characteristics of an agitated liquid. Each particle of the
solid is in constant agitated motion, separated from all other particles,
and is in intimate contact with the gas stream. Any chemical reaction
that takes place between the solid and the gas happens very quickly,
with no cold pockets or hot spots.
In the fluidization roaster, the gas is a recycled stream of flue
gas, into which regulated streams of air and fuel gas are being in-
troduced. The solid is copper concentrate, continuously being fed and
continuously overflowing the fluidization vessel. Both the fuel and the
oxygen are completely consumed; by elimination of excess air, the S02
content of the flue gas stream is greatly increased, to the point that
the gas is strong enough for feed to a sulfuric acid plant.
If the sulfur content of the concentrate is high enough, fuel is
needed only at start-up. With 20 percent sulfur in the feed, sufficient
heat is released by the sulfur to make additional fuel unnecessary. The
multiple-hearth roaster requires 24 percent sulfur. Operators of fluidi-
zation roasters, therefore, find it best not to process the ore into
super-quality concentrates, but to tailor the quality of the concentrate
to match the requirements of the roaster.
Three domestic copper smelters have adopted fluidization roastingJ
Being complex and highly instrumented units, they must be capable of
large throughput to justify the investment. Units with capacities from
700 to 1500 metric tons per day are in use.2
50
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2. Input Materials - Copper concentrate is the only input, usually
blended or produced to a quality that the roaster can most economically
process.
3. Operating Conditions - Because of the thorough mixing in the fluid
bed, temperature is held constant in all portions of the bed in the
range of 650 to 750°C. The pressure in the bed is slightly above atmos-
pheric, and in portions of the recycle stream the pressure may be 1
kilogram per square centimeter or more.
4. Utilities - The major utility used is electricity to compress the
recycled gas, to inject air, and to operate auxiliary devices. No
estimates of power consumption are published.
Fuel gas is the usual source of heat for starting the charge,
although some plants have fuel oil facilities for standby. Quantity
consumed is small.
Oxygen enrichment facilities are being considered for some of these
units in order to provide more operating flexibility.
5. Waste Streams - Fluidization roasters are always fitted with cy-
clone separators that catch the large amounts of dust that rise from the
bed and return the dust to the bed. Dust quantities can be as much as
75 percent of the feed.2,3 These cyclones are most properly considered
as part of the process, and the waste stream considered as the outlet of
the cyclones. This waste stream contains particulates and fumes of the
same chemical character as these from a multiple-hearth roaster (Process
No. 3); they are rich in volatile elements such as arsenic and contain
considerable copper. Data on total quantities of dust are not avail-
able; they are likely to be greater than emissions from a multiple-
hearth roaster because of the more complete separation of the smaller
particles from the body of the ore.
Sulfur dioxide concentrations in the gas from the fluidization
roaster are reported to be from 12 to 16 percent.2.4,5
The roasting operation produces no liquid or solid wastes.
6. Control Technology - All the smelters that operate fluidization
roasters use the S02 for production of sulfuric acid. Except for other
processing to recover liquid SO? or elemental sulfur, this is the only
known technology with which to dispose of such a concentrated gas
stream.
51
-------
Almost complete removal of particulates Is required before the gas
is introduced into a sulfur recovery process. Electrostatic precipita-
tors and wet scrubbers are in use with the operating fluidization
roasters. Since the dusts and condensed fumes contain valuable materials,
they are normally returned to the pyrometallurgical processing units,
usually to the reverberatory furnace, but some may be discarded.
7. EPA Source Classification Code - 3-03-005-02
8. References -
1. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
2. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation, Park Ridge, New Jersey. 1972.
3. McAskill, D. Fluid Bed Roasting: A possible Cure for Copper
Smelter Emissions. Engineering and Mining Journal, p. 82-86.
July 1973.
4. Compilation and Analysis of Design and Operation Parameters
for Emission Control Studies. Pacific Environmental Services,
Inc. (Individual draft reports).
5. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research
Triangle Park, North Carolina, EPA-450/2-74-002a. October
1974.
52
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PRIMARY COPPER PRODUCTION PROCESS NO. 5
Drying
1. Function - Copper concentrates that are not to be processed by
roasting must first pass through a dryer, whose only function is to
decrease the moisture content. Drying may be practiced to simplify
handling of the concentrate; and in recent years it has been practiced
to make the concentrate suitable for direct feed to the reverberatory
furnace (Process No. 6). Excessive moisture in the concentrate may
cause minor eruptions or explosions that may damage the furnace.
Some smelters continue to use their existing multiple-hearth
roasters, operated at much lower temperatures, to accomplish this drying
operation. Special dryers, such as rotary kilns, are also being used.
The dryer may be more conveniently located at the concentrator rather
than at the smelter (see Process No. 2).
2. Input Materials - The ore containing 5 to 25 percent moisture is
the only input.Analysis is given in Table 2-11.
3. Operating Conditions - Except at flame fronts, temperatures in the
drying operation do not exceed 150°C. Pressures are atmospheric.
4. Utilities - Fuel gas is most frequently used for ore drying, al-
though facilities for substitution of oil are usually provided. One
report calculates that for a plant yielding 91,000 metric tons of refined
copper per year, the drying heat from fossil fuels would be equivalent
to 17,200 kg-cal (20 kWh) per hour of dryer operation.1
Electricity is used for conveying materials and for mechanical
operation of a dryer. The report cited above estimates 38,700 kg-cal
(45 kWh) per hour of operation for the same size plant.1
5. Waste Streams - Dust generated by a drying operation, would be of
the same composition as the input concentrate. A foreign plant using a
multiple-hearth roaster for drying reports particulate emissions of 0.05
percent of the weight of the feed. No data for domestic plants have
been reported.
Small quantities of organic materials in the concentrate could be
decomposed or volatilized in the drying operation, but no data have been
reported. Emissions of metallic fumes or oxides of sulfur are unlikely.
53
-------
6. Control Technology - Dust from a drying operation frequently con-
sists of the fine particles present in the ore, and their collection is
complicated by the ready condensation of moisture in the warm effluent.
If a dryer is installed at a concentrator plant, the best control is to
remove the dust by wet scrubbing and return it to the final stages of
the flotation process. If a multiple-hearth roaster is used for drying,
balloon flues or other particulate removal equipment may be modified to
handle this wet dust. Bag filters generally produce a caked product
that must be redried; they are effective, although troublesome, collec-
tors.
Quantities of the organic materials in the concentrate are believed
to be small enough that these materials require no separate treatment;
they are believed to be nontoxic.
7. EPA Source Classification Code - 3-03-005-06
8. References -
1. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-74/032. February 1975.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
4. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington, D.C.,
EPA-R2-73-274a. September 1973.
5. Jones, H.R. Pollution Control in the Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
6. Little, A.D. Economic Impact of New Source Performance
Standards on the Primary Copper Industry: An Assessment. U.S.
Environmental Protection Agency, Washington, D.C. C-75072-20.
October 1974.
7. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW, June
1969.
54
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PRIMARY COPPER PRODUCTION PROCESS NO. 6
Reverberatory Smelti ng
1. Function - Copper smelting is the process of removing from a roast-
ed or dried ore concentrate much of its iron and some undesirable im-
purities, leaving a molten mixture that can be processed efficiently by
a copper converter. This is most often accomplished with a reverbera-
tory furnace. It may, however, be accomplished by other methods (see
Processes Nos. 7, 8 and 11).
The reverberatory furnace is a large horizontal chamber into which
ore and various fluxing materials are charged. The furnace is then
heated by direct firing. As the temperature of the charge increases, a
complex series of chemical reactions takes place, and the charge separates
into three fractions. One fraction is S02 gas, which along with other
volatiles, is mixed with the combustion gases and is carried out of the
furnace. A second fraction is a molten slag containing much of the
iron, which is tapped from the furnace and discarded. The third frac-
tion, called the matte, becomes the charge to the copper converter.
The charge to the reverberatory furnace is proportioned so that the
resulting matte typically contains 40 to 45 percent copper and 25 to 30
percent each of iron and sulfur.1 The matte contains most of the heavy
elements present in the charge, practically all the gold and silver, and
part of the arsenic and antimony.
Reverberatory smelting, the oldest of the copper smelting processes
now in use, is little different now than when it was first practiced in
1879. It is in use at all but two of the smelters in this country, in
one of two modifications, described as either "deep bath" or "dry
hearth". Some reverberatory furnaces are very large, capable of accep-
ting as much as 1800 metric tons of material per charge.^
2. Input Materials - The primary input material is the roasted or
dried concentrate, whose composition is not much different from the
concentrate received from the mill. Slags from the converter and anode
furnace also are added for reprocessing, as are flue dusts from dust
collection equipment throughout the smelter. Precipitates from hydro-
metallurgical operations or materials from refinery processing may be
added at this step.
Flux consists normally of sand high in silica content, and usually
limestone to make the slag more fluid. Sometimes "direct smelting ore"
is used, which adds both fluxing material and additional copper.
55
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Composition of one charge to a reverberatory furnace in Arizona is
reported as follows:2
Ore concentrate - 64%
Converter slag - 2%
Hydro, precipitator - 1%
Flue dusts - 1%
Silica flux - 2%
Limestone flux - 6%
This charge produced molten materials of which 47 percent was matte
and 53 percent was slag.
3. Operating Conditions - When possible, the concentrate is charged
into the furnace while still hot from the roaster (400°C or more).
Converter slag is charged as a liquid (1100°C approximately). Other
materials are usually charged at ambient temperatures. The reverbera-
tory furnace usually heats the mixed charge to at least 1000°C before
the matte forms and separates; temperatures up to 1300°C have been
reported.3 All operations are at or near atmospheric pressure.
4. Utilities - It is estimated that 90 percent of the energy consumed
in a smelting operation is added into the reverberatory furnace.4 It
is reported that 18 billion kWh of energy were used in domestic copper
smelters in 1973.^ Consumption of energy by this process is very high;
it is usually supplied by natural gas, but pulverized coal or fuel oil
can be used. It is estimated that 500,000 kilocalories of heat is
required to smelt 1 metric ton of concentrate if the charge is preheated
by a roasting operation. If the charge is not preheated, an additional
390,000 kilocalories are required.5 These values give credit for steam
generated by waste heat boilers, which are almost always installed with
a reverberatory furnace. In itself, the reverberatory furnace is
thermally inefficient, using uses more than 4 times the heat theoreti-
cally required.6
Noncontact cooling water is used by copper smelters, primarily for
the protection of equipment auxiliary to the roaster, converter, and
reverberatory furnace. Data that allocate this cooling load specifi—
cally to each process are not available. Reported data indicate that
the total cooling water consumption for smelting operations can vary
from 4000 to 61,000 liters per metric ton of copper product.
Contact cooling water is used at four smelters to granulate the
slag from the reverberatory furnace. One smelter uses 1,700,000 liters
of water per day for this purpose.l
56
-------
5. Waste Streams - It is reported that 20 to 45 percent of the sulfur
that enters with the ore concentrate is emitted by the reverberatory
furnace as S02.3>7»8 Although most smelter operators have attempted to
make operational changes either to increase or reduce this quantity, no
recent data are available. The gas is released as a dilute stream of
variable composition, reported as from 0.5 to 2.5 weight percent
SO?.3'8'9^0*11^2 Other constituents in the exit gas are shown in
Table 2-16, for unroasted and roasted concentrate feeds. The volume of
this gas is very large, since it consists primarily of the combustion
gases from the heating fuels. Temperature of the exit gases may reach
1150 to 1200°C.10
Between 14 and 40 kilograms of particulate matter are emitted in
this gas stream per metric ton of copper matte produced.8,11,13,14 One
analysis of the particulates showed 24 percent copper and the following
concentrations of other elements:^
Zinc
Cadmium
Manganese
Chromium
Nickel
Mercury
mg/1
44,000
310
100
45
35
2.5
Other investigations indicate that most of the volatilized arsenic,
selenium, lead, antimony, cadmium, chromium, and zinc emissions will be
generated in the reverberatory furnace.lOJ'»14,16,17,18
Fugitive dust is generated in this process as materials are loaded
into the furnace. No quantities are reported, but this is probably one
of the largest sources of dust in a smelting operation.
The only liquid waste from this process is the run-off from slag
granulation. Three complete analyses are reported as shown in Table
2-17. Liquid waste is most often generated as the overflow from a pond
into which the molten slag is dumped. Since the pond is an open body of
usually hot water, subject to rainfall and evaporation, quantity and
composition of the overflow may be highly variable.
One copper smelter is situated close to a market for the furnace
slag it produces; for all the others, the slag constitutes a large
quantity of solid waste, as much as 3000 kilograms per metric ton of
copper produced.15 Table 2-18 gives an analysis of potentially hazardous
elements found in a reverberatory furnace slag. The bulk of the slag is
a mixture of iron silicates, with other elements, also as shown in
Table 2-18.15
57
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Table 2-16. COMPOSITION OF REVERBERATORY FURNACE
EXHAUST GASES12
Component
Carbon dioxide
Nitrogen
Oxygen
Water
Sulfur dioxide
Green feed,
%
8.4
69.3
0.25 - 1.0
18.8
1.5 - 2.5
Calcined feed,
%
10.2
71.0
0.25 - 1.0
17.7
.6
58
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Table 2-17. EFFLUENTS FROM SLAG GRANULATION
(mg/1)
16
Parameter
PH
TDS
TSS
so4=
CN-
As
Cd
Cu
Fe
Pb
Hg
Ni
Se
Te
Zn
Oil and grease
Plant 103
7.7
140.
6.8
62.
0.005
3.7
0.001
0.12
0.04
0.04
0.0001
0.001
0.001
0.001
0.44
Plant 110
8.1
3800.
151.
310.
0.050
0.048
0.001
0.05
0.03
0.070
0.0001
0.06
0.54
0.023
0.0
Plant 102
6.4-7.6
0.030
5.70
0.042
0.604
340.
7.4
0.0001
0.16
0.040
0.100
36.
0.02
59
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Table 2-18. GENERAL RANGE OF REVERBERATORY FURNACE
SLAG COMPOSITION
Compound
or element
FeO
Si02
CaO
MgO
A1203
Copper
Sulfur
Composition
percent
34 to 40
35 to 40
3 to 7
0.5 to 3
4.5 to 10
0.4 to 0.7
1.0 to 1.5
Trace elements
Zinc
Magnanese
Antimony
Lead
Chromium
Selenium
Nickel
Cadmium
Mercury
Arsenic
Tellurium
Cobalt
Parts per million
Approximately 7800
Approximately 450
Approximately 400
Approximately 100
Approximately 100
Approximately 20
Approximately 25
Approximately 10
Less than 1.0
Trace
Trace
Trace
60
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6. Control Technology - Gases from the reverberatory furnace pass
through a waste heat boiler and then through an electrostatic preci-
pitator for particulate removal. The gases may pass through spray
coolers or balloon flues before entering the ESP units. The degree of
particulate removal ranges from 50 to 99.9 percent. Particulates col-
lected are recycled into the metallurgical process, normally as part of
the reverberatory furnace charge, but accumulations of trace elements
causes some flue dusts to be discharged or processed separately.
Quantities and their disposition are not reported.
At present, there is virtually no control of the $03 emissions from
reverberatory smelters. Intensive studies are underway to develop
scrubbing techniques that can be applied to large volumes of flue gas
containing small concentrations of $03. These represent the best
available control technology. One smelter absorbs the S02 from this
stream in dimethyl aniline and regenerates it as a concentrated stream
for further processing. One Canadian smelter uses an ammonia absorption
process on some smelter streams, but this system is not in use in this
country. Other scrubbing solutions, containing compounds of zinc and
aluminum, are used on smelter gases in Japan. Scrubbers using lime or
limestone, with and without magnesium addition, are being used on sulfur-
eontaining flue gases from coal-fired boilers in the United States, and
might be adopted for use in U.S. smelters, as has been done in Japan.
Another absorption process based on sodium sulfite-bisulfite is under
test. The only one of these processes specific to the domestic copper
industry is DMA absorption, described in Process No. 14.
Of the four smelters that practice slag granulation, one reports no
wastewater from this source, since the rate of evaporation at this
location necessitates a continuous water make-up to the quenching pond.
The other three smelters mix the water from slag granulation with other
wastes.'*19 Control of these mixed wastes is discussed in Section 6 of
this report.
Granulated slag is usually a coarse-grained material of low to
medium density, usually discarded near the smelter. A small amount may
find a market for use as road fill or concrete aggregate. Crushed slag
that has not been granulated also finds a small market for these same
purposes. Most slag is not granulated, but is simply poured out and
allowed to solidify. There is no easy way to naturalize or reclaim the
slag dumping areas, and there are no published reports on how this could
be done. It is generally assumed that the potential of secondary water
pollution from slag dumps is less than that from mine spoil or tailings
beds.
7. EPA Source Classification Code - 3-03-005-03
61
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8. References -
1. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
4. Fejer, M.E. and D.H. Larson. Study of Industrial Uses of
Energy Relative to Environmental Effects. U.S. Environmental
Protection Agency. Research Triangle Park, North Carolina.
July 1974.
5. Metallurgy Processing in 1974, Mining Congress Journal.
February 1975.
6. Treilhard, D.G. "Copper - State of the Art", Chemical Engineering
Journal. April 1975.
7. Halley, J.H. and B.E. McNay. Current Smelting Systems and
Their Relation to Air Pollution. Arthur G. McKee and Company,
San Francisco, California 35224. September 1970.
8. Jones, H.R. Pollution Control in the Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
9. Compilation and Analysis of Design and Operation Parameters
for Emission Control Studies. Pacific Environmental Services,
Inc. (Individual draft reports).
10. Measurement of Sulfur Dioxide, Particulate and Trace Elements
in Copper Smelter Converter and Roaster/Reverberatory Gas
Streams. Environmental Protection Agency, Washington, D.C.
EPA 650/2-74-111. October 1974.
11. Statnick, R.M. Measurement of Sulfur Dioxide, Particulate and
Trace Elements in Copper Smelter, Converter and Roaster/
Reverberatory Gas Streams. U.S. Environmental Protection
Agency, (NTIS), Research Triangle Park, North Carolina.
PB-238095. October 1974.
62
-------
12. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW. June
1969.
13. Vandegrift, A.E. (Dr.), L.O. Shannon (Dr.), P.G. Gorman, E.W.
Lawless, (Dr.), E.E. Sal lee, and M. Reichel. Particulate
Pollutant System Study - Mass Emissions, Volumes 1, 2, and 3.
U.S. Environmental Protection Agency, (NTIS), Durham, North
Carolina. PB-203 128, PB-203 522, and PB-203 521. May 1971.
14. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo- Environmental Specialists, Inc. August 1974.
15. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II Primay and Secondary Non-
ferrous Smelting and Refining. Calspan Corporation, Draft,
April 1975.
16. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b, February 1975.
17. Davis, W.E. National Inventory of Sources and Emissions:
Copper, Selenium, and Zinc. U.S. Environmental Protection
Agency, (NTIS), Research Triangle Park, North Carolina.
PB-210 679, PB-210 678, and PB-210 677. May 1972.
18. Phillips, A.J. The World's Most Complex Metallurgy (Copper,
Lead, and Zinc). Transactions of the Metallurgical Society
of AIME. Volume 224: pp. 657-668. August 1962.
19. Assessment of the Adequacy of Pollution Control Technology for
Energy Conserving Manufacturing Process Options. Industry
Assessment Report on the Primary Copper Industry. Arthur D.
Little, Inc. Draft. October 1974.
63
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PRIMARY COPPER PRODUCTION PROCESS NO. 7
Electric Smelting
1. Function - Copper smelting is the process of removing from an ore
concentrate a portion of its iron and sulfur content in order to produce
a molten mixture that can be treated efficiently in subsequent pro-
cessing. Smelting requires only the application of heat to produce
matte and slag from a charge of minerals. Three other smelting methods
are used (see Processes Nos. 6, 8, and 11). In electric furnace smel-
ting, heat is supplied by electricity.
Electric furnaces for copper smelting are similar to those used in
other high-energy industries. Two electric furnace installations in the
United States are used for copper smelting. Capacities are up to
1350 metric tons of total charge per day.
2. Input Materials - The principal input is copper ore concentrates
processed or blended to a suitable composition. Various fluxing mate-
rials are also required. The charge materials are similar to those
outlined for a reverberatory furnace (Process No. 6).
In addition, electric smelting requires the use of carbon elec-
trodes to conduct electric current into the layer of slag. These elec-
trodes are consumed during operation. Various types of carbon elec-
trodes can be used. In one U.S. smelter a proprietary paste carbon
mixture, is consumed at a rate of 2.5 kilograms per metric ton of
charge.2
3. Operating Conditions - The charge is usually heated to temperatures
between 1000° and 1300°C. and the electric furnace is operated at a
small negative pressure.' Electric furnaces are normally enclosed in a
large building.
4. Utilities - When hot concentrate is fed from a roasting process,
electric energy is consumed at a rate of 520,000 kg-cal (605 kWh) per
metric ton of total feed.2 Use of cold feed requires 850,000 kg-cal
(990 kWh) per metric ton of concentrate.3 No direct combustion of fuel
takes place in electric smelting.
The furnace must be cooled to protect some of the components from
high temperatures. Cooling is done partially by infiltration of air
into the furnace, but some external cooling is also required. Either
water or air can be used, the quantity depending on furnace design.
Operators of one electric smelter report that 111,000 liters per minute
of air is circulated each second to cool a furnace of 51,000 KVA trans-
former capacity.4
64
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5. Waste Streams - S02 in the effluent gas from an electric furnace is
more highly concentrated and temperatures are lower than in effluent
from a reverberatory furnace. S02 concentration can range from 3 to 8
percent.3 Within limits this can be adjusted to make the gas suitable
for sulfuric acid production. The electric furnace may produce small
amounts of hydrogen gas and carbon monoxide, but sufficient air is
allowed to infiltrate to keep these combustible materials oxidized.
Very small amounts of hydrocarbons released from the electrode compounds
will also burn.
Because of the lower gas volumes, and more uniform gas flow, emis-
sions of particulate matter would be expected to be lower than with a
reverberatory furnace; no published estimates are available. Particu-
late composition would be about the same as from a reverberatory furnace
(Process No. 6).
Slags and wastewaters from slag granulation would be similar to
those of the reverberatory furnace, although more complete removal of
copper and sulfur compounds from electric furnace slags is likely.
6. Control Technology - The two operating electric smelters in this
country use the gases from the furnaces for sulfuric acid manufacture.
One installation combines the gas with the effluent of a converter, and
the other with the effluent of a fluidization roaster. Acid manufacture
is the best available technology for $03 removal from electric furnaces,
since the sulfur content is high enough for that application.
Control of the slag as a solid waste, or lack of controls, is
described in reference to the reverberatory furnace (Process No. 6).
The small wastewater stream from slag granulation is invariably
mixed with other streams for treatment, as described in the section on
sulfide ore mining and concentrating.
7. EPA Source Classification Code - 3-03-005-03
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington, D.C.,
EPA-R2-73-274a. September 1973.
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
65
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4. Cole, R.C. Inspirations' Copper Smelter Facilities. Mining
Congress Journal. October 1974. p. 22-32.
66
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PRIMARY COPPER PRODUCTION PROCESS NO. 8
Flash Smelting
1. Function - Copper smelting is the process of removing from an ore
concentrate a portion of its iron and sulfur content, in order to pro-
duce a molten mixture that can be treated efficiently by a copper con-
verter. Until about 25 years ago, only two batch-type smelting pro-
cesses, were available, both inefficient in energy consumption (see
Processes Nos. 6 and 7). Since that time, a continuous flash smelting
process has been developed. This process performs the smelting function
at a much higher thermal efficiency while producing a continuous stream
of flue gas, much easier to control, with a high S02 content.' The
Noranda process (Process No. IT) is an alternate method which combines
most of the functions of roasting, smelting, and converting.
In flash smelting, ore concentrates are injected along with flux
and preheated air into a combustion chamber. Part of the sulfur of the
concentrate burns in a "flash" combustion while the particles are
falling through the chamber. The heat from this combustion maintains
smelting temperature. Matte and slag form in the chamber and separate
into layers. The matte is sent to a conventional converter for further
processing, and the slag, which contains too much copper to discard, is
also further processed (see Process No. 12).
One smelter in the United States is operating a flash smelting unit
that was developed in Finland. This version is in extensive use in
several other countries. Another flash smelter design, using pure
oxygen, is operating in Canada.
2. Input Materials - Copper concentrates especially tailored for flash
smelting are the primary input. Not all concentrates are suitable for
this process. The concentrates must be finely pulverized (50 percent
minus 200 mesh),2 and must contain very low concentrations of lead,
zinc, and other volatile metals. They must have a fairly high sulfur-
to-copper ratio, and thus are not high-grade concentrates. They are not
preroasted, unless they contain considerable arsenic, but must be dried.
Precipitates from hydrometallurgical operations cannot normally be
handled by a flash smelter.
Flux in the form of silica sand or crushed rock must be prepared in
a separate milling process to 50 percent through 140 mesh2 and must
also be dried. High grade "direct smelting" ores can be substituted if
available.
3. Operating Conditions - Temperature in the flash chamber is main-
tained at approximately 1100°CJ Pressure is approximately atmospheric.
67
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4. Utilities^ - Fuel consumption in the Finnish version of flash
smelting is only about two-thirds of that required by a reverberatory
furnace in equivalent product!on.2,3 Except for start-up or abnormal
operations, fuel is required only to preheat the combustion air. This
is reported as 7,560,000 kg-cal (30 million Btu) per hour for copper
production of 100,000 metric tons per year.2 Any fossil fuel can be
used.
Natural gas or oil is used to heat the furnace to start-up tempera-
ture. Oil can also be used if required to maintain smelting tempera-
tures with some concentrates.
Electrical power is used to operate feed and air injection equip-
ment as well as the complex instrumentation this process requires. An
estimate of 516,000 kg-cal (600 kWh) per hour of operation of a 91,000
metric ton per year copper plant production has been reported.4
5. Waste Streams - Flash smelting removes a large percentage of the
sulfur from the concentrate. From 50 to 80 percent is converted to S02,
which leaves as a 10 to 20 percent stream. In Finland the concentrates
are blended to produce a 14 percent $03 concentration, which is ideal
for a suitably-designed sulfuric acid plant.5
Particulates in the gas stream are expected to equal about 6 to 7
percent of the feed, which is about the same as in the reverberatory
furnace. Composition should be about the same as that of effluent from
a reverberatory furnace, except that content of volatile metals should
be lower since they are lower in the feedstock. Care is taken to keep
zinc and lead to a minimum in the concentrates, since they tend to plate
out within the flash chamber.
There are no solid or liquid wastes from flash smelting. The slag
is discharged to waste from the slag treatment process (Process No. 12).
6. Control Technology - An important objective in development of the
flash smelting process was sale of the sulfur. There was a good market
for sulfuric acid near the Finnish smelter;^ continuous and stable
production of S02 made acid production most efficient. Flash smelting
therefore was not developed with production of copper as the sole
consideration.
It is possible, for reasons of energy economy, that flash smelters
will be built in this country where there is no local market for sulfur
compounds. The best currently available technology for control of S02
emissions would still be sulfuric acid production, even if the acid were
then neutralized and discarded, wet scrubbing would be an expensive,
although satisfactory, alternative. The flash smelter in this country
uses the gas for acid manufacture.
68
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Complete removal of participates is required for sulfuric acid
manufacture, and recovered dusts would be blended back into the flash
smelter feed. Cyclones, balloon flues, electrostatic precipitators, and
wet scrubbers'afford satisfacotry methods for removal.
7. EPA Source Classification Code - 3-03-005-03
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
3. Metallurgy Processing in 1974, Mining Congress Journal,
February 1975.
4. Personnal Communication with Mr. Paterson. El ken - Spigerverket
a/s, New York, New York.
5. Treilhard, D.G. Copper-State of the Art. Chemical Engineering
and Mining Journal. 1975.
69
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PRIMARY COPPER PRODUCTION PROCESS NO. 9
Peifce-Smith Converting
1. Function - A copper converter produces crude blister copper metal
from the matte that is formed in the smelting process. The Peirce-Smith
converter is the type most often used by copper producers in this coun-
try. See also Process No. 10 for an alternate converting method. In
addition, the Noranda process (Process No. 11) performs most of the
functions of roasting, smelting and converting.
Matte is a molten mixture containing copper, iron, and sulfur. In
the converter, a flux is added to the matte and compressed air is blown
into the mixture through a series of openings called tuyeres. The
remaining sulfur is oxidized to S02 and leaves with the flue gases. The
iron forms into a slag that is returned to the smelting process. The
blister copper is removed for further processing.
2. Input Materials - Matte from the smelting process is the prinicipal
input. Scrap copper and brass being recycled are also introduced at
this step, as is scrap produced within the smelter from spills or ladle
"skulls", and materials from other processes with high concentrations of
metallic copper. Flux used in the converter is sand or crushed rock
with a high silica content. Sulfur is added if necessary to maintain
the proper ratio of copper to sulfur.
Table 2-19 shows an average charge and product distribution from
one copper converter.
3. Operating Conditions - To ensure that a slag of proper composition
is formed and separated from the molten copper, converter temperatures
are carefully controlled at 1175° to 1200°cJ>2 Tne converter operates
at atmospheric pressure.
4. Utilities - The converting process consumes no fuel, since oxida-
tion of the remaining sulfur furnishes enough heat to keep the mixture
at the proper temperature.
Electricity is used to rotate the converter to discharge the slag
and product.
Compressed air is required for the process. No data are reported
on the required quantities.
A small amount of cooling water is used for noncontact cooling of
some of the converter sections and auxiliaries.
70
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Table 2-19. MATERIAL BALANCE ON CONVERTERS
SMELTER IN ARIZONA1
percent
Material
Input
Output
Reverberatory matte
Silica
Scrap and brass, etc.
Reverts
Sulfur
Blister copper
Slag
Sulfur Dioxide
Flue dust
78
13
4
4
0.5
28
67
2
3
71
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Table 2-20. COMPOSITION OF CONVERTER DUSTy
Component
Cu
Fe
Pb
Bi
F
Sb
As
Se
Si
Mg
Mo
Al
°2
Cl
Te
S
Ca
Percent, weight
10 - 19.0
10 - 20.0
0.83 - 2.5
0.61
nil
nil
0.04 - 0.6
0.03 - 0.5
5.0 - 15.0
0.57
0.08
0.4 - 3.60
21.0
nil
0.005 - 0.01
12.0
1.0
72
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Table 2-21. PARTICLE SIZE DISTRIBUTION IN CONVERTER DUST
7
Mesh
-48 to + 65
-65 to + 100
-100 to + 150
-150 to + 200
-200 to +270
-270 to + 335
-325 to + 32 microns
-32 microns to + 25 microns
-25 microns
Percent
0.5
1.5
2.5
3.5
5.0
5.0
16.6
55.6
9.8
Table 2-22. PARTICULATE EMISSIONS ANALYSIS AT STACK OUTLET FOR
REVERBERATORY FURNACE AND CONVERTER3
Metal
Arsenic
Cadmium
Copper
Selenium
Zinc
Chromium
Manganese
Nickel
Vanadium
Boron
Barium
Mercury
Lead
Total
% .
0.038
0.008
5.6
0.014
1.1
0.006
0.023
0.0045
0.0023
0.12
0.03
0.0007
0,065
7.0115
Stack test data (5/13/71).
73
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Table 2-23. CONVERTER OFF-GAS COMPOSITION
7,11
Component
N2
°2
so2
so3
Dust
Percent by volume
80
11
6.9 - 9.8
0.05 - 0.07
0.0053 g/lit max.
74
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5. Waste Streams - The converter emits about 120 kilograms of particu-
late matter per ton of copper produced.3,4,5,6 Tables 2-20 and 2-21
provide data on converter dust from some Arizona smelters, and Table
2-22 gives an analysis of particulates from a smelter in Nevada.
Peirce-Smith converters are designed to be partially covered by a
hood that catches the S02 and particulate emissions from the converter,
but also draws in a considerable amount of uncontaminated air. By
stoichiometric calculations, the S02 content in converter gases varies
from 15 to 20 percent during various stages in the processing of a
charge, but when diluted by the excess air the resulting mixture con-
tains from 4 to 10 percent S02- »4,5,7,8,9,10 $OITie converters may
produce gas with S02 content as low as 2 percent.7.10 Because the mouth
of the Peirce-Smith converter is rotated from under the hood when flux
is added and when slag and copper are poured out, local losses of S02
occur during those periods.
Table 2-23 gives the composition of converter off-gas from an
Arizona smelter.
Fugitive dust and fumes are generated in a converting operation in
the area near the converter. No quantities have been reported.
There are no solid or liquid wastes from the converter process.
6. Control Technology - Gases from a Peirce-Smith converter are some-
times combined with the gas stream from a smelting or roasting process
for particle removal and further treatment. The converter S02 stream is
usually controlled through sulfuric acid plants.
Technology for control of a mixed gas stream including converter
off-gas is discussed with the various smelting and roasting processes
(Process Nos. 3, 4, 6, 7, and 8).
7. EPA Source Classification Code - 3-03-005-04
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
75
-------
3. Vandegrift, A.E. (Dr.), Shannon, L.J. (Dr.), Gorman, P.G.,
Lawless, E.W. (Dr.), Sallee, E.E. and Reichel, M. Particulate
Pollutant System Study - Mass Emissions, Volumes 1, 2 and 3.
U.S. Environmental Protection Agency, (NTIS), Durham, North
Carolina. PB-203 128, PB-203 522 and PB-203 521. May 1971.
pp. 500.
4. High, M.D. and Lukey, M.E. Exhaust Gases from Combustion and
Industrial Processes. U.S. Environmental Protection Agency,
(NTIS), Durham, North Carolina PB-204 861. October 1971.
5. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
6. Statnick, R.M. Measurement of Sulfur Dioxide, Particulate and
Trace Elements in Copper Smelter, Converter and Roaster/Rever-
beratory Gas Streams. U.S. Environmental Protection Agency,
(NTIS), Research Triangle Park, North Carolina. PB-238095.
October 1974.
7. Compilation and Analysis of Design and Operation Parameters
for Emission Control Studies. Pacific Environmental Services,
Inc. (Individual draft reports). November 1975.
8. Control of Sulfur Oxide Emissions in Copper, Lead, and Zinc
Smelting. Bureau of Mines Information Circular 8527, 1971.
9. Halley, J.H. and McNay, B.E. Current Smelting Systems and
Their Relation to Air Pollution. Arthur G. McKee and Company,
San Francisco, California 35224. September 1970.
10. Jones, H.R. Pollution Control in the Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
11. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. Department
of HE&W. June 1969.
76
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PRIMARY COPPER PRODUCTION PROCESS NO. 10
Hobdken Converting
1. Function - The Hoboken converter is one type of equipment used to
produce crude blister copper from the matte formed in the smelting
process. Its function is identical to that of the Peirce-Smith converter
(Process No. 9). The principal difference is that the flue that removes
the gas from the converter is an integral part of the converter construc-
tion instead of being a hood mounted above it. This design minimizes
infiltration of uncontaminated air, minimizes local losses of $02 from
the converter mouth, and allows production of a gas with a higher and
more uniform SO? content. The Noranda process (Process No. 11} also
provides most or the functions of converting, as well as roasting and
smelting.
One smelter in this country uses Hoboken converters.
2. Input Materials - These are the same as for Peirce-Smith converters,
consisting largely of matte from the smelters, plus silica flux.
3. Operating Conditions - These are also the same as for a Peirce-
SmithuhTrrT200^C~lrE~atmbspheric pressure.
4- Utilities - Same as for Peirce-Smith. No supplemental fuel is
required for the converting process.
5. Waste Streams - There are no published reports of the quantity or
nature of particulates from a Hoboken converter. Since particulate
matter is generated by the air being blown through the converter charge,
particulate emissions should be comparable to those of the Peirce-Smith.
The S02 content of the gas stream from this converter is at least 8
percent if three or more converters are operating, and may reach as high
as 13 percent.1 It is calculated that with oxygen enrichment, the SO?
concentration could be increased to 10 to 14 percentJ
6. Control Technology - Production of S02 by the Hoboken converter is
intermittent, but a battery of several of the converters will produce a
stream sufficiently constant in rate to allow the gas to be used for
sulfuric acid manufacture. The one domestic smelter using Hoboken
equipment mixes the gas stream with the gas from an electric smelter,
and after particulate removal, uses the gas for this purpose.
77
-------
7. EPA Source Classification Code - 3-03-005-04
8. References -
1. Cole, R.C. Inspirations' Copper Smelter Facilities. Mining
Congress Journal. October 1974. pp. 22-32.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
78
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PRIMARY COPPER PRODUCTION PROCESS NO. 11
Nbraiida
1. Function - Noranda is one design of a continuous smelter, which in
a single furnace combines most of the functions of roasting, smelting,
and converting. This process approaches a one-step method of producing
copper metal from ore concentrates. A Noranda installation is being
constructed in Utah. One other unit is operated by the developer in
Canada.
The Noranda furnace is a horizontal cylinder, into which a mixture
of concentrate and flux is continuously fed, along with fuel and oxygen.
The mixture reacts to form copper matte and slag, which separate into
layers. Additional oxygen-enriched air is blown through side-mounted
tuyeres into the matte layer, forming blister copper, which collects in
a third liquid layer below the matte. Slag and copper are intermittently
tapped from the furnace. Noranda slag contains 10 to 12 percent copper,
and is processed to recover the copper content (see Process No. 12).
Noranda does not completely eliminate the use of the copper con-
verter. Blister copper from Noranda contains from 1.5 to 2.0 percent
sulfur, and is usually batch treated in a standard converter to remove
additional sulfur prior to fire refining. If the concentrate contains
considerable impurity elements, the developer recommends that Noranda be
used as a smelter only, to produce a high-grade matte for separate
conversion to blister copper. A converter is also used to process matte
from Noranda slag treatment.
2. Input Materials - As normally used, Noranda will be fed with
"clean" concentrates that contain low concentrations of volatile ele-
ments, especially arsenic. Pulverized silica and limestone fluxing
materials are blended with the concentrate. Flue dusts may be mixed
with the charge. Many of the finer particles may be pelletized to
minimize particulates in the gas stream.
3. Operating Conditions - This process operates at approximately
atmospheric pressure.Temperatures in the U.S. installation have not
been reported, but will probably be higher than 1100°C, which is the
flue gas temperature from the Canadian smelter.
4. Utilities - A principal advantage of Noranda is its efficient
utilization of fuel. With oxygen enrichment, about 440,000 kilocalories
of heat are required to produce a metric ton of copper, which is about
22 percent of that required by a reverberatory furnace. The operating
installation in Canada uses fuel oil as a heat source, but other fuels
are acceptable.
79
-------
The fuel consumption reported above was based on enrichment of
combustion air to 50 percent oxygen, and of tuyere air to 35 percent.
Although oxygen enrichment is not necessary with the Noranda process,
best economy requires its use, and it has been reported that the U.S.
installation will use a higher degree of enrichment than in the Canadian
plant.
Small amounts of electricity are used for feed injection and
auxiliary services.
5- Waste Streams - With oxygen enrichment, gas from the furnace con-
tains about 23 percent S02- Air infiltration around the hood that
encloses the exhaust duct reduces the concentration to about 13 percent
S02. If oxygen enrichment is not used, furnace flue gases will contain
about 4 percent S02- Gas emission is interrupted about 5 percent of the
time during tapping operations.
Particulate emission rates have not been reported, but are probably
dependent on the size distribution in the feed. Since feed is con-
tinuously injected at high velocity into a moving gas stream, particu-
late loadings could be substantial.
The use of Noranda will cause an emission of a gas stream low in
S02 content from the associated converter operation. Details are
unreported.
There are no solid or liquid wastes from this process. Slag is
transferred to Process No. 12 for further treatment.
6. Control Technology - At the operating Canadian smelter, most of the
particulates are collected in a cooling chamber connected to the furnace
hood, and in an ESP unit. Most of the dust is pelletized and recycled.
The gas is further cleaned and used for sulfuric acid production. This
probably represents best control technology for this process. The
equipment to be used at the U.S. plant has not been reported.
Control technology for Noranda or converter off-gas cannot be
established until data of gas composition are known.
7. EPA Source Classification Code - None
8. References -
1. Environmental Considerations of Selected Energy Conserving
Manufacturing Process Options, Volume XIV, EPA 68-03-2198.
2. Advertising literature and letter, Noranda Mines Limited,
Toranto.
80
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PRIMARY COPPER PRODUCTION PROCESS NO. 12
Slag Treatment
1. Function - Slag from a flash or continuous smelter contains too
much copper to discard economically. Also, in flash or continuous
smelting there is no way to recycle the slag formed in the converters
and the anode furnaces. Among the various ways to reclaim the copper
content of these slags is use of an electric furnace. This is the
procedure to be used with the flash smelter now operating in the United
States, and the continuous smelter soon to be built.
In electric furnace slag treatment, coke is used to reduce sulfates
and metallic copper and to reconstitute the copper as a sulfide. A
molten matte is formed that can be recycled to a converter for produc-
tion of copper metal; the process leaves a slag low in copper content
that can be discarded.
2. Input Materials - The slags are similar to those from the rever-
beratory furnace (Process No. 6), the copper converters (Processes No. 9
and 10), and the fire refining furnaces (Process No. 17), except with
higher copper content. Flash smelting slags contain 1 to 2 percent
copper, and slags from Noranda, 10 to 12 percent copper.
Carbon electrodes, as described for electric smelting (Process No.
7) are consumed. Reported usage is 1.5 kilograms per metric ton of slag
processed.'
Iron pyrites are usually added to the furnace charge to adjust
sulfur content.
The coke is similar to that used in electric furnace operations in
other industries. High grade coal can be substituted. No reports of
range of addition are available.
3. Operating Conditions - Temperatures are maintained somewhere in the
range of 12005 to 1300°C.i Pressures are approximately atmospheric.
4- Utilities - Electric consumption is reported as 190,000 kg-cal
(220 kWh) per metric ton of slag treated.2 There is no reported use of
cooling water or air in the slag treatment furnace.
5- Waste Streams - Since this is a reducing furnace, it is expected
that S02 in the exit gases is negligible. Carbon monoxide and parti-
culates are present however, including metallic fumes of zinc and other
elements, and there will be some hydrogen if moisture is introduced into
the furnace along with the coke. There are no reported analyses of
these gases. Gas volumes are relatively small.
81
-------
Slag discharged from this treatment is primarily iron silicate,
similar to the slag from the reverberatory furnace (Process No. 6). No
analyses are available.
No liquids are discharged from this process.
6. Control Technology - In other industries, gases from reducing
electric furnaces are passed through a wet scrubber for cooling and
particulate removal, and are then either burned for fuel, incinerated
with no recovery of the heat, or discharged through a stack. Venturi
scrubbers are normally used to cool the gas quickly and minimize the
possibility of explosions. This is a preferred design, but combustion
of hot gases prior to particulate removal is sometimes practiced, which
may be cooled with water sprays and passed through an ESP for particu-
1 ate removal.
Particulates will probably be sluiced into a tailings pond and
discarded, since they should be low in volume. They may contain quanti-
ties of trace elements, however, and their proper disposal warrants
further study.
Slags from the slag treatment furnace are discarded, with or with-
out granulation, as outlined for the reverberatory furnace (Process No.
6).
7. EPA Source Classification Code - None
8. References
1. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
2. Personal Communication with Mr. Paterson El ken - Spi\erverket
a/s. New York, New York.
82
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PRIMARY COPPER PRODUCTION PROCESS NO. 13
Contact Sulfuric Acid Plant
1. Function - An acid plant catalytically oxidizes S02 gas to sulfur
trioxide, and absorbs it in water to form sulfuric acid. S02 gas may be
controlled by DMA absorption (Process No. 14) or elemental sulfur pro-
duction (Process No. 15).
Contact sulfuric acid plants are continuous steady-state processing
units that are operated in other industries using S02 made by burning
elemental sulfur. They may be used with waste S02 streams if the gas is
sufficiently concentrated, is supplied at a reasonably uniform rate, and
is free from impurities.
The heart of a sulfuric acid plant is a fixed bed of vanadium
pentoxide or other special catalyst. All other components of the plant
are auxiliary to this catalytic converter, which oxidizes the S02. The
other components clean and dry the stream of gas, mix the proper amount
of oxygen with it (unless sufficient oxygen is present), preheat the gas
to reaction temperature, and remove the heat produced by the oxidation.
The plant incorporates one or two absorbers to contact the gas with
water to form the acid. If only one absorber is provided, this is
described as a single-contact sulfuric acid plant. If two are provided,
the second is placed between stages of the converter, and this is a
double-contact plant. The second absorber allows a larger proportion of
the SOz to be converted into acid, and thus removes more S02 from the
gas stream if the initial concentration is high.
Thirteen of the copper smelters in this country operate contact
sulfuric acid plants to treat all or part of the gases from the metal-
lurgical operations.
2. Input Materials - Most contact sulfuric acid plants operate most
efficiently with a constant gas stream that contains 12 to 15 percent
S02- Performance almost as good can be achieved in plants that are
designed for 7 to 10 percent S02 content. The ability of a plant to
convert .most of the S02 to sulfuric acid declines either as gas streams
become weaker in $03 or as the flow rate or concentration become less
consistent. A concentration lower than 4 percent $02' is extremely
inefficient, since sufficient catalyst temperature cannot be maintained.
Certain modifications of the process, which add heat by combustion of
fuel, can provide better conversion at low S02 concentrations.
The gas that enters the catalyst bed must be cleaned of all partic-
ulate matter, must be almost completely dried, and must contain no gases
or fumes that act as poisons to the catalyst. The acid plant is always
supplied with special scrubbers to remove final traces of objectionable
83
-------
materials. Table 2-24 provides information on the acceptable limits of
these impurities.
Clean water is required to react with the SO^ to form sulfuric
acid. It may be necessary to deionize the water in a special ion
exchange system in order to avoid excessive corrosion or to meet acid
quality specifications. Steam condensate may also be used.
3. Operating Conditions - The catalyst bed operates properly only if
temperatures are maintained between 450° and 475°C. Pressures do not
usually exceed 2 kg/cnr. The plants are usually not enclosed in a
building.
4. Utilities - Noncontact cooling water is required. At one plant
producing 1500 metric tons of acid per day, about 12 million liters of
water are required each day.2
A small amount of electricity is required for pumps and blowers.
In certain patented modifications, heat from combustion of natural
gas is used to provide better efficiency at low SO;? concentrations.
Natural gas or oil is also required to heat any acid plant to operating
temperature following a shutdown.
5. Waste Streams - Single-contact sulfuric acid plants using weak gas
streams can at best absorb only 96 to 98 percent of the S02 fed to them.
The remaining quantity passes through to the atmosphere. Efficiencies
as low as 60 percent have been reported.^ In addition, it is likely
that some S02 may be vented without treatment in some smelters, since an
acid plant cannot instantly change the flow to match the intermittent
production typical in the copper industry. Of gas that is treated, it
is reported that absorber exit gases may contain from 0.01 to 0.5 per-
cent S02-4 Total flow rates may range from 34,000 to 68,000 std. cu.
meters per hour.5
Double-contact acid plants provide a higher percentage of S02
removal, if they are fed gas with a higher S02 content. Efficiencies
higher than 99 percent have been reported. Exit gas S02 concentration
is usually still within the same range as shown above.
In sulfuric acid plants, it is difficult to prevent some loss of
$03, in the form of a fine mist of sulfuric acid, with the absorber exit
gases. This is usually 0.02 to 0.04 kilogram of $03 per metric ton of
100 percent acid produced.
The scrubbing columns that clean the waste gas stream create off-
grade weak acid that cannot be sold. The amount is estimated as 4 to 8
liters for each 10 cubic meters of gas treated.6 Table 2-25 provides
typical analyses for acid plant blowdown.
84
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Table 2-24. ESTIMATED MAXIMUM IMPURITY LIMITS FOR METALLURGICAL
OFF-GASES USED TO MANUFACTURE SULFURIC ACID1
Approximate limit, (mg/Nm )a
Substance
Chlorides, as Cl
Fluorides, as F
Arsenic, as AS203
Lead, as Pb
Mercury, as Hg
Selenium, as Se
Total Solids
H2SO Mist, as 100% acid
Water, as H20
Acid Plant Inlet
1.2
0.25
1.2C
1.2
0.25
50C
1.2
50
-
Gas Purification System Inletb
125d
25e
200
200
2.5f
100
10009
-
400 x 103
Notes:
(a) Basis: dry off-gas stream containing 7% sulfur dioxide.
(b) For a typical gas purification system with prior coarse dust removal.
(c) Can be objectionable in product acid.
(d) Must be reduced to 6 if stainless steel is used.
(e) Can be increased to 500-if silica products in scrubbing towers are
replaced by carbon; must be reduced if stainless steel is used.
(f) Can be increased to 5-12 if lead ducts and precipitator bottoms
are not used.
(g) Can usually be increased to 5000-10,000 if weak acid settling tanks
are used.
85
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Table 2-25. RAW WASTE CHARACTERIZATION: ACID PLANT SLOWDOWN'
00
Parameter
pH
IDS
TSS
S04=
Cn-
As
Cd
Cu
Fe
Pb
Hg
N1
Se
Te
Zn
011 and Grease
Flow, 106
Production
Flow/ Prod
Units
pH
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
kg/metric ton
I/day
metric ton/day
kg/metric ton
Plant 1
2.0-2.5
[0.99]a
0.0000
0.044
0.0002
0.0001
0.0014
0.0051
0.0000
0. 0000
0.0001
0. 0000
0.0017
0.0000
0.147
311. (62)
2,400.
Plant 2
1.8
78.5
0.102
7.69
0.0000
0.129
0.0014
0.0018
0.0015
0.0142
0.0000
0.0000
0.0000
0.0000
0.215
--
4.16
528. (264)
15,800.
Plant 3
2.0
410.
3.74
64.0
0.0024
0.004
0.0276
[188.2]
0.1116
0.2501
0.0002
0.0030
0.0268
~
0.436
0.0
10.1
655. (393)
25,700.
Average
2.0
244.
1.92
36.0
0.0008
0.059
0.0097
0.0010
0.0382
0.0898
0.0001
0.0010
0.0090
0.0000
0.218
0.0
~
—
14,700.
a Bracketed values not used in averaging computation.
-------
In this industry, most participate matter from gas cleaning equip-
ment is recycled in dry form or as a water slurry back to the metallur-
gical processes. The small quantities of particulate removed by the
acid scrubbing operations, however, are mixed with a stream of weak
sulfuric acid and cannot readily be recycled. They are discharged with
the acid plant blowdown.
In some sections of the country it is difficult to sell the product
acid, even for less than the cost of manufacture. Therefore, it may be
less expensive to neutralize and discard the acid than to absorb the
costs of shipment to a distant user. Thus, the product acid can itself
become a waste stream.
An acid plant does not produce solid wastes directly, but the
gypsum formed in neutralization of acid can constitute a significant
solid waste.
6. Control Technology - In this country the S02 in the tail gas from
the sulfuric acid plant is not controlled. When S02 emissions are high,
best control may be to increase operating efficiency by adding addi-
tional catalyst stages or by adding heating equipment to maintain proper
catalyst temperature. Changes in the metallurgical operations to pro-
duce a stream of higher S02 concentration at a more uniform rate are
also good controls, if this is possible. Scrubbing of the weak S02
stream for final $62 absorption may also be necessary.
Mist eliminators in the form of packed columns or impingement metal
screens can minimize acid mist emissions. Manufacturers claim elimina-
tion of all but 35 to 70 mg of mist per cubic meter of gas, and the
units sometimes do better. To prevent formation of plumes of mist
during periods of abnormal operations, however, electrostatic precipi-
tators are often used. Better regulation of feed rate and quality also
minimizes acid loss.
As frequently happens in this industry, acid plant blowdown is
sometimes mixed with other waters for treatment or recycle. Table 2-26
lists the practices of the existing smelter acid plants. The practices
outlined for plants 1 and 3 appear to describe the best available con-
trol technology, since by recycle to hot ESP units, the heavy metals
content of this waste partially returns to the metallurgical processing.
If volumes of strong acid must be neutralized, treatment with
limestone, followed by more precise pH adjustment with lime, and dis-
charge to a pond for in-perpetuity storage of the resulting gypsum is
the only tested and economical method of disposal.
7. EPA Source Classification Code - None
87
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Table 2-26. ACID PLANT SLOWDOWN CONTROL AND TREATMENT PRACTICES
Plant
Code
1
2
3
4
5
6
7
8
9
10
11
Discharge
0
0
Oa
0
0
0
0
0
oa
3.4 I/sec
50-190 l/secd
Control and/or treatment practice
Blowdown neturalized with ammonia and used
to precondition converter gases prior to
hot ESP. No discharge.
2/3 of blowdown to reverb brick flue spray
chamber for cooling reverb gases, other 1/3
used to precondition converter gases prior
to hot ESP. Any excess is solar evaporated
on slag dump. No discharge.
Blowdown from packed tower used in open
tower blowdown to clarifier. One-half
recycled to packed tower, other half to
two-stage ammonia neutralization facility.
Then 2.2 I/sec to converter hot ESP for gas
preconditioning and 0.6 I/sec to hot ESP
for gas preconditioning (joins 0.6 T/sec
DMA purge). No discharge.
Blowdown to tailings pond. Pond water
recirculated to mill concentrator. No
discharge.
Blowdown from new scrubbers and mist pre-
cipitators to recycle and tailings thickener
underflow. No discharge.
Blowdown used in mill concentrator circuit.
No discharge.
Blowdown to settling pond and either re-
cycled or wasted. No discharge.
Blowdown to acid ponds and reused in copper
precipitation leach facility. No discharge.
Blowdown currently used to blend fluid- bed
roaster feed. Anticipate closed circuit,
but will eventually send to proposed treat-
ment facility.
Blowdqwn to lime pond, then to tailings pond.
Eventual (8 km of ponds) discharge.
Blowdown to go to new treatment facility
with subsequent discharge.
Anticipated, practice under construction.
88
-------
8. References -
1. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
3. Confidential information from EPA.
4. Control of Sulfur Oxide Emissions in Copper, Lead, and Zinc
Smelting. Bureau of Mines Information Circular 8527, 1971.
5. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW, June
1969.
6. Vandegrift, A.E. (Dr.), Shannon, L.J. (Dr.), Gormena, P.G.,
Lawless, E.W. (Dr.), Sallee, E.E., and Reichel, M. Particu-
late Pollutant System Study -Mass Emissions, Volumes 1, 2, and
3. U.S. Environmental Protection Agency, (NTIS), Durham,
North Carolina. PB-203 128, PB-203 522, and PB-203 521. May
1971.
89
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PRIMARY COPPER PRODUCTION PROCESS NO. 14
DMA SOp Absorption
1. Function - The DMA absorption process scrubs S02 from a stream of
gas, then releases the S02 as a concentrated stream. The principal
applications have been to concentrate streams too weak for efficient use
in sulfuric acid manufacture, to absorb surges in waste gas flow that
could not otherwise be handled by the acid plants, and to manufacture
liquified $03 for sale. Sulfuric acid (Process No. 13) or elemental
sulfur production (Process No. 15) may also be used to control $03
emissions.
Waste gases, after first being cleaned of particulate matter and
dried, pass through a scrubber where most of the $03 is absorbed by
dimethyl aniline (DMA). The gases are then scrubbed with sodium car-
bonate to remove the remaining S02, then with weak sulfuric acid to
reclaim the DMA in the gas stream. The gases are then released to a
stack. In a series of chemical operations, the DMA is recovered for
recycling, and the S02 is recovered as dry, 100 percent sulfur dioxide,
which is compressed, cooled, and stored as a liquid.
The DMA system has the advantage of being relatively insensitive to
changes in S02 concentration of the gas stream, and to changes in gas
flow rate. If part of a waste gas stream is sent directly to an acid
plant at a constant rate, the DMA can handle the remaining gas, which
may be of variable composition, with uneven flow. Then the concentrated
S02 from the DMA plant can be bled back into the acid plant stream as
required to maintain a constant and higher S02 concentration. Thus the
acid plant operates more efficiently and more of the S02 in the waste
gas stream is recovered.
Three smelters in this country operate DMA absorption processes,
handling waste streams that contain from 1 1/2 to 10 percent S02-
Efficiencies up to 99 percent removal from a 5 percent gas stream have
been reported.2 Plant capacities range up to 180 metric tons of liquid
S02 per day.
2. Input Materials - Waste gas containing S02 is the principal input.
Designers of the process do not recommend its use on streams weaker than
2 to 3 percent S02- The gases must be cleaned and dried as described
for the contact sulfuric acid plant (Process No. 13).
A constant feed of soda ash (sodium carbonate) is required for
this process. One smelter reports use of 16 kilograms per metric ton of
S02 produced.
90
-------
Sulfuric acid, 98 percent concentration, is used for drying and
scrubbing. The same smelter reports consumption of 18 kilograms per
metric ton of S02 produced.
This smelter also reports a small loss of the expensive dimethylani-
line, 0.5 kilogram per ton of SC^.
3. Operating Conditions - Feed gas must be cooled to ambient tempera-
turesprT^FTouRA~aBsorption. Temperature in some of the regeneration
steps may reach 150°C. Pressures in the waste gas stream are near
atmospheric, and may approach 3 kg/cm2 in parts of the regeneration
system.
4. Utilities - Electricity is normally used to drive pumps and blowers.
The process is efficient in energy conversion; treatment of a 5 percent
S02 gas stream requires about 160 kWh per metric ton of S02 produced.2
Steam is required at a rate of 1.0 to 1.5 tons per ton of S02
produced.2 Noncontact cooling water is required in the amount of 1250
liters per metric ton of Sfy- A small amount of process water is
needed to compensate for purge and evaporation.
5. Waste Streams - Exit gas contains 0.05 to 0.3 percent S02 and no
particulate matter. Temperatures are approximately ambient. The DMA
process adds only an insignificant quantity of carbon dioxide to the
stream, and very small amounts of DMA or its decomposition products may
escape the third stage of scrubbing.
The waste gas stream may carry with it an entrained mist of dilute
sulfuric acid from the third stage.
This process requires scrubbing of the gas stream in a weak sul-
furic acid column and thus may produce a liquid waste blowdown stream
similar to that from a sulfuric acid plant. Normally, however, the same
scrubber is used to treat gases that feed both the acid plant and DMA.
A liquid waste, continuously purged from the DMA process, consists
of water, sodium sulfite or bisulfite, and sodium sulfate. The quantity
is about 18 kilograms per metric ton of S02 produced, when treating gas
with 5 percent S02 content. The stream typically contains about 4.5
percent dissolved solids, 25 mg/1 DMA, and 18 mg/1 suspended solids,
with a pH around 5.8.1
The process produces no solid wastes.
6. Control Technology - Following DMA absorption, further treatment of
the waste gas stream for S02 removal is not normally required. An elec-
trostatic precipitator is usually installed to eliminate acid mist
carryover.
91
-------
If a separate dryer is used for DMA gas treatment, disposition of
the blowdown would be the same as that for the sulfuric acid plant
blowdown. Best available control technology is to neutralize this
stream and recycle it as coolant to a hot ESP unit, thus returning the
metals content to metallurgical processing.
The purge stream from the DMA process is the only waste of this
character generated by the primary copper industry. It is a clear
stream with a BOD and a COD and is quite concentrated with non-recov-
erable minerals. Each of the three operating DMA installations handles
the purge stream differently. One adds it to the concentrator circuits;
one mixes it with the acid plant blowdown, which is .in turn sent to a
hot ESP unit; the third uses activated carbon to absorb the DMA content,
then uses it as part of a fluid-bed wet feed blending, which returns it
directly to the metallurgical processing. This last alternative may be
the best control technology, with or without the activated carbon ab-
sorption. The sodium may then combine into the slag where it will not
increase alkali content of the concentrator water, thus reducing poten-
tial of recycle. An excess of sodium salts may plate out in a hot ESP
unit. Further study to establish the best disposition of this stream is
indicated.
7. EPA Source Classification Code - None
8. References -
1. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
2. Fleming, Edward P. and Fitt, T. Cleon. Liquid Sulfur Dioxide
from Waste Smelter Gases. I&EC, November 1950, Vol. 42, No.
11 pp. 2253-2258.
3. Halley, J.H. and McNay, B.E. Current Smelting Systems and
Their Relation to Air Pollution. Arthur G. McKee and Company,
San Francisco, California 35224. September 1970.
92
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PRIMARY COPPER PRODUCTION PROCESS NO. 15
Elemental Sulfur Production
1. Function - Sulfur dioxide from a waste stream can be converted into
elemental sulfur by one of several methods. Other S02 removal methods
are sulfuric acid production (Process No. 13) and DMA absorption (Process
No. 14). Sulfur is easier to store than sulfuric acid and is less
expensive to transport. Although the market is also better than for
acid, the sulfur now produced by this process is not competitive with
cheap mined sulfur from the Texas coast.
All variations of the process use high temperatures to oxidize
methane to carbon dioxide, while simultaneously reducing the S02 to
sulfur. All use special catalysts and are sophisticated multi-step
processes, efficient in energy conservation. Like the ammonia plants or
oil refinery units that they resemble, they are most efficient as large-
capacity installations.
None of the smelters in this country now include sulfur production
facilities, although one plant is being constructed.
2. Input ^Materi alls - Gas must enter the process at constant flow rate
and composition.TFfe gases must contain 5 to 7 percent S021 and must be
free from particulate matter and metal-containing fumes.
.Methane from natural gas is added in the stoichiometric ratio of 1
kilogram to each 8 kilograms of S02- Additional natural gas is required
as fuel.
3. Operating Conditions - Although temperatures and pressures differ
in the various process modifications, most will operate between 1000°
and 1500°C at pressures less than 25 kg/cm2. One variation requires
1250°C for proper equilibrium.2 The equipment is not normally enclosed
in a building.
4. Utilities - Natural gas fuel is assumed for most designs, and
electricity is required for pumps, blowers, and compressors. One varia-
tion incorporates electrostatic precipitators as integral components,
which require electric power. No utility estimates have been reported.
Cooling water is required for portions of the process.
5. Waste Streams - All these processes claim removal of more than 90
percent of the S02 from the gas stream, and one claims up to 95 percent.
The waste gas stream can therefore be expected to contain no more than
0.7 percent S02- Slight emissions of H2S gas may occur in some of the
process variations.
93
-------
No liquid wastes are expected from this process. The water formed
by the reaction normally escapes by evaporation into the waste gas
stream.
This process produces no solid wastes. Fugitive dust from the
product sulfur may be possible.
6. Control Technology - The only further control of the waste gases
from this process is wet scrubbing, as described in connection with the
reverberatory furnace (Process No. 6).
7. EPA Source Classification Code - None
8. References -
1. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
2. Fleming, E.P. and Fitt, T.C. High Purity Sulfur from Smelter
Gases. Industrial and Engineering Chemistry. Volume 42, No.
11. March 1950. pp. 2249-2253.
94
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PRIMARY COPPER PRODUCTION PROCESS NO. 16
Arsenic Recovery
1. Function - Most of the arsenic present in a copper ore concentrate
will be volatilized in the roasting and smelting processes, and will
appear in the dusts collected from the electrostatic precipitators and
other particulate removal equipment. One smelter, treats those dusts to
extract arsenic for sale. Since demand for arsenic as a commercial item
is very small, this one smelter satisfies most of the U.S. demand for
this material.
The recovery process consists of placing the collected dusts in a
Godfrey roaster, in which they are heated until the arsenic vaporizes.
The vapors are condensed in chambers, and are then resublimed and con-
densed to yield an arsenic trioxide product more than 99 percent pure.
Arsenic metal is also occasionally produced by reducing the oxide with
carbon in an atmosphere deficient in oxygen.1
All these operations are batch-type and are done on a small scale.
2. Input Materials - Flue dusts from the multiple-hearth roasters and
reverberatory furnace at this location are the principal input. Dusts
of high arsenic content from other smelters were also used at one time,
but are no longer accepted by this smelter. Flue dusts are charged into
the furnace along with a small amount of pyrite to minimize conversion
to arsenites.1
Charcoal in small quantity may be used for arsenic metal produc-
tion.
3. Operating Conditions - Arsenic trioxide is assumed to be completely
vaporized from the dusts when the temperature reaches 650° to 700°C. It
recondenses in the cooling chambers at around 200°cj Atmospheric pres-
sures are used.
4. Utilities - Gas-fired burners are used to heat the charge, and non-
contact cooling water to assist in condensing. No quantities have been
reported.
Water is used to wash down dust and spills within the plant.
5. Waste Streams - Because there is no mechanical movement of material
within the Godfrey roaster, few particulates are generated during the
operation. Fugitive dusts are generated during the handling of the dry
materials.
95
-------
The gas stream from the roaster contains carbon dioxide and water
vapor, and may contain small amounts of S02 and arsenic fumes. No
analysis has been reported.
A water waste is generated by daily washdown of the plant to
remove settled dusts from materials handling. Table 2-27 gives the
analysis of this water.
No solid wastes are generated by this process. After treatment for
arsenic removal, the remaining dusts are returned to the smelting fur-
nace.
6. Control Technology - Fugitive dust emissions in and from the plant
building are controlled by the use of fabric filter baghouses on venti-
lating air streams.
Effluent process gas streams are currently routed through an electro-
static precipitator before being vented to a tall stack. In the near
future, a fabric filter baghouse will be installed to remove particulate
from the waste gas.
The washdown from this plant mixes with another waste stream and
discharges to a pond. Table 2-27 indicates that up to a kilogram of
arsenic may enter the pond each day from this source. The degree to
which it becomes soluble has not been reported.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Development Document for Proposed Effluent Limitations Guide-
lines and New Source Performance Standards for the PHmary
Copper, Lead, and Zinc Segment of the Nonferrous Metals
Manufacturing Point Source Category (Draft). U.S. Environ-
mental Protection Agency. Washington, D.C. Contract No.
68-01-1518. December 1973.
96
-------
Table 2-27. ANALYSIS OF ARSENIC PLANT WASHDOWN WATER'
Parameter
As
Cu
Zn
Pb
Cd
Hg
Se
Te
Ni
Fe
so4=
Cn-
Oil and grease
PH
Concentration
rng/1
310.
88.4
37.0
7.7
1.05
0.0003
0.04
0.43
0.75
9.4
340.
0.01
0.04
3.8 to 4.4
97
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PRIMARY COPPER PRODUCTION PROCESS NO. 17
Fire Refining and Ariode Casting
1. Function - Impure or "blister" copper from the converters must be
refined to remove impurities. This is partially accomplished by fire
refining, which is the last major process that occurs at a copper smelt-
er.
Blister copper is placed in a fire refining furnace, a flux is
usually added, and air is blown through the molten mixture. This blow
oxidizes most of the remaining sulfur, vaporizes some impurities, and
converts others into a slag. The mixture is then "poled" with wooden
logs, or otherwise treated to reduce the excess oxygen in the mixture.
The copper is then poured into molds and cooled with water sprays or by
immersion in a tank of water. The resulting anodes are sent to an
electrolytic refinery for further processing.
A small percentage of the copper may be subjected to more complete
fire refining to produce ingots or special castings for direct sale.
This fire refined copper is used for manufacture of alloys and other
special purposes, and contains no impurities other than oxygen in
significant amounts. Anode copper is less completely refined, but is
more than 99 percent pure. The general range of analysis is shown in
Table 2-28.
2. Input Materials - Copper from the converting operation (Processes
Nos. 9 and 10) is the principal input, usually charged into the fire
refining furnace while still molten.
Various slag-forming fluxes may be added. These include silica
sand and sodium carbonate.
Wooden poles are still frequently used for the reducing step of the
process. The wood decomposes when thrust into the molten copper, pro-
ducing a variety of carbonaceous products that remove oxygen from the
mixture. Instead of wood, some smelters use hydrogen, natural gas, or
ammonia for reduction.
The casting molds are sprayed with a mold dressing of silica flour
or potassium alum to keep the castings from sticking.^
3. Operating Conditions - Temperature in the furnace is around 1100°C.
Pressure is atmospheric.
4. Utilities - If molten blister copper is charged to the furnace,
additional fuel is required only in small amounts. Gas or oil is used
for heating, or to melt the charge if cold feed is used. In a plant
producing 91,000 metric tons of copper per year, fuel consumption for
this process has been estimated at 8600 kg-cal (10 kWh) per hour of
operation.'
98
-------
Table 2-28. GENERAL RANGE ANALYSIS OF ANODE COPPER*'
Constituent
Copper
Oxygen
Sulfur
Arsenic
Antimony
Bismuth
Lead
Nickel
Selenium
Tellurium
Gold
Silver
Platinum
Palladium
Content, %
99.0-99.6
0.1-0.3
0.003-0.01
0.003-0.2
0.001-0.1
0.001-0.01
0.01-0.2
0.01-0.2
0.01-0.06
0.001-0.02
3. 4-102. 6b
68-3080b
N.A.
N.A.
Extremes omitted
J g/metric ton
i.A. - not available
99
-------
Compressed air is used to oxidize the molten mixture in the fur-
nace. No quantities have been reported.
Water is used for direct cooling of the casting machine and the
copper anodes. This is usually a recirculated stream or is reclaimed
water from combined sources. Quantities are shown in Table 2-29 for
five smelters.
5. Waste Streams - Gases from the fire refining furnace may contain
fumes of zinc and cadmium^ that probably condense within the exhaust
duct. Concentration of S02 has been estimated at 0.38 kilogram per
metric ton of copper treated.4 There are no recorded data giving the
quantity of this waste gas, but particulate loading has been reported as
5-20 kilograms per metric ton of copper produced.2 Gas temperature is
about 1000°C.5
The water used for anode cooling is reported to pick up additional
amounts of arsenic, copper, and zinc, and also to pick up aluminum and
chlorides, probably from mold dressing compounds. Table 2-30 lists the
data reported for one anode cooling operation. The "net change" re-
presents the difference between inlet and outlet concentrations.
There are no solid wastes from this process. All slags are re-
turned to the metallurgical processing.
6. Control Technology - No control of waste gas from the fire refining
process is being exercised by any of the operating copper smelters.
Apparently it is assumed that this is one of the cleaner gas streams
from pyrometallurgical operations because of the relative purity of the
input copper.
Table 2-31 lists the controls of contact cooling water being prac-
ticed by the domestic smelters. This list includes water used for
cooling of both anodes and blister copper direct from the converter.
Discussion of the control of mixed waste streams appears in the section
on water management of the sulfide mining and concentrating industries.
7. EPA Source Classification Code - 303-005-05
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Compilation of Air Pollutant Emission Factors. AP-42. U.S.
Environmental Protection Agency. Research Triangle Park,
North Carolina. March 1975.
100
-------
Table 2-29. WATER REQUIREMENTS FOR COPPER REFINERIES5
Plant
A
B
C
D
E
Water intake, liters
per metric ton
of metal producted
4,000
9,000
13,000
3,000
6,000
Water consumed
(Intake minus discharge),
liters per metric
ton metal producted
4,000
700
1,200
1,900
0
101
-------
Table 2-30. WASTE EFFLUENTS FROM ANODE COOLING WATER?
Parameter
Chloride
Aluminum
Arsenic
Copper
Zinc
Flow, 106
I/day
Production,
metric ton/day
Net change,
mg/1
8.7
0.12
0.01
8.53
0.25
0.95
265
Net loading
kg/day
7.8
0.11
0.01
8.07
0.24
kg/metric ton
0.029
0.0004
<0.0001
0.030
0.0009
Source: RAPP.
102
-------
Table 2-31. CONTACT COOLING WATER CONTROL AND TREATMENT PRACTICES6
Plant
code
1
2
3
4
5
6
7
8
9
10
11
12
Discharge •
0
0
0
0
Intermit-
tent
0
0
0
0
0
0
5670 m3/day
Control and/or treatment practice
Anode casting: water in closed circuit with
cooling tower, cooling tower blowdown joins
blowdown from wire-bar casting cooling tower
blowdown, entire blowdown to side-stream
filter, anticipate total water recycle.
Anode casting: water directly reused in mill
concentrator circuit. No discharge.
Anode casting: water collected in mill tail-
ings thickener, all flow recycled (with some
evaporation) to mill concentrator. No dis-
charge.
Blister cake cooling: air cooled with some
water spray; spray water totally recycled
from cooling pond. No discharge.
Fire-refined (cathode) - shape casting: water
mostly recycled, with small intermittent
discharge.
Fire-refined casting: water to thickener,
overflow recycled. No discharge.
Anode casting: water in closed circuit with •
cooling tower, blowdown to evaporation pond.
No discharge.
Anode casting: water in closed circuit with
cooling tower, blowdown reused in mill con-
centrator. No discharge.
Anode casting: water to tailings thickener,
reused in mill concentrator. No discharge.
Anode casting: water all used in mill con-
centrator circuit. No discharge.
Anode casting: water in closed circuit with
100 percent circulation. No discharge.
Anode casting: water collected in slag
settling pond, part is recirculated for
slag granulation 53,000 m3/day. Remainder
5700 m^/day discharged to tailings ponds.
Eventual (8 km of ponds) discharge.
103
-------
Table 2-31 (continued). CONTACT COOLING WATER CONTROL
AND TREATMENT PRACTICES6
Plant
code
13
14
15
Discharge
18 I/sec
•v-340 m3/day
(125 I/sec,
45 min/day)
0
Control and/or treatment practice
Anode casting: once-through water, part used
for shot copper cooling, remainder discharged.
Shot copper cooling: Intermittent flow, all
discharged. Plan to treat water in proposed
treatment facility with anticipated discharge.
Blister cake cooling: water consumed during
spraying and air cooling. No discharge.
104
-------
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
4. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
5. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
6. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-74/032-b, February 1975.
105
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 18
Electrolytic Refining
1. Function - Although copper produced at a smelter contains less than
1 percent impurities, this is too much to meet most of today's quality
specifications. The electrolytic refinery reduces the impurities to
approximately 0.05 percent.
The refining is done by passing a direct current of electricity
through two copper electrodes that are immersed in a bath of acidic
copper sulfate solution. The anode is a casting of impure copper from
the smelter, and the cathode is a "starting sheet" stripped from a
previously refined block of electrolytic copper. The electric current
causes the copper to dissolve from the anode and deposit at the cathode.
Impurities either will not dissolve in the electrolyte, or will not
plate out at the cathode, so they collect either as slimes in the bottom
of the cell or as soluble ions in the electrolyte.
Seven electrolytic refineries are operating in the United States.
Four are located near a copper smelter, and the others are distant from
smelters.
2. Input Materials - The principal input is the anode castings from
copper smelters. About 85 percent of the anodes in use at any one time
are directly from the smelter. The remainder are made at the refinery
by melting and re-casting partially electrolyzed anodes.
The electrolyte is sulfuric acid, which is either fresh acid or
acid reclaimed from the electrolyte purification process (Process No.
18).
Various additives are used to ensure a smooth deposit at the
cathode. Chlorides are added to cause silver to precipitate irto the
slimes.
3. Operating Conditions - Electrolytic cells are normally maintained
at 60 to 65°CJPressures are atmospheric.
4. Utilities - Electric power is the only source of energy. Approxi-
mate ly~T5(57000~ to 190,000 kg-cal are required to produce a metric ton of
cathode copper.2 The direct current required for the cells is produced
within the refinery by rectifiers or motor-generator sets. Additional
electricity is required for the auxiliary materials handling equipment.
Water is used for washing the electrodes as they are removed from
the cells. In most refineries, this is specially treated water, usually
steam condensate, since untreated water contains minerals that affect
106
-------
the quality of the product. This same water is used for make-up to the
electrolyte system to replace that lost in purge and evaporation.
5. Waste Streams - There is no appreciable pollution of the air from
an electrolytic refinery.
Most refineries reclaim the copper from impure solutions (see Pro-
cess No. 19), but two do not, and therefore create a substantial liquid
waste directly from the electrolytic cells. Table 2-32 gives the range
of composition of the electrolyte solution, along with the composition
of the refined copper and the slime that is recovered from the bottom of
the cells.
Slime is periodically cleaned from the cells and processed for
recovery of the gold, silver, and other valuable elements (see Process
Nos. 21 and 23). Because this slime represents a product of consider-
able value to the copper industry, procedures at most smelters are
designed to place as much of these valuable elements as possible into
the anode copper.
In any plant that handles highly corrosive liquids in large quan-
tities, there are leakages, spills, and occasionally major discharges,
frequently not expected and normally not included in waste tabulations.
No solid wastes are produced by an electrolytic refinery.
6. Control Technology - Of the two refineries that do not reclaim the
impure electrolyte solution, one treats this stream separately by plac-
ing it in a lined pond and allowing it to evaporate to dryness. Cli-
matic conditions at this site make this procedure workable. The other
refinery follows a practice often used in the copper industry and mixes
it along with other wastes into a tailings pond, where lime is added to
neutralize the acid. This refinery is also in an arid section of the
country. Both refineries report no discharge into public waters.
Ultimate disposal of the solids from these evaporations has not been
reported.
Most refineries have demonstrated an economic benefit from the
reclamation and partial recycle of spent electrolyte; this represents
the best available control technology (see Process No. 19).
There are usually no controls specifically designed to handle
spills, leakages, and accidental discharges of electrolyte.
7. EPA Source Classification Code - 303-005-05
107
-------
Table 2-32. GENERAL RANGE ANALYSIS OF ELECTROLYTE, REFINED
COPPER, AND ANODE SLIME3'2'3
Constituent
Sulfuric acid
Copper
Oxygen
Sulfur
Arsenic
Antimony
Bi smuth
Lead
Nickel
Selenium
Tellurium
Gold
Silver
Platinum
Palladium
Iron
Electrolyte,
g/i
170-230
45-50
0.5-12
0.2-0.7
0.1-0.5
2.0-20.0
Refined copper,
%
99.95
0.03-0.05
0.001-0.002
0.0001-0.001
0.0002-0.001
0.00001-0.00002
0.002-0.0010
0.0001-0.002
0.0003-0.001
0.0001-0.0009
0.68-0.242b
1. 71-17. lb
tr.
tr.
tr.
Raw slime, %
(dry basis)
20-40
2-6
0.5-4.0
0.5-5.0
tr.
2.0-15.0
0.1-2.0
1.0-20.0
0.5-8.0
1714-10286
34285-274283
N.A.
N.A.
0.1-0.2
Extremes omitted.
g/metric ton.
tr. = trace.
N.A. = Not available.
108
-------
8. References -
1. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b, February 1975.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
109
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 19
Electrolyte Purification
1. Function - In operation of the electrolytic cells in a refinery,
certain impurity elements become dissolved in the electrolyte solution.
If a portion of the electrolyte is not removed from the circulating
stream, concentrations of these impurities will become so high that they
begin to deposit with the refined copper. In most cases, the purge
stream is processed to recover some of its constituents.
All but two of the refineries in this country remove the copper
from the purge stream. This is done in special "liberator" electrolytic
cells that use insoluble lead anodes and sheets of copper as cathodes.
The copper and frequently some of the impurities deposit on the cathode.
The plates of copper return to the metallurgical processing either
within the refinery or in the smelter, depending on quality. Some of
the remaining impurities collect in the liberator cells as a sludge.
A few refineries recover a portion of the sulfuric acid from the
purge stream by use of dialysis equipment. The dialyzers provide a
partial separation of the acid and produce a stream in which impurities
are more concentrated. The acid is returned to the electrolyte circula-
tion.
Effluent from the liberator cells or the dialysis equipment may be
concentrated further by removing water in vacuum evaporators. Con-
centration of the acid produces a sludge, which has a high concentration
of nickel sulfate and usually also contains iron and zinc. This sludge
can be filtered out, and then part of the acid can be returned to the
electrolyte system, or it may be discarded or sold.
Various smelters may practice all, part, or none of these treat-
ments. Three refineries recover nickel. One refinery consumes all
spent electrolyte in an associated chemical operation.!
2. Input Materials - The input is the purge stream from the recircu-
lating electrolyte. The range of analysis is given in Table 2-32
(Process No. 18). This shows it to be a stream of warm, concentrated
acidic copper sulfate solution, containing also nickel, arsenic, anti-
mony, and bismuth. It also contains smaller quantities of iron, cobalt,
zinc, lead, selenium, tellurium, and other elements.
3. Operating Conditions - Temperatures are less than 100°C and pres-
sures are atmospheric, except in some evaporation operations.2
110
-------
4. Utilities - Electricity is used to drive pumps and mechanical
equipment, and to operate the liberator cells. Since the average con-
centration of salts in the liberator electrolyte is much less than in
the main electrolytic cells, the liberator requires 2 to 5 times as much
current to remove the same amount of copper. Usage is reported as 350
to 700 kWh per metric tonJ
Steam may be required for vacuum production, and fossil fuels may
be used for direct-fired evaporation.
5. Waste Streams - The principal characteristic of the waste streams
from electrolyte treatment is that, combined, they must contain almost
all the arsenic, antimony, and bismuth that comes in with the anode
copper.3 Some of these elements may be returned to the smelter with
liberator cathodes, even though there is no way to dispose of them
there. Examination of Table 2-28 (Process No. 1.7) shows that a ton of
arsenic enters the electrolytic refinery with as little as 500 tons of
anodes, and the best analysis shown in this table would provide a ton of
arsenic waste for each 40,000 tons of copper.
Some of the arsenic is known to escape from the second stage of the
liberator cells as arsine (AsHs).^ This very poisonous gas can accumu-
late to dangerous levels if the liberator cells are not well ventilated.
Arsine will slowly oxidize in the atmosphere to arsenic trioxide and
water. This is the only atmospheric pollutant recognized from an elec-
trolytic refinery, but quantities apparently have not been measured.
This is also the only reference to arsenic in a waste stream from this
process.
Unless all of these elements are returned to the smelter with the
liberator cathodes, the arsenic, antimony, and bismuth must exit with a
purge of electrolyte acid. It appears that accumulation of these ele-
ments must result in either continuous or occasional disposal of a
quantity of "black acid", regardless of the extent of electrolyte treat-
ment employed.
In some refineries evaporated water from the electrolyte consti-
tutes another waste stream, which usually is mixed with volumes of steam
condensate and direct cooling water in barometric leg discharge devices.
Table 2-33 provides an analysis from such a source.
No solid wastes are discharged from this process.
6. Control Technology - Arsine formed in the liberator cells can be
readily oxidized or can be scrubbed to form a liquid waste. Best con-
trol technology cannot be evaluated unless an order of magnitude of the
quantity being released is known. If the amount is fairly large, it
should be possible to design an oxidation process that could recover
this as dry arsenic trioxide.
Ill
-------
Table 2-33. WASTE EFFLUENTS FROM NiS04 BAROMETRIC CONDENSER1
Parameter
PH
Alkalinity
COD
Total Solids
Dissolved Solids
Suspended Solids
Sulfate (as S)
Arsenic
Cadmi urn
Copper
Iron
Lead
Nickel
Z'inc
Flow, 106
I/day
Production,
metric ton/day
Intake,
mg/1
6.5
90
750
21,080
21 ,060
18
1,722
<0.010
<0.20
<0.20
<0.50
<0.50
<0.50
<0.20
Discharge,
mg/1
6.6
450
24,000
24,000
18
1,060
<0.010
<0.20
<0.20
1.30
<0.50
<0.50
0.48
11.4
415
Net change,
mg/1
neg
2,920
2,920
neg
<1.3
<0.48
Net loading
kg/day
<15
<5.4
112
-------
If it is possible to sell or give away the black acid to the
fertilizer industry, as has been reported, this would transfer the
impurity elements into the gypsum ponds from phosphate rock treatment.
This would be a better solution than disposition in a local tailings
pond, which is already loaded with metal ions from other sources. The
neutral to acidic nature of phosphate ponds may also cause a greater
degree of precipitation of arsenic and antimony than would occur in the
alkaline water designed to precipitate copper.
For the effluent from a vacuum evaporator condenser, the best
control technology is to avoid overloading the evaporator. Because of
the low volatility of sulfuric acid and other components of this stream,
evaporation of the water should be easy if the evaporator is properly
designed, instrumented, and operated.
7. EPA Source Classification Code - None
8. References -
1. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-74/032-b. February 1975.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
4. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
113
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 20
Melting and Casting Cathode Copper
1. Function - Refined copper from the electrolytic cells is melted and
recast into the shapes required by fabrication industries. There is
usually also a final adjustment of the oxygen content of the finished
product.
Special equipment used for these operations ranges from direct-
fired reverberatory furnaces to continuous casting machines. Electric
arc and induction furnaces may be used to melt or hold the molten cop-
per. The trend in this process is toward continuous or semicontinuous
equipment to provide closer control of product quality and to minimize
energy requirements.
2. Input Materials - The principal input is cathodes from the electro-
lytic cells. These are washed free of electrolyte and slime prior to
delivery to this process.
Mold dressings such as bone ash may be used in some operations.
For production of certain grades of copper, special oils, graphite, and
phosphorous-copper alloys may be added at various stages in the process.
Use of reverberatory furnaces for this process requires addition of
a flux and possibly a "poling" operationJ This modification is com-
parable to fire refining and anode casting (Process No. 17).
3. Operating Conditions - Depending on the process details, tempera-
tures range~TromnrT301:rrTo 1215°C.2 Pressures are atmospheric. Special
reducing atmospheres may be used in some operations. Open molds are
usually cooled with water sprays to around 150°C.
4. Utilities - Electricity or fossil fuel may be used for melting.
The newest electrical furnaces are reported to operate at high thermal
efficiencies; power consumption is rated at 250 to 300 kW for main-
taining a molten charge of 55 to 90 metric tons of copper in an electric
arc furnace.3
Electricity is used to power materials handling and casting equip-
ment.
Both contact and noncontact cooling waters are used to cool the
casting equipment and the cast shapes. One refinery reports a water
usage of 320,000 liters per day.4
114
-------
5. Waste Streams - Reverberatory furnaces, still occasionally used for
refined copper melting, produce a gaseous discharge to the atmosphere;
the quality of this emission has not been reported.
Table 2-34 provides the analysis of water used for cooling the
refined copper shapes at two refineries. Another report showed an
increase in chlorides of 58.6 grams per liter,3 but the origin of this
ion was not defined.
There are no solid wastes from this process.
6. Control Technology - There is no control of air emissions from a
melting or casting operation, and specific control is probably not
required.
Water from this process is often cooled prior to discharge, but is
not usually otherwise specially treated.
7. EPA Source Classification Code - 3-03-005-008 - finishing opera-
tions.
8. References -
1. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
4. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
115
-------
Table 2-34. ANALYSIS OF WATER USED TO COOL REFINERY SHAPES
(Concentrations in mg/1)
Parameter
PH
TDS
TSS
so4
As
Cd
Cu
Fe
Pb
Hg
Se
Te
Zn
Oil and
grease
Plant X
Inlet
water
7.6
1430.
0.0
240.
0.001
0.001
0.30
0.02
0.007
0.00350
0.001
0.001
0.0
Wirebar
cooling
7-8
1250.
12.5
240.
0.001
0.001
0.69
0.13
0.007
0.00425
0.001
0.067
2.0
Semi contin-
uous cake
casting
8.
1400.
0.0
270.
0.001
0.001
0.18
0.04
0.003
0.0001
0.001
0.001
0.0
Plant Y
Inlet
water
7.1-7.6
0.2
0.5
0.001
0.0008
0.021
1.2
0.078
). 00004
0.040
0.35
0.14
Wirebar cooling
recycle
8.0-8.4
0.1
0.4
0.001
0.0021
3.5
1.7
0.068
0.00004
0.040
0.088
0.1
116
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 21
Slime Acid Leach
1. Function - The first step in treatment of slimes from the cells of
an electrolytic refinery is removal of the copper. This may be by
direct roasting (Process No. 23), or the slimes may be first leached
with acid to extract a portion of the copper prior to the roasting step.
The acid leach is accomplished in a pressure filter, through which
sulfuric acid is circulated. Copper dissolves in the acid as a solution
of copper sulfate. This solution is either mixed with the electrolyte
in the refinery cells,' or with the electrolyte purge to the liberator
cells, or may be used for copper sulfate production (Process No. 22).
2. Input Materials - The primary input is the slime from the electro-
lytic cells (Process No. 18), which may contain 20 to 40 percent cop-
per. 2 Small particles of metallic copper will be present.
Sulfuric acid is the leach solvent. Concentration and quantity of
the acid vary with the slime composition.
3. Operating Conditions - There is normally no heating of the circu-
lating solution, but chemical action may cause a slight temperature rise
above ambient. Pressures are atmospheric to slightly higher, not ex-
ceeding 1 kg/cm^.
4. Utilities - Electricity is used for acid circulation in small
quantity.Water or steam condensate is used to wash the leached slimes
prior to transferring them to the roaster.
5. Waste Streams - A minor evolution of S02 in this process is due to
the reaction of copper metal with the acid.3
Except for accidental spills or pump leakage, there are no liquid
or solid wastes. All materials are transferred to other processes.
6. Control Technology - If this were a larger-scale process, control
of S02 by blending witn other streams (if available) or by scrubbing
would be the best control technology. Since quantities are small, none
of the refineries control this emission except by local ventilation.
7. EPA Source Classification Code - None
117
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8. References -
1. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategbry and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-74/032-b. February 1975.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Leigh, A.M. Precious Metals Refining Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 95-110.
118
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PRIMARY COPPER PRODUCTION PROCESS NO. 22
CuSO. Precipitation
1. Function - The function is to precipitate copper sulfate in crystal
form as a salable by-product. The solution from water or acid leach
constitutes part or all of the source. In this process, copper powder
is first added if there is excess acidity, and excess water is evaporated.
When the mixture cools, crystals of copper sulfate form. The concent-
rated liquor either returns to the electrolytic cells or is transferred
to chemical operations for the manufacture of other products. The
crystals may be heated to remove water of hydration prior to sale.
2. Input Materials - The input is the leached solution from Process
No. 21, containing copper sulfate and sulfuric acid, or from Process No.
24, containing copper sulfate in water.
3. Operating Conditions - Atmospheric evaporators are usually used,
with boiling temperatures less than 125°C. Crystallization occurs in
atmospheric vessels. The crystals may.be heated as high as 600°C after
separation from the mother liquor if anhydrous copper sulfate is being
produced.'
4. Utilities - This is primarily a chemical type process, using either
direct gas-fired or steam-heated evaporation equipment, noncontact
cooling water for crystallization, and electricity for solution transfer
and auxiliaries. Utility usage is not reported, but quantities are
small.
5. Waste Streams - Use of copper powder for neutralizing excess acid
will cause a slight evolution of $03, which will be stripped into the
atmosphere during evaporation.2
Water evaporated from the solution will condense as a wastewater,
or will be lost into the atmosphere if direct-fired evaporators are
used. Some carryover of entrained solution could occur. There are no
reports of the waste from this source.
The process produces no solid wastes.
6. Control Technology - No controls are currently associated with this
proce?s~i If quantities of S02 evolution were greater, scrubbing or
mixing with another stream for combined S02 treatment would provide
adequate control.
There is no report on the disposition of water from this source.
119
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7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Leigh, A.H. Precious Metals Refining Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 95-110.
120
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PRIMARY COPPER PRODUCTION PROCESS NO. 23
Slimes Roasting
1. Function - Roasting of slimes from the cells of an electrolytic
refinery allows removal of the copper content. A portion may be removed
by acid leach of the slimes (Process No. 21). Heating the slimes in a
strong acid environment converts the remaining copper to soluble copper
sulfate, which can be removed by a subsequent water leach process.1
Roasting also converts some of the silver and tellurium to soluble salts
and volatilizes some of the selenium.
2. Input Materials - The principal input is the slime materials,
either direct from the electrolytic cells or as residue from acid leach.
Fluxes in the form of sulfuric acid and sodium sulfate are used to
ensure complete reaction of almost all the copper present in the slime,
which may be as much as 40 percent by weight of the slimes.2 One report
gives the sulfuric acid consumption as 1.74 kilogram of acid per kilo-
gram of slime treated.2
Muds from the scrubber (Process No. 26) are also recycled to this
roaster.3
3. Operating Conditions - Temperatures in the roaster are maintained
between 540° and 650°C.Pressures are atmospheric.1
4. Utilities - Gas or oil is used for heating, and electricity for
driving mechanical equipment. Quantities are not large, because of the
small scale of this equipment.
5. Waste Streams - The gas leaving the roaster contains highly miner-
alized particulates and fumes. Roasting breaks down silver and copper
selenides, releasing S602-1 Arsenic, tellurium, and trace amounts of
lead also are present as fumes. The stream contains S02 and dusts,
which consist of all the elements present in the slime. This gas stream
normally passes to the scrubber (Process No. 26), but any loss can
represent a hazardous waste. No analyses of this stream have been
reported.
Solids from the roaster contain the most valuable metals, and are
usually carefully transferred to the water leach equipment (Process
No. 24). There is no solid waste.
No liquid wastes are generated.
121
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6. Control Technology - Proper transfer of the highly mineralized
gases from this process to the scrubber is the best control.
7. EPA Source Classification Code - None
8. References -
1. Leigh, A.M. Precious Metal Recovery Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 95-110.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032. February 1975.
122
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PRIMARY COPPER PRODUCTION PROCESS NO. 24
Slime Water Leach
1. Function - The objective of this process is to reprecipitate all
the silver and tellurium that has been made water-soluble in the roast-
ing process, and to dissolve and separate all the soluble copperJ
Powdered copper is added to roasted solids in calculated quantity.2
The mixture is then slurried with water in a tank, and by a cementation
reaction, the silver and tellurium are precipitated. The mixture is.
allowed to stand to cause these reactions to approach completion and to
allow the solids to settle. The liquid is then decanted off, and the
slurry is filtered. The liquid solution of copper sulfate returns to
the electrolytic cells or is used for copper sulfate production (Process
No. 22). The filter cake is transferred to the Dore" Furnace (Process
No. 25).3
2. Input Materials - Roasted slimes from Process No. 23 is the prin-
cipal ~Tnpurt!Powdered copper in slightly less than stoichiometric
proportions is added.
3. Operating Conditions - The temperature in the leach tank is less
than 100°C, but is not carefully control led.3 Pressures are atmospher-
ic, rising to less than 3 kg/cm^ during filtering.
4. Utilities - No external heat is added to this process. Heat in the
hot solids from the roaster adds incidental heat.
Either deionized water or steam condensate is used to prevent
introduction of foreign elements into the electrolyte solution.
5. Waste Streams - Except for accidental spills, no wastes are gener-
ated by this process.
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
123
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Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
3. Leigh, A.H. Precious Metals Refining Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. Feburary 25 - March 1, 1973). p. 95-110.
124
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PRIMARY COPPER PRODUCTION PROCESS NO. 25
Dbre Furnace
1. Function - This process separates the trace elements contained in
the slimes into several distinct fractions, each of which is either sold
or further treated. The most valuable fraction is a Dore" metal, con-
sisting primarily of silver, gold, and the platinum group metals.
The equipment is a special small reverberatory furnace, which
removes groups of elements in separate slag-producing steps. The filter
cake from the water leach process is mixed with a silica flux, charged
into the furnace, and heated. A slag forms, containing primarily the
lead, iron, arsenic, and antimonyJ This "sharp slag" is withdrawn and
can be sent for further processing to a lead smelter. Sodium salts are
then added to the furnace, and a soda slag forms. This slag contains
selenium and tellurium and any residual arsenic and antimony? and is
further treated (see Process No. 27). An oxidative slag is then formed
by blowing air through the molten metal,3 removing bismuth and any
remaining copper. This returns to the copper smelter. At least one
refinery performs a final cleanup using Portland cement, which returns
to the Dore furnace at the start of the next charge.
The Dore" metal, that remains may be sold to a specialty processor,
or may be further refined (see Process No. 29). Table 2-35 gives the
approximate range of analysis.
Table 2-35. DORE METAL ANALYSIS2
Gold
Silver
Copper
Palladium
Platinum
Lead
Tellurium
Selenium
8 to 9%
90 to 92%
0.5 to 1.0%
0.16 to 0.18%
0.05 to 0.009%
0.02%
0.003%
0.00002%
2. Input Materials - Filter cake from the water leach process is the
primary input.The slimes at this stage are fairly low in copper con-
tent; they contain about 18 percent water2 and no sulfur.
Silica sand is the first flux, and a 2:1 mixture of sodium car-
bonate and sodium nitrate is the second. Quantities depend on the
analysis of the filter cake.4 Portland cement is used in very small
quantities.2
3. Operating Conditions - Temperatures in the furnace rise as high as
1400°C. Pressures are atmospheric.1
125
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4. Utilities - Gas or oil is the fuel to a Dore* furnace J>2 Com-
presseel air is used in the third stage slagging operation. Electricity
is not required except for auxiliary purposes.
5- Waste Streams - Flue gas temperatures may reach 1370°C.2 The
stream may be high in particulate matter and in fumes containing sele-
nium, tellurium, and some arsenic, antimony, and lead. These are
normally sent to a wet scrubber (Process No. 26). The precious metals
content of the particulate matter is high enough that care is taken to
collect them, but no analyses have been reported.
This process produces no solid wastes if all the slags are pro-
cessed or recycled as outlined above. The soda slag would be especially
troublesome if it became a solid waste, since it is rich in soluble
oxidized salts of arsenic, antimony, tellurium, and selenium.
There are no liquid wastes from the Dore* furnace.
6. Control Technology - Wet scrubbing of the gases for removal of
particulates and fumes is the best control of this gas stream.
Care should be taken in the handling of slags from the Dore*
furnace to avoid secondary water pollution from this source.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Leigh, A.M. Precious Metal Refining Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 95-110.
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
4. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
126
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PRIMARY COPPER PRODUCTION PROCESS NO. 26
Scrubber
1. Function - Gases from the si imes roaster and the Dore" furnace
contain particulates in quantities that justify their recovery for
further processing. The gases also contain fumes, especially of sele-
nium, which hydrolyze in the water scrubber, allowing their separation
for sale.
The scrubbers are generally of the water spray typej with the
water continuously recirculating. As solid material accumulates, peri-
odic blowdown is performed. The amorphous selenium is often removed by
flotation,^ or occasionally the blowdown is combined with the soda slag
leach liquor (Process No. 27). Muds from the scrubber are recycled to
the slimes roaster (Process No. 23).
2. Input Materials - Flue gases from the Dore" furnace and the slimes
roaster are the principal inputs.
If flotation recovery of selenium from the blowdown is practiced,
methylamyl alcohol and liquid colloid glue are used as flotation rea-
gents. 2
3. Operating Conditions - The gases entering the scrubber are ex-
tremely hot, about 1000° to 1300°C. The water sprays are at ambient
temperatures.^ Pressures are near atmospheric.
4. Utilities - Water is used as make-up to replace evaporation losses,
and electricity is used to drive the exhaust blower. Quantities are not
large.
5. Waste Streams - Gases leaving the scrubber may contain particulates
and fumes that were not removed. Selenium is expected to be a major
constituent.
If all the scrubbing liquor and particulates are recycled to pre-
vious operations, no liquid or solid wastes are produced.
6. Control Technology - Most refiners find it economical to install
electrostatic precipitators on the scrubber effluent to remove the
highly metalliferous dusts and fumes that escape collection. The use of
a more efficient venturi-type scrubber would also be an acceptable
control.
7. EPA Source Classification Code - None
127
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8. References -
1. Development Document for Interim Final Effluent limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
2. Leigh, A.M. Precious Metals Refining Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 75-110.
3. Particulate and Sulfur Dioxide Emission Control Cost Study of
the Electric Utility Industry. U.S. Environmental Protection
Agency. Washington, D.C. 68-01-1900.
128
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PRIMARY COPPER PRODUCTION PROCESS NO. 27
Soda Slag Leach
1. Function - The soda slag, the second slag removed from the Dore*
furnace, is rich in selenium and tellurium, both of which are marketable
by-products.1 The function of this leach process is to selectively
dissolve these elements from the slag.2
The slag is leached in a tank of water, which becomes alkaline be-
cause of the sodium oxide content of the slag. Selenium and tellurium
dissolve as sodium selenite and tellurite.3 The resulting solution is
filtered from the insoluble components of the slag, and the solids are
returned to the Dore" furnace for reprocessing. The leached solution is
further treated (see Process No. 28).
2. Input Materials - The soda slag from the Dore" furnace is the only
input.
3. Operating Temperature - Residual heat in the slag and chemical
action between the slag and the water cause some temperature increase
during leaching, to less than 100°C.2»4 Pressures are atmospheric
during the leach, and less than 2 kg/cm2 during filtration.
4. Utilities - Water is required as the leaching solvent.
The literature does not state whether supplemental heat is required
during this step.
Electricity in small quantity is used to pump the leached slurry
through the filter.
5. Waste Streams - There are no gas, liquid, or solids wastes from
this process. All materials at this stage are valuable and are care-
fully handled.3
6. Control Technology - None is required.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Leigh, A.M. Precious Metal Recovery Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 95-110.
129
-------
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
4. Hallowell, O.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
130
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PRIMARY COPPER PRODUCTION PROCESS NO. 28
Selenium and Tellurium Recovery
1. Function - The processing solutions which have become rich in
selenium and tellurium are treated in laboratory-scale equipment to
recover these elements as by-productsJ Both are valuable for use in
manufacture of electrical and electronic products.
The alkaline solutions are made acidic with sulfuric acid to a pH
of 5.5 to 6.5.2,3 Tellurous acid (^TeOa) precipitates, and is removed
by filtration. Then the solution is treated by bubbling S02 through it.
Selenium and any remaining tellurium precipitate in elemental form, and
can be selectively separated by several stages of precipitation and
filtration. Both are dried to become marketable products, or they may
be further purified prior to sale.
The crude tellurous acid is dissolved in caustic, treated with
sodium sulfide to precipitate impurities, and filtered. The clear
solution is again acidified, and the pure tellurous acid again pre-
cipitates. When filtered and dried, it can be sold in this form or may
be further processed to elemental tellurium.
A number of purification and reduction processes are used to pro-
duce pure materials. All are very small-scale operations.
Only a very small percentage of the tellurium in the original
copper ore is reclaimed. Ninety percent is lost during ore flotation,
while in each subsequent processing step, from 20 to 60 percent of the
tellurium that remains is lost.4 Selenium recovery is reported to be
much higher - recovery of 80 percent, with the remainder lost to slags,
flue dusts, and gases.4
2. Input Materials - The principal input is the filtered solution from
soda slag leaching (Process No. 27), to which are added selenium and
tellurium extracted or floated from scrubber precipitates (Process
No. 26).
Laboratory-grade reagents are normally used, such as sulfuric acid,
sodium hydroxide, compressed and liquified SOp, and others. Quantities
are very small.
3- Operating Conditions - Temperatures may reach 450°C during some
purification steps. Pressures are normally atmospheric.
4. Utilities - A variety of laboratory utilities may be employed.
Consumption of each is negligible.
131
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5. Waste Streams - No gas or solid wastes are generated by this pro-
cess in anything other than trace amounts.
Liquids are low in volume and normally discharge through standard
laboratory waste systems.
6. Control Technology - None is applicable.
7. EPA Source Classification Code - None
8. References -
1. Leigh, A.H. Precious Metal Recovery Practice (Presented at
International Symposium on Hydrometallurgy. Chicago, Illi-
nois. February 25 - March 1, 1973). p. 95-110.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
4. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
132
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PRIMARY COPPER PRODUCTION PROCESS NO. 29
Pore" Metal Separation
1. Function - In a series of complex chemical and electrochemical
laboratory operations, the Dore* metal, a mixture primarily of silver,
gold, and the platinum group metals, is separated into specification
grades of each of these metal sJ
A special small electrolytic cell, the Moebius cell, is used to
separate the silver, which is further processed to produce bullion bars
of 1000 troy ounces each, analyzing 99.97 percent silver.2*3
Mud from the Moebius cell is melted into anodes and processed in
another special electrolytic device, the Wohlwill cell, which produces
gold of marketable quality.
The remaining electrolyte is chemically processed to separate
platinum, palladium, and occasionally other metals. Iridium, rhodium,
ruthenium, and others may be present.
2. Input Materials - The principal input is Dore* metal (see Process
No.
Small quantities of many inorganic chemicals are used. The list
includes sulfuric, nitric, and hydrochloric acids, powdered iron and
copper metals, and sulfur dioxide.
3. Operating Conditions - Temperatures during the various steps of
processing range up to 1300°C in the casting of the metals, but most
operations are at less than 100°C. No unusual laboratory pressures are
employed.
4. Utilities - Electricity is used for the electrochemical operations,
and either electricity or gas for operation of the casting furnace.
Utility consumption is negligible in comparison to other processes in
this industry.
5. Waste Streams - Also in comparison to other processes, wastes are
insignificant. No" losses of unusual metallic elements occur in this
process. There are minor evolutions of nitrous oxides, sulfuric acid
mists, and other acid fumes, and occasional liquid discharges of elec-
trolyte acids in quantities of a few gallons at most. No solid wastes,
are produced, although residues may occasionally be returned to the
Dor6 furnace.
6. Control Technology - Local ventilation is the only control exer-
cised for this process.
133
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7. EPA Source Classification Code - None
8. References -
1. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Primary Copper Smelting Subcategory and the Primary Copper
Refining Subcategory of the Copper Segment of the Nonferrous
Metals Manufacturing Point Source Category. Environmental
Protection Agency, EPA 440/1-75/032-b. February 1975.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Leigh, A.M. Precious Metal Refining Practice (Presented at
the International Symposium on Hydrometallurgy. Chicago,
Illinois. February 25 - March 1, 1973). p. 95-110.
134
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PRIMARY COPPER PRODUCTION PROCESS NO. 30
Heap and Vat Leaching
1. Function - Heap and vat leaching are simple forms of hydrometal-
lurgy, in which valuable metals are dissolved from ores to form water
solutions. In heap leaching, the ores are placed in a dump, or pile, on
the ground. In vat leaching, they are placed in tanks, with or without
pretreatment.
Heap leaching is an accelerated form of natural weathering. It is
usually applied to low-grade ore and mine overburden that contains less
than 0.4 percent copper.'.2,3 This material is placed in an area pro-
vided with drainage ditches and basins, and is alternately flooded with
sulfuric acid solution and allowed to drain. This procedure causes
rapid oxidation of the copper minerals. Soluble copper sulfate is
formed, and washes from the heap with the acid solution. From 70 to 82
percent of the copper in these low-grade ores can be recovered.3»4 The
liquor that seeps from the heap has a pH of 1.5 to 2.55 and may contain
from 1.0 to 18 grams of copper per liter.4*5
Vat or percolation leaching is routinely applied to oxidized copper
minerals, which occur as partially weathered deposits in the mine.
These minerals cannot easily be reclaimed by conventional smelting
processes, so they are selectively mined and are placed in concrete
vats. They,are also subjected to alternate flooding with sulfuric acid
and draining.
The technique of hydrometallurgy allows the extraction of copper
from low-grade ores without evolution of sulfur dioxide. These can be
small operations that require less capital expenditure than pyrometal-
lurgical processing. The overall cost to produce a ton of copper is
greater, however, and there is no way in simple leaching to recover the
precious metals content of the ores.
One proprietary process is developed for vat leaching of a roasted
ore (see Process No. 34).
2. Input Materials - The principal input is the ore materials, as
described above. In most cases, these would otherwise be waste mate-
rials, unprofitable to process by conventional techniques.
Sulfuric acid has been the only solvent used for simple leaching
since it is not only inexpensive and nonvolatile, but also has a slight
selective action for copper. Consumption will vary, but to extract a
metric ton of copper from a 1 percent ore body containing oxide, car-
bonate, sulfate, or silicate minerals will require about 4400 liters of
96 percent acid.4
135
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3. Operating Conditions - Since the process occurs in an open, outdoor
area, it operates at atmospheric pressure and ambient temperatures.
4. Utilities - Diesel fuel and electricity are used in the materials
handling operations, and electricity in pumping the leach solution.
Process water must be added to most of these operations, since in this
country they are located in arid regions with high evaporation losses.
In 1973, water usage in heap leaching ranged from 920 to 4850 cubic
meters per metric ton of crude copper precipitate produced.4 For vat
leaching, consumption was 50 to 200 cubic meters of water per metric ton
of copper precipitate.4
5. Waste Streams - Wastes from heap leaching include fugitive dusts
from materials handling, and quantities of highly mineralized solid
wastes containing residual sulfuric acid. It is usually difficult to
separate these wastes from those of the mining process, as discussed in
more detail in Process No. 1.
Vat leaching produces a large amount of tailings of waste rock that
is sluiced into a tailings pond. This material is comparable to the
waste from a concentrator plant (Process No. 2). Frequently the same
pond is used for both concentrator and vat-leached tailings.
The circulating stream of a leaching operation eventually becomes
so rich in iron that it must be discarded. No analyses have been
reported; the volume is reported as varying from 350,000 to 1,000,000
liters of spent liquor per metric ton of copper produced.6
6. Control Technology - Most installations mix the discharge of this
process with mining or concentrating wastes. Control is described in
connection with Process Nos. 1 and 2.
7. EPA Source Classification Code - None
8. References -
1. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
2. Copper Hydrometallurgy: The Third-Generation Plants. Engi-
neering and Mining Journal. June 1975.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
136
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4. Roberts, R.W. San Xavier Vat Leach Plant Operation. Mining
Congress Journal. December 1974.
5. Gardner, S.A. and Warwick, G.C.I. Pollution-Free Metallurgy:
Copper via Solvent-Extraction. Engineering and Mining Jour-
nal. April 1971.
6. Davis, W.E. National Inventory of Sources and Emissions:
Copper, Selenium, and Zinc. U.S. Environmental Protection
Agency, (NTIS), Research Triangle Park, North Carolina.
PB-210 679, PB-210 678, and PB-210 677. May 1972.
137
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PRIMARY COPPER PRODUCTION PROCESS NO. 31
Cementation
!• Function - The cementation process converts soluble copper into a
metallic precipitate through a chemical reaction with metallic iron. It
is used to recover copper from strong solutions created by other pro-
cesses, especially those from heap and vat leaching.
This process is based on the relative activity of a metal to become
a soluble ion. Metals can be listed in a continuous electromotive
series; iron, having higher activity, will preferentially replace a
copper ion in solution and thus produce an insoluble precipitate, often
called "cement copper". In a typical application, liquor draining from
a heap leaching operation flows through a trough that is filled with
scrap iron. Part of the copper precipitates, and the liquor is recycled
back to the heap. It is reported that 94 percent of the copper can be
recovered by this methodJ»2
Periodically, the trough is cleaned and the cement copper is sent
to a smelter for processing. This is usually a mixture of copper with
iron compounds and other insoluble minerals. It usually is around 70
percent copper and is rarely more than 90 percent.'>3
In many of its applications, cementation is being replaced by
solvent extraction and electrowinning techniques (see Process No's. 32
and 33).
The term cementation is also applied in this industry to other
similar chemical reactions. Zinc metal is used in cementation of gold
and copper, and copper powder is used in cementation of silver.4
2. Input Materials - Aqueous liquors containing dissolved copper are
the principal input. The process is efficient only with fairly concen-
trated solutions.
Scrap iron is most commonly used for cementation if it can be
obtained. Because it is becoming difficult to obtain sufficient scrap
of good quality, a process for manufacture of a sponge iron for this
process is in the final steps of development (see Process No. 35).
3. Operating Conditions - Cementation processes normally operate at
atmospheric pressure and ambient temperatures.
4- Utilities - No utilities are consumed unless special pumps are
required to cause the liquor to flow through the cementation tanks.
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5. Waste Streams - Atmospheric pollution from the cementation process
is negligible.TFere may be tiny amounts of hydrogen gas created by a
side reaction of acid with the iron.
No liquid wastes can be directly assigned to this process. In most
of its applications, cementation is used as an auxiliary to a larger
liquid handling operation.
Solid wastes may be assigned to cementation by the presence of
scrap iron partially used, discarded, or abandoned, that causes some of
these operations to resemble a junk yard.
6. Control Technology - No controls are specific to this process.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Roberts, R.W. San Xavier Vat Leach Plant Operation. Mining
Congress Journal. December 1974.
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
4. Development Document for Interim Final and Proposed Effluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source
Category Volumes I and II. Environmental Protection Agency,
Washington, D.C., EPA/1-75/032-6. February 1975.
139
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PRIMARY COPPER PRODUCTION PROCESS NO. 32
Solvent Extraction
!• Function - As applied in the copper industry, solvent extraction is
a method to produce a concentrated copper solution, relatively free of
other metal ions, from a solution of copper that does contain other
dissolved metals. The process uses a special mixture of organic sol-
vents and an agitated vessel. When the solvents are mixed with the
impure solution, the copper combines with the solvents as a complex.
Agitation of the vessel is then stopped and the solvents, now containing
the copper, form a separate layer. The water layer is drained off.
Sulfuric acid is then added and mixed with the solvent. This breaks
down the complex and regenerates the solvent for reuse. The copper is
withdrawn as a solution in the acid.
Solvent extraction has been applied to liquors from vat leaching,
and it is being incorporated into some of the developing hydrometallur-
gical processes^ (see Process No. 36). The concentrated acid solution
can be directly treated by electrowinning (see Process No. 33). This
process can also be made continuous, rather than batch, to adapt it to
large scale operations.
It is reported that about 95 percent of the copper values in a
solution can be extracted by this technique.2
2. Input Materials - Water solutions of copper are the primary input.
There is no published information on the composition ranges that can be
efficiently treated with this process.
The solvents mixture is kerosene containing about 12 percent of a
proprietary chemical made by General Mills called LIX.3»4 The total
rate of recycle is not known, but losses of 0.1 liter per 1000 liters of
impure solution have been reported.4 TWO other chemicals, "Kelex" and
"Shell 529" are also being advertised for this application.
Concentrated sulfuric acid is required (normally recycled through
electrowinning cells), but no usage quantity has been reported.
3. Operating Conditions - No special temperature limits are reported;
it is assumed that ambient temperatures and atmospheric pressures are
satisfactory.
4. Utilities - A small amount of electricity is required for agitation
and liquid pumping.
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5. Waste Streams - There are no reports of atmospheric pollution from
this process. For operating safety, evaporation of the solvent is
undoubtedly minimized.
The manufacturer of the LIX solvent states that small amounts of
iron, arsenic, and zinc are extracted along with the copper.1 The
procedure for disposal of these materials has not been reported. It is
likely that there will be a bleed of the concentrated acid to prevent
accumulation of these other elements.
The loss of solvent is reported as 1 liter per 10,000 liters of
raffinate.2 It is likely that this is almost entirely kerosene, which
has a slight water solubility. The more expensive chelating compounds
should stay largely dissolved in the kerosene layer. No confirmation of
this has been published.
No solid wastes are generated by this process.
6. Control Technology - No special controls are indicated. The
possible acid blowdown should be of a quality that could be reused in
other processes. The organic loss would be biodegradable if this waste
stream were combined into other wastewaters.
7. EPA Classification Code - None
8. References -
1. In Clean-Air Copper Production^ Arbiter is First off the Mark.
Engineering and Mining Journal. 1973.
2. Gardner, S.A. and Warwick, G.C.I. Pollution-Free Metallurgy:
Copper via Solvent-Extraction. Engineering and Mining Jour-
nal. April 1971.
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
4. Ion Exchange: The New Dimension in Copper Recovery Systems.
Engineering and Mining Journal. June 1975.
141
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PRIMARY COPPER PRODUCTION PROCESS NO. 33
Electrowinning
1. Function - Electrowinning is a process for the extraction of re-
latively pure copper metal from a solution containing copper ions. This
is an electrolytic process, similar to the cells of an electrolytic re-
finery (see Process No. 18), except than an inert anode is used. Copper
metal deposits at the cathode, and the water in the solution is electro-
lytically decomposed, liberating oxygen at the cathode, and regenerating
the sulfate ion as sulfuric acid.
If the copper solution is relatively pure, the copper produced by
electrowinning is comparable to the best electrolytic copper, assaying
99.9 percent plus.' If impure solutions direct from vat leaching are
used, the purity is equivalent to that of the anode copper from a
conventional smelter and the product thus requires electrolytic refining
prior to sale.
2. Input Materials - Electrowinning is in use to recover copper direct
from vat leaching solutions (Process No. 30), and from purified solution
from solvent extractions (Process No. 32). It is also being tested as a
part of some of the more sophisticated hydrometallurgical processes that
are being developed, in which chlorides rather than sulfates will be the
input materials (see Process No. 36).2
Additives to produce a uniform cathode deposit are necessary. They
are the same as for electrolytic refining. One report lists glue for
electrowinning being added at a rate of 0.02 to 0.06 kilograms per
metric ton of cathode copper.3
3. Operating Conditions - Electrowinning cells are maintained at about
60 to 65°C and at atmospheric pressure.2>3
4. Utilities - Electrowinning requires 8 to 10 times as much electric
current as does an electrolytic refining cell to produce the same amount
of copper.J In a loop with a solvent extraction process, 2100 kg-cal
are required to produce a kilogram of copper.4 in direct electrowinning
of a vat or heap leach solution, 2400 kg-cal per kilogram of copper are
required.•* These high values reflect the energy required to dissociate
water into its elements and are the principal reason that the simple
hydrometallurgical processes have been more expensive than conventional
smelting.
A small amount of water is used to clean the cathodes after re-
moving them from the cell.
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5. Waste Streams - The oxygen produced at the anode of an electro-
winning cell can be considered either as an atmospheric emission or as a
by-product. If it is discarded to the atmosphere, there are no delete-
rious environmental effects.
A small amount of liquid waste may be discharged in connection with
cleaning of the completed cathodes. No reports of this source have been
published.
The possible larger purge of electrolyte, necessary to prevent
accumulation of other elements, was discussed in connection with Process
Nos. 30 and 33.
There are no solid wastes from this process.
6. Control Technology - No special controls are applicable to the
minor wastewater that may develop from this process.
7. EPA Source Classification Code - 3-03-005-0
8. References -
1. Gardner, S.A. and Warwick, G.C.I. Pollution-Free Metallurgy:
Copper via Solvent-Extraction. Engineering and Mining Jour-
nal. April 1971.
2. Atwood, G.E., and Curtis, C.H. Hydrometallurgical Process for
the Production of Copper. U.S. Patent No. 3,785,944. January
15, 1974.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
4. Ion Exchange: The New Dimension in Copper Recovery Systems.
Engineering and Mining Journal. June 1975.
143
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PRIMARY COPPER PRODUCTION PROCESS NO. 34
Sulfation Roasting
1. Function - One company has developed a hybrid process that will use
a fluidizatlon roaster (Process No. 4) to prepare a calcine especially
suited to vat leaching (Process No. 30). This is the technique of
sulfation roasting. In this process, concentrate is roasted to oxidize
the copper to copper sulfate and the iron to iron oxide, and to remove
the excess sulfur as sulfur dioxide. Roasting is a batch rather than a
continuous operations.
2. Input Materials - This plant will use an ore concentrate that is
predominantly chalcopyrite (CuFeS2).l It is expected than any sulfide
concentrate could be used. The concentrate is blended with Fe203,
which moderates the heat produced by the exothermic oxidation reaction
and promotes sulfation of the copper. A ratio of two parts concentrate
to one of iron oxide is believed to be about optimum.2
3. Operating Conditions - Temperatures are kept much lower than in
conventional roasting. Instead of 760°C, the range is 400 to 600°C.2>3
Pressures are approximately atmospheric.
4. Utilities - Gas or oil is used to preignite the charge and to
maintain temperature.
Noncontact cooling water is used to regulate temperatures of the
roaster.
Air or oxygen is injected through the bottom of the charge for
oxidation. Twenty percent above theoretical amount of oxygen will be
required for the duration of each batch.2
5. Waste Streams - It is believed that emission of metallic fumes will
be considerably less than in a conventionally operated roaster. Par-
ticulate emissions following the roaster cyclones have not been esti-
mated.
Organic flotation reagents are expected to be volatilized into the
exit gases and oxidized. The gas stream is expected to contain 8 percent
S02 and 4 percent oxygen.4 Gas temperature should be less than 400°C.
There will be no solid or liquid wastes from this process.
6. Control Technology - A 225 metric ton per day single-contact sul-
furic acid plant will remove from 70 to 80 percent of the S02 from this
process.4,5 This operation will require complete particulate removal,
but it is not known what devices will be used.
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7. EPA Source Classification Code - 3-03-005-02
8. References -
1. Haver, P.P. and Wong, M.M. Lime Roast-Leach Method for Treat-
ing Chalcopyrite Concentrate. U.S. Bureau of Mines, Washing-
ton, D.C. 8006. 1975.
2. Foley, R.M. Method of Treating Copper Ore Concentrates. U.S.
Patent No. 2,783,141. February 26, 1957.
3. Haskett, P.R., Bauer, D.J., and Lindstrom, R.E. Copper Re-
covery from Chalcopyrite by a Roast-Leach Procedure.
4. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
5. Potter, J. Personnel Communication on Hydrometallurgical
Processes. Bureau of Mines. Salt Lake City, Utah. 1976.
145
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PRIMARY COPPER PRODUCTION PROCESS NO. 35
Sponge Iron Plant
!• Function - One company is building a sponge iron plant to accompany
their sulfation roasting installation (Process No. 34).' After leaching
of the copper from the calcine produced by sulfation roasting, the leach
residue contains mostly iron oxide plus smaller amounts of copper and
precious metals. This process will partially reduce the iron, which
will then be used for cementation of liquor from the vat leaching of
oxidized copper ores. The precipitate from this process is expected to
contain the precious metals from the concentrate originally fed into the
sulfation roaster,2 and will therefore provide a means of recovery.
The sponge iron will be produced in a kiln by reduction with coal.
The iron is not to be high-purity grade, but will be adequate for cemen-
tation.
2. Input Materials - The principal input is the residue from vat
leaching of sulfation roasted concentrates.
Coal is to be used in the proportion of one ton of coal for each
two tons of sponge iron produced.
3. Operating Conditions - Kiln temperatures are expected to be ap-
roximately 1100°C.^Pressures are approximately atmospheric.
4. Utilities - Gas or oil is used to heat the charge until the coal is
ignited, and is then used only if required to maintain temperature.
Combustion air is allowed to enter the kiln in carefully regulated
amounts. Air quantity is calculated to be 1.5 tons of air per ton of
iron produced.3
5. Waste Streams - No emission data are available, since this process
is not yet in operation. Particulates and fumes of volatile metals
would be expected, in a gas containing appreciable carbon monoxide.
The process will generate a solid waste in the form of a slag.
There should be no liquid waste.
6. Control Technology - It is not yet known what atmospheric control
devices will be employed with this process.
The slag is expected to be discarded in a waste dump also used for
wastes from ore-concentrating operations.
146
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7. EPA Classification Code - None
8. References -
1. Potter, 0. Personnal Communication on Hydrometallurgical
Processes. Bureau of Mines. Salt Lake City, Utah, 1976.
2. Hydrometal1urgy Makes Advances in Copper Processing. Engi-
neering and Mining Journal. 1973.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
147
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PRIMARY COPPER PRODUCTION PROCESS NO. 36
CLEAR Reduction
1. Function - An advanced hydrometallurgical process has been devel-
oped with the trade name of CLEAR (Copper Leaching, Extraction, and
Refining). The descriptions in Process No's. 36 through 38 outline the
three sections of this method.
The first step of the CLEAR system, called a reduction, could also
be called a leaching operation. It uses as a solvent a water solution
of cupric, ferrous, and sodium chlorides.! Primarily the cupric chlo-
ride is active during this step, reacting with copper ore minerals to
form cuprous chloride, additional ferrous chloride, and elemental sul-
fur. The sulfur is a solid material and remains with the leach residue.
Since only about half the copper in the concentrate is solubilized in
this first leach, it is treated a second time in the oxidation process
(Process No. 38).
Although in this process, leaching of fresh concentrate does occur,
the principal purpose of this step is to prepare the liquor for electro-
winning. It may be reduced with other materials if necessary, then
filtered and sent to electrowinning eel Is.2 These cells are similar to
those described in Process No. 33, but are slightly modified for chloride
service, operating at a slightly lower temperature, around 55°C. After
electrowinning removes part of the copper, the liquor is sent to the
Regeneration-Purge step (Process No. 37).
2. Input Materials - A typical Arizona ore concentrate is the primary
input. Chalcopyrite is the predominant ore mineral.
The leach liquor at this stage contains about 8 percent cupric
chloride, 12 percent ferrous chloride, and 13 percent sodium chloride.'
It is received directly from the oxidation leach of the previous batch.
To complete the reduction of the liquor, scrap iron or copper,
sodium sulfite, or sulfur dioxide may be added.2
3. Operating Conditions - The CLEAR process operates at higher tem-
peratures than some other hydrometallurgical processes; 107°C has been
reported.1 Pressures are atmospheric.
4- Utilities - Although there are no published reports, the source of
heat is probably steam. Electricity is also undoubtedly required for
materials handling and pumping.
148
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5. Waste Streams - Some loss of hydrochloric acid vapor may occur, but
the leaching step is enclosed to minimize this emission. Dust may arise
from materials handling.
The process generates no intentional waste streams; with the corro-
sive solutions, however, accidental losses of liquids are likely.
6. Control Technology - Any losses of hydrochloric acid vapors can be
controlled by scrubbing with an alkaline solution.
7. EPA Classification Code - None
8. References -
1. Atwood, G.E., and Curtis, C.H. Hydrometallurgical Process for
the Production of Copper. U.S. Patent No. 3,785,944. January
15, 1974.
2. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometal-
lurgy. Chemical Engineering. Volume 83, No. 1: pp. 79-81.
January 5, 1976.
149
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PRIMARY COPPER PRODUCTION PROCESS NO. 37
CLEAR Regeneration — Purge
1. Function - Exhausted and stripped solvent from the electrowinm'ng
cells that follow the reduction process (Process No. 36) is oxidized
with air to remove excess iron and to prepare the solution for the
pressurized leach operation (Process No. 38). Air is blown through the
solution, and ferrous iron is oxidized to ferric chloride and ferric
hydroxide. Copper in the solution acts as a catalyst for this oxida-
tionJ Sulfates are also formed and collect into an insoluble compound
similar to the mineral jarosite.2. The jarosite and ferric hydroxide
are filtered out and discarded.
2. Input Materials - Liquor from the electrowinm'ng cells is the only
input. At this stage, the solution contains about 6 percent cuprous
chloride, 8 percent ferrous chloride, and 14 percent, sodium chloride.3
3. Operating Conditions - This is a pressurized process, operating at
107°C, and 2.7 kg/cm* (40 psig).3
4. Utilities - Source of heat has not been reported. Either steam or
direct firing could be applicable.
Compressed air is required, and process water is added at this step
to compensate for evaporation and losses. Quantities are unknown.
5. Waste Streams - Although no data have been reported, a gaseous
stream must be released from this step, carrying hydrochloric acid vapor
and steam.
Solids removed by filtration are discarded. Composition is reported
to be primarily an iron sulfate/hydroxide mixture.3 Other elements
leached from the concentrate will be present. The quantity of solids
produced is about 2 to 4 percent of the total weight of spent electro-
lyte.-3
Liquids will probably drain to waste from the solids.
6- Control Technology - The gas stream from this oxidation operation
is undoubtedly processed to recover the vaporized and entrained mate-
rials. This is probably accomplished by external cooling, condensation,
and water scrubbing. No details have been disclosed.
The solid wastes from this process are mixed with the wastes from
the oxidation step (Process No. 38) and sluiced into a settling pond.
In locations other than the arid region where this plant is operating,
secondary water pollution could be substantial.
150
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7. EPA Classification Code - None
8. References -
1. Potter, J. Personnal Communication on Hydrometallurical
Process. Bureau of Mines. Salt Lake City, Utah, 1976.
2. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometal-
lurgy. Chemical Engineering. Volume 83, No. 1: pp. 79-81.
January 5, 1976.
3. Atwood, G.E., and Curtis, C.H. Hydrometallurgical Process for
the Production of Copper. U.S. Patent No. 3,785,944. January
15, 1974.
151
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PRIMARY COPPER PRODUCTION PROCESS NO. 38
CLEAR Oxidation
1. Function - This is the principal leaching operation of the CLEAR
system, in which partially leached ore concentrate is contacted with
freshly regenerated leach solution. Primarily the ferric chloride is
active during this step, reacting with copper ore minerals to form
ferrous chloride, cupric chloride, and elemental sulfur.
The sulfur is a solid material that remains with the leach residue.
It is reported that 98 percent of the residual copper is extracted in
this step.1 The solids from this process are discarded, although
cyanide treatment for gold recovery may be performed if the gold analysis
warrants it (see Process No. 2).
The leach solution is filtered and sent to the reduction step
(Process No. 36).2
2. Input Materials - Solid residue from Process No. 36, and liquor
from Process No. 37 are the only input materials. The ratios have not
been disclosed. The liquor at this stage contains about 6 percent
cupric chloride and 15 percent each of ferric and sodium chlorides.'
3. Operating Conditions - This is a high-temperature, pressurized 2
leaching operation.Temperatures of 140°C and pressures of 2.7 kg/cm
are used.1*
4. Utilities - Source of heat has not been reported; either steam or
direct firing could be applicable. Electricity is also undoubtedly
required, but again no information has been published.
5. Waste Streams - Details of the process have not been disclosed in
sufficient detail to establish whether emissions of gases to the atmos-
phere occur. Hydrochloric acid vapors could be generated.
The major waste of the CLEAR system, consisting of the solid
residue, is discharged from this step. This residue may be a large
fraction of the original concentrate. It is expected to be 99 percent
free of copper' and much reduced in iron; it must contain considerable
colloidal sulfur and soluble chlorides. It may contain cyanides if gold
extraction was performed. No analyses have been published.
6. Control Technology - The process will likely incorporate an opera-
ting control of hydrochloric acid emission, since even small concentra-
tions of this very corrosive gas can damage plant equipment. If needed,
scrubbing with an alkali can further reduce the concentration.
152
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Solid wastes from this process are sluiced into a settling pond.
The location of this first application is such that natural evaporation
should dispose of the water content, and secondary water pollution
should be minimal. In other locations, however, this waste could cause
severe secondary pollution in the form of an acidic seepage high in
chlorides, sulfates, and heavy metals.
7. EPA Source Classification Code - None
8. References -
1. Atwood, G.E., and Curtis, C.H. Hydrometallurgical Process for
the Production of Copper. U.S. Patent No. 3,785,944. January
15, 1974.
2. Potter, J. Personnel Communication on Hydrometallurgical
Process. Bureau of Mines. Salt Lake City, Utah, 1976.
3. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometal-
lurgy. Chemical Engineering. Volume 83, No. 1: pp. 79-81.
January 5, 1976.
153
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PRIMARY COPPER PRODUCTION PROCESS NO. 39
Cymet Leaching
1. Function - An advanced hydrometallurgical process is being devel-
oped with the trade name of Cymet. The descriptions in Process Nos. 39
through 42 represent best guesses about this method as it now exists. A
pilot plant was built in 1974 but was never started. In 1975, substan-
tial changes were made, but no information was officially released,
pending patent coverages. It is not known whether the changes have been
successfully applied to the pilot plant, or whether plans to build a
full-scale process are still being considered.
The Cymet process is expected to be based on a leach using a solu-
tion of ferric chloride and hydrochloric acid. This solution will react
with minerals such as chalcopyrite to produce cuprous chloride, ferrous
chloride, and elemental sulfur.
The sulfur is a solid material that remains with the leach residue.
The solution contains copper and iron, reported to be 99 percent of the
copper from the concentrate and 84 percent of the iron.' It is also
reported to extract 60 percent of the gold content and 96 percent of the
silverj but the chemistry of this extraction has not been published.
The concentration of copper in the leach liquor is reported as 5.5
percent by weight.2
2. Input Materials - Copper concentrate is the primary input. It must
be milled to at least 90 percent minus 270 mesh, but for thorough copper
extraction it must be 100 percent minus 325 mesh.3
The leach solvent is approximately 36 percent ferric chloride and
1 percent hydrochloric acid in water. From 2 1/2 to 5 tons of leaching
solution must be added to each ton of concentrate.1
3. Operating Conditions - The leaching temperature is 70 to i>0°C,2»3
and pressures are atmospheric.4
4- Utilities - Steam is probably used to maintain leach temperature
and electricity for materials handling. Details of utilities have not
been published.
5. Waste Streams - Unless scrubbing devices are integral with the
process, there would be an emission of hydrochloric acid fumes at the
operating temperatures. Details have not been reported.
Solid residues from leaching are estimated to be almost 0.6 ton of
waste for-each ton of copper produced. Composition has not been
reported.
154
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Except for accidental spills, there will be no liquid wastes from
this process.
6. Control Technology - If hydrochloric acid fumes are created by this
process, they can be controlled by scrubbing with an alkaline solution.
Fumes of hydrochloric acid usually cannot be tolerated for purely
economic reasons, since they are very corrosive to equipment.
It is expected that the residues from this process will be slurried
to a tailings pond. In this first application, natural evaporation will
remove the water. Composition of the solids are at this time insuffi-
ciently characterized to permit control evaluation; some consideration
should be given to the elemental sulfur that will be present in these
tailings. There are no data to show whether this will form secondary
pollution on long standing.
7. EPA Source Classification Code - None
8. References
1. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometal-
lurgy. Chemical Engineering. Volume 83, No. 1: pp. 79-81.
January 5, 1976.
2. Druesi, P.R. Process for the Recovery of Metals from Sulfide
Ores through Electrolytic Dissociation of the Sulfides. U.S.
Patent No. 3,673,061. June 27, 1972.
3. Kruesi, P.R. Cymet Copper Reduction Process.
4. Copper Hydrometallurgy: The Third-Generation Plants. Engi-
neering and Mining Journal. June 1975.
5. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974
155
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PRIMARY COPPER PRODUCTION PROCESS NO. 40
Cyrnet Crystal 1ization
1. Function - It is believed that following the ferric chloride leach
in the Cymet process (Process No. 39), cuprous chloride is to be sepa-
rated from the leach liquors by crystallization. The developer of the
process has released no details, since this section could be the one on
which patent coverage is being sought.
It is known that cuprous chloride is less soluble in water than
ferrous chloride, and that if the solution were evaporated, crystals of
copper chloride would form preferentially on cooling. They might then
be separated by filtration. It is likely that the process is not this
simple, since in other industries it has been very difficult to grow
crystals to filterable size in similar mixtures of inorganic salts.
It is thought, however, that the end product of this process is a
dry solid composed primarily of cuprous chloride.1
2. Input Materials - The only known input material is the leach liquor
from Process No. 36.
3. Operating Conditions - No details have been released. A closed
process to minimize oxidation of ferrous iron is likely. High-vacuum
crystallization has been successfully applied to similar solutions.
4. Utilities - No details have been reported.
5. Waste Streams - Atmospheric emissions of acid are possible, and a
liquid waste stream is probable. No details have been reported. Solid
wastes are unlikely.
6. Control Technology - The process is insufficiently defined to
evaluate waste control technology, but losses of hydrochloric i»:id may
require additional equipment for recovery or treatment.
7. EPA Source Classification Code - None
8. References -
1. Potter, J. Personnal Communication on Hydrometallurgical
Processes. Bureau of Mines. Salt Lake City, Utah, 1976.
156
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PRIMARY COPPER PRODUCTION PROCESS NO. 41
Cymet Reduction
1. Function - In the overall Cymet hydrometallurgical process, cuprous
chloride could be converted to metallic copper by direct reduction with
hydrogen gas. This would also produce hydrogen chloride, which could be
absorbed in water to form hydrochloric acid. It is suspected that, in
Cymet, this reduction will take place in a fluidized bed, using the same
principle as the fluidization roaster (Process No. 4)J Most similar
direct hydrogenation processes require a catalyst for this reaction, but
it is possible that the copper product can act as its own catalyst.
This process would produce a finely divided copper powder that
would then be separated by a process not yet disclosed from the unreacted
copper chloride powder. The developer claims copper of high purity will
be produced. The technology by which to further purify or solidify the
product has not been announced.
2. Input Materials - It is assumed that dry cuprous chloride from
Process No. 40 is the primary input and that the only other input is
hydrogen gas. Quantities are not known. ,
3. Operating Conditions - No data have been released.
4. Utilities - No data are available, although fluidized bed processes
require sizeable amounts of electricity to circulate high-velocity gas
streams.
5. Waste Streams - It is possible that this process will not recover
all the hydrogen chloride gas as hydrochloric acid and may generate a
waste stream of this gas.
There should be no liquid or solid wastes.
6. Control Technology - Absorption of HC1 by scrubbing with an alkali
provides adequate control.
7. EPA Source Classification Code - None
8. References -
1. Potter, J. Personnel Communication on Hydrometallurgical
Processes. Bureau of Mines. Salt Lake City, Utah, 1976.
157
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 42
Cymet Sol vent Regeiierati on
1. Function - The designer of the Cymet process states that it is a
closed-loop system. One or more regeneration steps will therefore be
required to recycle the hydrochloric acid produced in Process No. 41 and
to oxidize the ferrous iron produced in Process No. 39 back to ferric
chloride. The system must also dispose of the iron that enters the
process with the concentrate, as well as the elemental sulfur produced
during the leach reaction.!
Details of these procedures have not been announced. Theoreti-
cally, if 25 percent of the liquor from the crystallization step is
treated to extract the HC1, that portion could be discarded as stable
ferric oxide or hydroxide. The remaining 75 percent would represent the
amount of iron that would have to be oxidized and combined with all the
reclaimed hydrochloric acid to keep the process operating with no loss
of the chlorine intermediate. This theoretical calculation makes the
assumption that the original concentrate was pure chalcopyrite.
2. Input Materials - Two process streams are involved: the mother
liquor from the crystallization step from Process No. 40 and the hydrogen
chloride or hydrochloric acid from Process No. 41.
3. Operating Conditions - They are not known.
4. Utilities - They are not known.
5. Waste Streams - The form of the wastes is not known. Iron in at
least a one-to-one molecular ratio to copper must be separated. Sulfur
must also be separated in at least a two-to-one molecular ratio.
6. Control Technology - No evaluation can be made.
7. EPA Classification Code - None
8. References -
1. Potter, J. Personnel Communication on Hydrometallurgical
Processes. Bureau of Mines. Salt Lake City, Utah, 1976.
158
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PRIMARY COPPER PRODUCTION PROCESS NO. 43
Arbiter Leaching
1. Function - An advanced hydrometallurgical process is being developed
with the trade-name of Arbiter. Processes Nos. 43 through 45 describe
the three sections of this process. Copper is leached from ore con-
centrates with ammonia in a strong oxidizing environment, forming soluble
copper ammonium sulfate and ferric oxide and hydroxideJ>2,3,4,5 jhe
leach solution is then sent to Process No. 44 for further treatment.
This leach is claimed to be able to extract as much as 99 percent of the
copper in the concentrate.3
2. Input Materials - This process is designed to leach any sulfide
mineral of copper, but chalcocite (Cu£S) requires less leach time than
chalcopyrite (CuFeS2) and energite (Cu3AsS4).
The leaching solvent is a water solution of ammonia and ammonium
sulfate. Gaseous oxygen is also required during the leaching step.
About 1.3 tons of oxygen are required to produce a ton of copper.
3. Operating Conditions - The Arbiter leach takes place at about 0.7
kilograms per square centimeter pressure and 70°C.5
4- Utilities - Steam is used to maintain leach temperature and elec-
tricity is required for pumping and auxiliary services.
5. Waste Streams - The leaching step of the Arbiter process creates no
air pollution.Alimonia that escapes from the leach tank is captured and
recycled.
Solid wastes are reported to be 1.8 tons of leached tailings per
ton of copper produced.6 During most of the testing of the Arbiter
system, only a portion of the copper has been extracted from the ore
concentrate, and the solids have been transferred to pyrometallurgical
processing for further treatment.
There are no liquid wastes from this process. All solutions are
further treated.
6. Control Technology - As this process has been practiced to date,
controls have not been required. In other applications, the leached
tailings, which do not contain elemental sulfur, can be discarded after
washing as are flotation tailings.
7. EPA Source Classification Code - None
159
-------
8. References -
1. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometal-
lurgy. Chemical Engineering. Volume 83, No. 1: pp. 79-81.
January 5, 1976.
2. Hydrometallurgy. Chemical Engineering. January 1976.
3. Price, F.C. Copper Technology on the Move. Chemical Engi-
neering.
4. In Clean-Air Copper Production, Arbiter is First off the Mark.
Engineering and Mining Journal. June 1975.
5. Copper Hydrometallurgy: The Third-Generation Plants.
Engineering and Mining Journal. June 1975.
6. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research
Triangle Park, North Carolina, EPA-450/2-74-002a. October
1974.
160
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PRIMARY COPPER PRODUCTION PROCESS NO. 44
Arbiter Precipitation
1. Function - Arbiter differs from other hydrometallurgical systems in
that the route to metallic copper is chemical rather than electrochemical.
In this step, solution from Process No. 43 is filtered, boiled to drive
off excess ammonia and treated with SQ^ gas. An insoluble copper ammonium
sulfite is formed, which when settled and filtered from the solution
contains up to 97 percent of the copper leached from the concentrate.
The crystals are sent to Process No. 45 for conversion to metallic
copper.''2
2. Input Materials - Liquor from the leaching step is the primary
input. Sulfur dioxide gas is required to acidify the liquor to a pH
between 3 and 5.
3. Operating Conditions - Crystal formation takes place at 40 to 55°C
and at approximately atmospheric pressure.2
4. Utilities - Non-contact cooling water is required to remove the
heat of reaction and electricity is used for solution pumping.
5. Waste Streams - The waste or by-product stream is produced from
this process, consisting of a water solution of ammonium sulfate con-
taining copper in concentration from 0.3 to 1.5 grams per liter, and
containing free sulfuric or sulfurous acid to a pH of less than 5.'
A second waste is produced consisting of the backwash and precoat
of the leach liquor filter. This is probably a slurry consisting
primarily of fine rock particles. No details have been announced.
There is no atmospheric emission from this process step. Ammonia
and S02 are released from process vessels, but are captured and recycled.
6. Control Technology - Unless all the waste liquor from this process
is recycled to the leach step, there is undoubtedly a recovery process
in use to salvage its copper content, since as much as 3 percent of the
product can remain in the filtrate. The remaining solution of ammonium
sulfate can be crystallized and sold as fertilizer, although the market
is limited, or could be hot-treated with lime to .salvage the ammonia,
and to form gypsum which would be discarded. There are reports that
this second method is the one being used in the operating plant with
either this stream or a mixed stream including a similar solution from
Process No. 45.
The filter backwash is probably mixed with the tailings from the
leach step. This practice should not add appreciably to waste loading.
161
-------
7. EPA Source Classification Code - None
8. References -
1. Arbiter, N., D. Milligan, and R. McClincy. Metal Production
from Copper Ammine Solution with Sulfur Dioxide. IChemE
International Symposium on Hydrometallurgy, Manchester,
England. 1975.
2. Arbiter, N. and D. Milligan. Reduction of Copper Ammine
Solutions to Metal with Sulfur Dioxide. International Sympo-
sium on Copper Metallurgy. American Institute of Mining
Metallurgical and Petroleum Engineers. Annual Meeting, Las
Vegas, Nevada. 1975.
162
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 45
Arbiter Decomposition
1. Function - Crystals of copper ammonium sulfite are autoclaved in
the presence of sulfuric acid, and decompose to form ammonium sulfate,
copper powder and S02 gas. The gas is vented from the autoclave, and
copper is filtered from the liquid. Conversion to copper metal is
essentially 100 percentJ
The copper powder may be sold in this form for certain applications
or may be melted and cast into wirebars. It is also reported that less-
pure powder can be made by less critical operation of the Arbiter
processes, which can then be fire-refined to commercial grade copper.2
2. InpuVMaterials - Crystals from Process No. 44 are the primary
input. Sulfuric acid is also used, but the literature does not com-
pletely define its application.
3. Operating Conditions - Temperatures in the autoclave are 147 to
150°C. Pressures are unreported, but are probably near 4 kilograms per
square centimeter.
4. Utilities - Steam and electricity are required in this step, but
details are unavailable.
5. Waste Streams - A solution of ammonium sulfate is produced by this
process, similar to the first-stage waste from Process No. 44 but con-
taining less soluble copper.
Sulfur dioxide is vented from the autoclave. One publication of
the developer2 shows this gas stream vented, not recycled. The chemistry
of the process, however, indicates that there is a net absorption of S02
by the Arbiter system and recycle of this stream is likely.
There are no solid wastes produced in this step.
6. Control Technology - The ammonium sulfate solution is shown on one
drawing of the developer as mixing with the similar solution from
Process No. 44.2 Control of this stream is the same as described in the
previous process.
Gas from this step should consist of only S02 and steam, and recycle
to Process No. 44 would be best control.
7. EPA Source Classification Code - None
163
-------
8. References -
Arbiter, N., D. Milligan, and R. McClincy. Metal Production
from Copper Ammine Solution with Sulfur Dioxide. IChemE
International Symposium on Hydrometallurgy, Manchester,
England. 1975.
Arbiter, N. and D. Milligan. Reduction of Copper Ammine
Solutions to Metal with Sulfur Dioxide. International Sympo-
sium on Copper Metallurgy. American Institute of Mining
Metallurgical and Petroleum Engineers. Annual Meeting, Las
Vegas, Nevada. 1975.
164
-------
3.0 LEAD INDUSTRY
INDUSTRY DESCRIPTION
The market for products of the lead industry continues to decrease,
principally because of public awareness that lead and its compounds are
cumulative poisons. Lead pigments are now rarely used in paints.
Although the manufacture of tetraethyl lead for gasoline additives
continues as a major market, its use is being restricted. In recent
years also, other materials have replaced lead in such applications as
joining material for cast iron pipe, and in plumbing and most other
construction applications.
Therefore, although new methods are being developed for lead pro-
duction, none has reached commercialization and lead is now being
produced by pyrometallurgical processes that have changed little in the
last 75 years. Only one major development during that period has been
adopted by the industry: updraft sintering machines have replaced almost
all the downdraft types.
The best news for some lead-producing companies came in 1965, when
the "New Lead Belt" of southeastern Missouri was discovered. Mining of
this deposit began in 1967 and it now produces more than 80 percent of
the ore that is mined in this country specifically for lead. Sections
of this deposit produce almost pure galena, analyzing more than 70
percent lead and containing only very small amounts of other metals.
Three of the six U.S. lead smelters are near this new lead belt in
Missouri. The others are in Idaho, Montana, and Nebraska. The industry
employs about 7000 people, with two-thirds in mining and concentrating.
Raw Materials
Lead is most often found in nature as galena (PbS2), the primary
sulfide of lead. These are rarely the pure deposits, since the lead-
bearing compound is usually mixed with pyrite, sphalerite, and pyrrho-
tite. Most of these deposits contain very little copper, gold, or
silver.
Oxidized lead ores also occur, composed primarily of anglesite and
cerussite, which are the weathered products of galena. Table 3-1 lists
the important lead ore minerals, together with others in which the lead
is combined with phosphorus, vanadium, and other elements.
165
-------
Table 3-1. LEAD MINERALS, BY NAME, AND COMPOSITION
Mineral
Galena
Anglesite
Cerussite
Pyromorphite
Vanadinite
Crocoite
Wulfenite
Linarite
Composition
PbS
PbS04
PbC03
Pb5ci(po4)3
Pb5ci(vo4)3
PbCr04
PbMo04
PbO-CuO-S03 H20
Lead, %
88.6
68.3
77.5
76.3
73.0
63.9
56.4
166
-------
Almost all the lead is produced in this country from domestic ores.
Very little is from imported concentrates. A considerable quantity is
processed first in zinc smelters, the residues then being sent to lead
smelters for lead recovery.
Products
Lead bullion, more than 99.9 percent pure, is the primary product
of this industry. Antimonial lead, a less ductile product, is also
produced. By-products may or may not be produced, depending on the
source of the ore and on market conditions. Some smelters are able to
produce materials rich in gold, silver, arsenic, cadmium, and bismuth,
or to process some of these elements to saleable products. The lead and
zinc industries are so interrelated that allocation of the by-products
by industry is difficult.
Companies
Four companies own and operate the six lead smelters in this coun-
try. Each of these companies also operates a zinc smelter; the lead
and zinc operations are not located near each other, except in Kellogg,
Idaho.
Table 3-2 lists the six smelters, with applicable data. Two of the
smelters are relatively new plants, and the St. Joe Minerals facility
has been extensively remodeled in recent years.
Environmental Impact
The primary lead industry emits sulfur dioxide to the atmosphere
and creates wastewaters containing heavy metals in solution. The slag
fuming operations may be a major source of emission of toxic metal fumes
to the atmosphere. Two slags are produced by these operations that may
be of major importance in discharge of heavy metals into watercourses.
Unlike the copper industry, most lead smelters are located in areas
where rate of rainfall exceeds evaporation. Surface evaporation of
wastewaters is not possible. Water pollution from these facilities is a
principal concern.
Lead smelting produces most of the sulfur dioxide emissions in the
sintering process, with lesser amounts from blast furnaces. None of the
lead smelters exercise complete control of these emissions, and a few
practice no controls for SO.
167
-------
Table 3-2. DOMESTIC PRIMARY LEAD PRODUCERS
CTl
00
Company name
Amax, Inc
Asarco, Inc.
The Bunker Hill Co.
St. Joe Minerals
Location
Boss, Missouri
East Helena, Montana
Omaha, Nebraska
Glover, Missouri
Kellogg, Idaho
Herculaneum, Missouri
Description
Lead smelter
Lead smelter
Lead smelter
Lead smelter
Lead smelter
Lead smelter
Capacity
ton/yr
140,000
350,000
192,000
90,000
125,000
230,000
Number of
employees
5300*
240
325
166
5475*
620
Date plant
built
1968
1888
1870
1968
1918
1892
* Including zinc mining personnel.
-------
INDUSTRY ANALYSIS
This industry analysis examines each production process to define its
purpose and its environmental effects. Each process is analyzed as
follows:
1. Function
2. Input Materials
3. Operating Conditions
4. Utilities
5. Waste Streams
6. Control Technology
7. EPA Classification Code
8. References
This section includes only the processes that are now operating in
the United States or that are under construction. Figure 3-1 is a
flowsheet showing the processes, their interrelationships, and their
major waste streams.
169
-------
Alt
I SOUD
Figure 3—1. Lead industry flow sheet.
170
-------
Figure 3-1 (continued). Lead industry flow sheet.
171
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 1
Mining
1. Function - Ore deposits containing economically recoverable amounts
of lead are taken from the ground and transported to an ore concentra-
tion plant. Most lead ore is obtained from underground mines that use
normal underground stoping methods!>2 such as .block caving, room-and-
pillar with and without rock bolting, and cut-and-fill with timber
supports. After the ore is cut from the deposit, it is taken to the
surface by rail tram, trackless shuttle cars, or conveyor belts and is
then transported to ore concentrating facilities by rail car, truck,
belt conveyor, or a combination thereof.
2. Input Materials - The major lead-containing minerals, with com-
pos i tTonlmarTelid~~cohtent, are presented in Table 3-1. The most common
are galena (lead sulfide), cerussite (lead carbonate), and anglesite
(lead sulfate). Galena ore deposits are the most abundant in nature and
are the most frequently used in the United States as a source of lead.
The deposits usually contain other elements such as zinc, gold, cadmium,
antimony, arsenic, and bismuth. In a few areas, however, such as south-
eastern Missouri, the ore deposits are characterized by simple mineraliza-
tion and virtual exclusion of other minerals.
The economically important deposits of lead ore in the United
States occur either as cavity fillings or replacements, the origin of
which is associated with deep-seated igneous rock.
A mixture of ammonium nitrate and fuel oil (AN-FO) is used for
blasting operations. Sodium nitrate is added to the mixture to increase
blasting power.3
3. Operating Conditions - Mining is done at ambient conditions.
4. Utilities - Electricity is used for operation of equipment in
underground mining and transport.
Diesel fuel and electricity are required for ore transport equip-
ment at the surface.
Specific energy requirements for the mining equipment are not
reported.
A small quantity of water is required for miscellaneous uses such
as equipment washing, dust control spraying, and sanitation facilities.
172
-------
5. Waste Streams - The mining of lead ore generates dust in drilling,
blasting, loading, and transport operations. Estimated average fugitive
dust emission is 110 grams per metric ton of ore, based upon observa-
tions from several types of nonferrous mining.
Substantial amounts of solid waste result from underground mining
operations, the estimated average for 1973 being 0.13 ton per ton of ore
mined.^ This waste material consists of the country-rock bed surround-
ing the ore body plus low-grade lead ore contained in it. The normal
method of disposal is to pile this spoil in a location near the mouth of
the mine.
Wastewater from lead mining results from several sources, such as
seepage of surface water, interception of aquifers, run-off from spoil
dumps, and water sent into the mine for utility purposes. The water is
pumped from the mine at a rate necessary to maintain mining operations.
The required pumping rate bears no relation to the ore output and is
subject to seasonal variation. The rate can range from a few cubic
meters to thousands of cubic meters (hundreds to millions of gallons per
day).
Along with small amounts of oil and hydraulic fluid resulting from
spills or leaks, the wastewater contains dissolved and suspended solids
that reflect the composition of the ore being mined. An analysis of
wastewater from a Missouri mine is given in Table 3-3. In general,
chemical characteristics of the water are typical of those from any
sulfide mine. These characteristics are described in Section 6 of this
report.
Other constituents and their concentrations, typical of mine waste-
waters, are given below.^»°
Component
Mercury
Cadmium
Chromium
Manganese
Iron
Sulfate
Chloride
Fluoride
Concentration, mg/1
0.001 to 0.002
< 0.002 to 0.058
< 0.010 to 0.17
< 0.02 to 57.2
< 0.02 to 2.5
63.5 to 750
< 0.01 to 57
0.063 to 1.2
6. Control Technology - Fugitive dust emissions are controlled by the
manual use of water sprays as needed.
The solid waste or spoil generated by the mining operation is often
used as support and landfill material for highway construction. When it
cannot be so used, it is placed in a spoils dump located so that it will
not contaminate a stream or underground aquifer.
173
-------
Table 3-3. ANALYSIS OF A MISSOURI MINE WATER*
pH
7.8
7.8
7.8
8.1
8.1
Cu
yg/i
4
4
0.5
0.7
Pb
yg/i
65
20
36
31
53
Zn
yg/i
8
15
19
22
36
Mg
mg/1
13
14
15
12
13
Ca
mg/1
30
27
29
30
30
Na
mg/1
32
42
53
54
K
mg/1
4.2
3.8
3.2
3.0
174
-------
Wastewater is generally treated by liming and impoundment as
practiced in copper mining. Management of wastewater is discussed in
detail in the section covering all sulfide mines. Since water from
Missouri mines is already basic, liming may not be required for pH
adjustment. Water from western lead mines is acidic and is treated
similarly to that from copper mines. After treatment, the wastewater is
reused in ore milling operations.
7. EPA Source Classification Code - None
8. References -
1. Mineral Facts and Problems, Washington, D.C. U.S. Department
of the Interior, Bureau of Mines, 1970.
2. Minerals Yearbook. Washington, D.C. U.S. Department of the
Interior, Bureau of Mines, 1973.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
4. Development Document for Interim Final and Proposed Effluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source Cate-
gory Volumes I and II. Environmental Protection Agency,
Washington, D.C., EPA/1-75/032-6. February 1975.
5. An Interdisciplinary Investigation of Environmental Pollution
by Lead and Other Heavy Metals from Industrial Development in
the New Lead Belt of Southeastern Missouri. Bobby G. Wixson,
et al. University of Missouri, Rolla and Columbia, Missouri.
June 1974.
6. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a.
175
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PRIMARY LEAD PRODUCTION PROCESS NO. 2
Concentrating
1. Function - Concentrating is the process whereby the lead-containing
portions of the ore produced by the mine are isolated from the fractions
low in desirable mineral content. Except for high-grade galena ore
produced in southeastern Missouri, ore concentration is required to
produce feed material suitable for subsequent metal recovery processes.
The process consists of milling the ore by crushing and grinding, fol-
lowed by separation into two or more fractions. The fractions rich in
desired minerals are called concentrates, and the fractions low in
mineral content are called gangue.
Separation is achieved by gravity and froth flotation methods. The
gravity method achieves separation because of differences in specific
gravity of the lead-rich minerals and the gangue particles. The flota-
tion method achieves separation by the use of compressed air and chemi-
cal additives that create a froth in which finely divided mineral par-
ticles are floated from the gangue. In some applications, the flotation
method serves as a supplement to gravity separation to improve the
concentrate.
Lead producers of the Mississippi Valley and the eastern United
States use gravity separation because there is considerable difference
in specific gravities of the ore minerals and the gangue. Since the
milled ore particles need not be as small as those required for flota-
tion, the milling costs are lower. Two modes of gravity separation are
commonly used, jigging and float-sink. In jigging, the crushed ore
particles are fed to an agitated, water-filled jigging chamber where the
heavier ore particles gravitate to the bottom and the lighter gangue is
displaced to the top and removed. The float-sink mode utilizes a
liquid medium, such as an aqueous ferrosilicon suspension, with a spe-
cific gravity between that of the lead mineral and the gangue. The
mineral particles sink, while the gangue floats to the top for ^emoval
by skimming.
Flotation is practiced chiefly by lead mines in the western United
States. The method is described in detail under Process 2 of copper
production. The concentrate recovered from the flotation cells contains
45 to 78 percent lead, the percentage depending on the type and grade of
crude ore and its susceptibility to flotation. The concentrate also
contains varying amounts of other valuable elements.
The ore concentrate from the flotation cells requires dewatering
before shipment to smelters. The slurry is fed to thickeners, and floc-
culating agents such as lime and alum are added to improve the settling
rate and fines collection. The thickened slurry of about 50 percent
solids is vacuum filtered and dried to a product containing 6 to 15
percent moistureJ
176
-------
Typical analyses of southeastern Missouri and western lead ore
concentrates are presented in Tables 3-4 and 3-5, respectively.
2. Input Materials - Lead content of the sulfide ores fed to concen-
trating plants ranges from 3 to 8 percent, except for the high-grade
Missouri ores in which lead content exceeds 70 percent.3 Mineralogy of
the western ores is complex and diverse.
Table 3-6 lists the flotation chemicals and the amounts required
for processing lead ore, indicating also some of the less commonly used
agents.
3. Operating Conditions - All concentrating operations take place at
atmospheric pressure and ambient temperatures.
4. Utilities - Although water usage varies with the degree of pro-
cessing, usage is approximately 4 cubic meters per metric ton of ore
processed.5
Electricity is used to operate grinding equipment and generate
compressed air.
5. Waste Streams - Except for local noise, fugitive dust emissions are
the only type of pollutant warranting consideration. Compositions of
the dust are not specified. Crushing operation generate, on the aver-
age, 3.2 kilograms (7 Ib) of particulate emissions per ton of ore pro-
cessed; 0.9 kilograms (2 Ib) is attributable to the crushing and grind-
ing operations and 2.3 kilograms (5 Ib) to material transport and
storage.2»5,6
Liquid waste from the concentrating operation is in the form of a
tailings slurry discharged to the tailings pond. Approximately 4 cubic
meters (1,060 gallons) of tailings slurry is discharged per metric ton
of ore processed.5,7
The flotation and conditioning chemicals are present in the waste-
water either as a floating layer or a solute. In general, lead sulfide
flotations are run at elevated pH levels (8.5 to 11) with frequent pH
adjustments with hydrated lime. This alkaline wastewater dissolves only
small amounts of heavy metals, but can carry the mineral particles in
suspension.
Wastewaters leaving a concentrating operation contained metals as
shown in Table 3-7. These were the only metals investigated, and others
could have been present in greater than normal concentrations. Concen-
trations of calcium, magnesium, sodium, and potassium in mill waters are
significantly higher than those in stream water.
177
-------
Table 3-4. TYPICAL SOUTHEASTERN MISSOURI LEAD CONCENTRATE ANALYSES
(percent by weight)
1
Ag
44.2
43.5
Pb
74.8
76.1
Cu
0.64
0.85
Zn
1.05
1.29
Fe
2.08
1.04
Ni
0.10
0.2
Co
0.06
0.08
S
15.1
15.4
As
0.009
0.006
Insol
1.1
1.3
CaO
1.34
0.94
MgO
0.90
0.75
00
-------
Table 3-5. WESTERN LEAD CONCENTRATE ANALYSES'
Constituent
Pb
Zn
Au
Ag
Cu
As
Percent
45-60
0-15
0-.05 kg/ ton
0-1.4 kg/ton
0-3
0.01-0.40
Constituent
Sb
Fe
insolubles
CaO
S
Bi
Percenta
.0.01-2.0
1.0-8.0
0.5-4.0
tr-3.0
10-30
tr-0.1
Unless otherwise indicated; tr = trace.
179
-------
Table 3-6. FLOTATION CHEMICALS
Chemical
Amount used,
kg/metric ton of ore
Na^CO- (conditioner)
CaO (conditioner)
CuS04 (activater)
Sodium isopropol xanthate (collector)
Pine oil (frothers)
NaCN (depressant)
0.45 - 0.9
0.9 - 18.
0.36 - 0.55
0.0045 - 0.09
0.09
0.045 - 0.14
LESS COMMON FLOTATION REAGENTS
Reagent
Purpose
Methyl isobutyl-carbinol
Propylene glycol methyl ether
Long-chain aliphatic alcohols
Potassium amyl xanthate
Dixanthogen
Isopropyl ethyl thionocarbonate
Sodium diethyl-dithiophosphate
Zinc sulfate
Sodium dichromate
Sulfur dioxide
Starch
Frother
Frother
Frother
Collector
Collector
Collectors
Collectors
Zinc depressant
Lead depressant
Lead depressant
Lead depressant
180
-------
Table 3-7. LEAD MILL WASTEWATER ANALYSIS7
Component
Concentration, mg/1
Mercury
Lead
Zinc
Copper
Cadmium
Chromium
Manganese
Iron
<0.001
0.107 to 1.9
0.12 to 0.46
0.014 to 0.36
0.005 to 0.011
0.002 to 0.02
0.03 to 0.169
0.03 to 0.53
181
-------
Water content of the gangue material from flotation is adjusted to
facilitate hydraulic transport to a tailings pond. Tailings contain
residual solids from the ore, dissolved solids, and excess mill rea-
gents. Typical quantities are 0.9 to 1.1 tons per ton of ore milled.
The main component is dolomite, with small quantities of such constit-
uents as lead, zinc, copper, mercury, cadmium, manganese, chromium, and
i ron.
6. Control Technology - The methods used by copper producers in
concentrating plants are also used at lead concentrators. Details are
given in Process No. 2 of Section 2 of this report, and in Section 6,
water management of sulfide concentrating plants.
To preserve the Ozark area where the new lead belt mining district
of Missouri is located, special attention has been given to wastewater
treatment. Flotation segments in the wastewater are biotically degraded
by algae growth. Use of meandering streams before final discharge to
natural watersheds increases exposure to the algae and provides a good
medium for algae growth. The algae sink to the bottom of the stream and
therefore also act as a solids collector.7 Disposal of algae in event
of excessive growth is not discussed.
7. EPA Source Classification Code - 3-03-010-04
8. References -
1. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/032-a. February 1975.
2. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
3. Mineral Facts and Problems, Washington, D.C. U.S. Department
of the Interior, Bureau of Mines, 1970.
4. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
5. Development Document for Interim Final and Proposed Effluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source Cate-
gory Volumes I and II. Environmental Protection Agency,
Washington, D.C., EPA/1-75/032-6. February 1975.
182
-------
6. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
7. An Interdisciplinary Investigation of Environmental Pollution
by Lead and Other Heavy Metals from Industrial Development in
the New Lead Belt of Southeastern Missouri. Bobby G. Wixson,
et al. University of Missouri, Rolla and Columbia, Missouri.
June 1974.
183
-------
PRIMARY LEAD PRODUCTION . PROCESS NO. 3
Sintering
1. Function - The ore concentrate is treated by a sintering process to
make it suitable for subsequent blast furnace operation. Sintering is
the roasting of blended and pelletized ore concentrate mixtures. The
purposes of sintering are as follows:
1) To provide a feed of proper ratio of lead, silica, sulfur, and
iron for smelting;
2) To convert metallic oxides into oxides or sulfates amendable
to smelting;
3) To drive off volatile oxides such as S09, SO,, As^O,, and
Sb203; and ^ J ^ J
4) To produce a firm, Nporous clinker that is easily fed to a
blast furnaceJ
The process consists of three consecutive steps:
1) Blending of ore concentrates with direct-smelting ores, sinter
recycle, flue dust, and fluxes;
2) Pelletization of the blended mixture; and
3) Roasting of pelleted material.
Blending balances the smelter charge and permits control of im-
purity levels of Zn, Cu, As, Sb, and Bi. Pelleting is achieved by
mixing the blended charge with 6 to 8 percent by weight water in a pug
mill and feeding the mix to a rotating pelletizing drum. Resulting
pellets are 3 to 5 millimeters (1/8 to 1/4 inch) in diameter.
Pellets are introduced onto a horizontal metal belt which takes
them through a sintering machine of 135 metric-ton-per-day capacity. As
the pellets proceed through the sintering machine, they are heated and
undergo oxidizing reactions that convert sulfides to oxides and sul-
fates. Lead silicate forms, and oxides combine to form low-melting-
point silicate complexes, which bind the ore particles together.
The resulting sinter is broken into pieces ranging up to 25 milli-
meters (5 inches) diameter.2 The crushed sinter is screened for removal
of fines, which are recycled to charge preparation. The screened pro-
duct is stored for blast furnace reduction. Table 3-8 gives typical
ranges of components in the sintered product.
184
-------
Table 3-8. SINTER ANALYSIS
1,3,4
Component
Ag
Cu
Pb
S
Fe
Si02
CaO
Zn
Sb
Cd
Weight,
percent
0.03-0.07
0.3-4.5
28-50. 0
0.75-2.0
12-15.5
10.0-15.6
9.0-10.5
4.0-12.5
0.01-1.5
Tr-0.04
Table 3-9. SINTER MACHINE FEEDV
Percent by weight
Ore concentrate
Misc. lead materials
Flux diluent
Sinter recycle
3.1.47
12.44
19.86
36.20
185
-------
2. Input Material - Lead concentrates are the main input material for
sintering. Typical analyses of western and Missouri lead concentrates
are presented in the concentrating process (Process No. 2).
Collected flue dusts, recycled sinter, and smelter residues must be
considered part of the charge for sintering. These sulfide-free fluxes
are added to maintain a specified sulfur content (5 to 7 percent by
weight) in the charge. Silica and limestone are used when needed.
Coke fines, in the amount of 1 percent by weight of the total
charge, are mixed with this charge.6
Typical feed to a sinter machine is given in Table 3-9.
3. Operating Conditions - Temperatures in both the updraft and down-
draft sinter roasting machines reach approximately 600°C. Pressure is
atmospheric.
4. Utilities - In both the updraft and downdraft sintering machines
gas- or oil-fired burners are used to ignite the charge. Energy con-
sumed in the sintering process amounts to 0.5 million kilogram-calories
per ton of lead produced (2 million Btu). A breakdown allocates 40
percent to coke consumption and 60 percent to gas or oil consumption,
gas being used more than oil.7
Water may be added for pelletizing the charge if the moisture
content is below required limits. Air is injected through the charge
while oxidizing in the sintering machines. No quantities are given for
air injection.
Electricity is the power source for fans, feed conveyors and gen-
eral operating equipment. Approximately 20 percent less power is re-
quired for the updraft fans than for downdraft.8
5. Waste Stream - Particulate emissions are approximately 100 to 250
kilograms per metric ton of lead produced in sinter machines.9 Analysis
of the flue dust is roughly 40 to 70 percent Pb, 10 to 20 percent Zn,
and 8 to 12 percent sulfur.9 Depending upon concentrate composition,
the flue dust contains various amounts of Sb, Cd, Ge, Se, Te, In, Tl,
Cl, F, and As.8
In the lead smelting process, only sintering emits enough S02 to
create a serious air pollution control problem. About 85 percent of the
sulfur is removed from the concentrate during sintering. About 50
percent of the remainder is discharged as SO? from subsequent opera-
tions; the balance goes into the slag as sulfates.8,11
186
-------
In the sintering process most of the sulfur is eliminated at the
front end of the conveyor. By the time the charge reaches the end of
the sintering machine, little S02 is being emitted. If the exit gases
are taken from the machine in a single stream, the $02 concentration is
about 2 percent.8,9 The exit gases can be split into two streams, one
predominantly from the front and the other from the rear. This pro-
cedure produces both a weak and a strong S02 stream, 0.5 and 5.7 percent
S02 respectively.8 Only updraft sintering machines have incorporated
this engineering modification. Table 3-10 and 3-11 give weight analysis
of particulate effluents and size distribution, respectively.
Offgases also contain organic vapors from flotation reagents in the
concentrate. The compounds formed from these flotation chemicals by
reactions caused by the sintering temperatures are not known. Traces of
HF and SiF4 may be found in these gases. The volume of gases emitted is
a function of machine size and material throughput, ranging from 0.25 to
0.50 standard cubic meters per minute per square meter of bed area (0.8
to 1.6 scfm/sq ft).9 Temperatures of the gases normally range from 150
to 400°C.5,8 Flow rates may range from 58,000 to 66,000 standard cubic
meters per hour (34,000 to 37,000 scfm).9
Table 3-12 gives a typical analysis of gases from a sintering
machine. Although a small percentage of arsenic trioxide in the gaseous
form is given off in these exit gases, amounts are unknown.
6. Control Technology - The particulate control devices discussed in
the copper section, Section 2, Process Nos. 3 and 6, are used for gases
from the sintering process. Particulates from the sinter machine are
collected by several different methods. Four of the six plants use
baghouses, and the other two use an ESP. Efficiencies range from 95 to
99.8 percent. Table 3-13 gives current atmospheric controls on lead
sintering processes.
The strong gas stream collected from separation of updraft exit
gases is the only stream amenable to S02 control by sulfuric acid pro-
duction. In both the weak stream and the combined single stream con-
centrations are too low for contact sulfuric acid production. A de-
tailed discussion of sulfuric acid production is presented in the copper
section, sulfuric acid production (Process No. 12).
Three plants are now controlling S02 in sinter machine off-gases by
use of single-contact sulfuric acid plants, which reduce total S02
emissions by 70 to 80 percent.
Estimated cost of control of the rich sinter machine off-gases by
a single contact acid plant is $0.01 per kg of lead produced. Currently
no lead smelters practice control on weak S02 streams. The best avail-
able control technology for these streams would be chemical scrubbing,
as described in Section 2, Process No. 6.
187
-------
Table 3-10. GRAIN LOADING AND WEIGHT ANALYSIS OF INPUT FEED AND
EFFLUENTS-UPDRAFT LEAD SINTERING MACHINE12
Grain
loading, g/Nm3 (0°C)
16.3
Weight
Pb
Si02
Fe
CaO
MgO
Zn
S
Cu
As
Cd
Se
inerts
analysis, %
35-50
8-11
9-13
7-10
0.7-1
4-6
0.7-1
tr
tr-30
tr
tr
6-8
Table 3-11. TYPICAL SIZE PROFILE OF EFFLUENTS,
UPDRAFT LEAD SINTERING MACHINE12
Size,
micron
20-40
10-20
5-10
< 5
% weight
15-45
9-30
4-19
1-10
188
-------
Table 3-12. ANALYSIS OF SINTER MACHINE EXHAUST GASES
(MISSOURI LEAD OPERATING COMPANY)13
so2
°2
co2
N2
so3
Dust content
Temperature
Moisture content
Range, % by volume
4-7
4-9
3-4
84-85
0.05-0.2
57 g/Nm3
200-350°C
25 percent by vol .
189
-------
Table 3-13. ATMOSPHERIC CONTROL SYSTEMS ON PRIMARY LEAD
SINTERING MACHINES8
Plant
Control system
Bunker Hi 11/Kellogg, Idaho
Amax, Inc./Boss, Missouri
St. Joe/Herculaneum, Missouri
ASARCO/E. Helena, Montana
ASARCO/Glover, Missouri
ASARCO/E1 Paso, Texas
Updraft sintering machine produces
two gas streams: strong gas stream
to acid plant. Weak gas stream
joined with blast furnace and hygiene
air, then goes to a baghouse and out
the stack.
Updraft sintering machine produces
two gas streams: strong gas stream
to acid plant. Weak gas stream
jointed with blast furnace before
exiting out 200 ft stack.
Updraft sintering machine produces
two gas streams: strong gas stream
to acid plant. Weak gas stream joins
other gases, then thru baghouses
and to stack.
Updraft sintering machine. Gases to
water spray, ESP, then dilution air
added and released to stack.
Updraft sintering machine. All gases
to water spray and baghouse, then
out stack.
Downdraft sintering machine?. Gases
treated by scrubbers, a spray chamber,
and a baghouse, then out stack.
190
-------
7. EPA Classification Code - 3-03-010-01.
8. References -
1, Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA 440/1-
75/032-a. February 1975.
2. Davis, W. E. National Inventory of Sources and Emissions:
Copper, Selenium, and Zinc. U.S. Environmental Protection
Agency, (NTIS), Research Triangle Park, North Carolina. PB-
210 679, PB-210 678, and PB-210 677. May 1972.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
4. Hallowell, J. B., R. H. Cherry, Jr. and G. R. Smithson, Jr.
Trace Metals in Effluents from Metallurgical Operations. In:
Cycling and Control of Metals, U.S. Environmental Protection
Agency, Cincinnati, Ohio. November 1972. pp. 75-81.
5. Systems Study for Control of Emissions Primary Nonferrrous
Smelting Industry. Arthur G. McKee & Co. for U.S. Department
of HE&W. June 1969.
6. Fejer, M. E. and Larson, D. H. Study of Industrial Uses of
Energy Relative to Environmental Effects. U.S. Environmental
Protection Agency. Research Triangle Park, North Carolina.
July 1974.
7. Copper Hydrometallurgy: The Third-Generation Plants. Engi-
neering and Mining Journal. June 1975.
8. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina. EPA-450/2-74-002a. October 1974.
9. Jones, H. R. Pollution Control in the Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
10. Phillips, A. J. The World's Most Complex Metallurgy (Copper,
Lead and Zinc). Transactions of the Metallurgical Society of
AIME. Volume 224: pp. 657-668. August 1962.
191
-------
11. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume Il-Primary and Secondary Non-
ferrous Smelting and Refining. Calspan Corporation, Draft,
April 1975.
12. Duncan, L. J. and Keitz, E. L. Hazardous Particulate Pollu-
tion from Typical Operations in the Primary Nonferrous Smelt-
ing Industry. Presented at the 67th Annual Meeting of the Air
Pollution Control Association. Denver, Colorado. June 9-13,
1974.
13. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
192
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 4
Acid Plant
1. Function - The sintering machine produces the only lead smelter
exit gases that are amenable to production of sulfuric acid. A detailed
description of this process is given in the copper section, contact
sulfuric acid plant, Section 2, Process No. 13.
2. Input Materials - Inputs are the, same as those discussed in copper,
sulfuric acid plant (Process No. 13).
3. Operating Conditions - Same.
4. Utilities - Same.
5. Waste Streams - Same.
6. Control Technology - The same types of mist eliminators are used in
lead contact acid plants as in copper acid plants. Particulate scrub-
bers are part of the input gas cleaning system.
Liquid waste effluents are treated by liming and settling in
cooling ponds. Overflows are recycled to slag granulation or discharged.
Tables 3-14 and 3-15 give the treatments now practiced for control of
acid plant blowdown and scrubber wastewaters, respectively, by the three
primary lead smelters having acid plants.
7. EPA Source Classification Code - None
8. References -
1. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/032-a. February 1975.
193
-------
Table 3-14. WASTEWATER TREATMENT AT PRIMARY LEAD ACID PLANTS
1
Plant
Liquid effluent treatment
Discharge
2
3
Enters water treatment plant,
limed, thickened, and filtered,
and sent to reservoir for
recycle.
Recycled to slag granulation.
Enters liming sump, then
passed to lime bed, then
to a cooling pond.
0
273 m /day
(72,000 GPD)
Table 3-15. SCRUBBER WASTEWATER TREATMENT AT PRIMARY LEAD PLANTS
1
Plant
Treatment
Discharge
2
3
Enters water treatment plant,
limed, thickened, filtered,
and then sent to reservoir for
recycling.
Recycled from a cooling tower.
Sent to a lime sump then
to a settling pit. Most
is recycled.
0
Undetermined
194
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 5
Blast Furnace
1. Function - Sintered feed is reduced in the blast furnace process to
produce a crude lead bullion. Specified amounts of coke, limestone, and
other fluxing materials are charged with the sinter through a water-
jacketed shaft at the top of the furnace. The material settles to the
furnace bottom, which is supported by a heavy refractory material.
Air is injected into the charge through side-mounted tuyeres to
effect a more complete formation of metallic oxides and thereby raise
the temperature of the charge. At the operating temperature of the
furnace, coke and resulting carbon monoxide reduce most of the metallic
oxides to yield a molten mass of metal. Some of the metallic impurities
interact with the fluxing materials to form a slag composed mainly of
iron and calcium silicates. Depending upon the composition of the
charge, material in the blast furnace can separate into as many as four
distinct liquid layers.
Copper, present in most lead ores, reacts with residual sulfur to
form a matte that separates into a layer beneath the slag. The matte
typically assays 44 to 62 weight percent copper, 10 to 20 percent lead,
and up to 13 percent sulfur.1 If the sinter is high in arsenic and/or
antimony content, a speiss layer will form under the matte. Speiss
compounds are arsenides and antimonides of iron and other metals. The
bottom layer of lead bullion is 94 to 98 weight percent lead plus
varying amounts of other metals such as copper, tin, arsenic, antimony,
silver, and gold. Typical ranges of composition are shown in Table
3-16.
Upon completion of the process, the crude bullion is charged to
dressing kettles,3 the matte and speiss are sold to a copper smelter,
and the slag is discharged to a fuming furnace. A typical slag analysis
is shown in Table 3-17.
The capacity of most large blast furnaces is 1360 metric tons (1500
U.S. tons) of charge materials per day.
2. Input Materials - The blast furnace charge is made up of sinter,
fluxes, coke, and sundry materials recycled from other smelting opera-
tions. The relative amounts of these materials are presented in Table
3-18.
Normally, coke comprises 8 to 15 weight percent of the furnace
charge. If the blast air is enriched with oxygen, coke consumption is
reduced 10 percent with a 10 to 20 percent increase in smelting rate.
195
-------
Table 3-16. LEAD BULLION COMPOSITION^'0'^
Component
Ag
Au
Cu
S
Pb
Fe
Zn
Sn
As
Sb
Bi
Wt. percent
0.13-0.31
1.6-3.1a
T.0-2.5
0.25
94-98
0.6-0.8
tr.
tr.
0.7-1.1
1.0-1.75
0.01-0.03
a Value for Au in g/metric ton.
tr. - trace
196
-------
Table 3-17. TYPICAL BLAST FURNACE SLAG ANALYSIS2'3'4
Component
Ag
Cu
Pb
FeO
CaO
Zn
insol
MnO
As
Sb
Cd
F
Cl
Ge
S
Weight percent
T .'56-4.69*
0.10b
1.5-3.5
25.5-31.9
14.3-17.5
13.0-17.5
22. 6-26. 5d
2.0-4.5
0.10
0.10
0.10
trc
trc
trc
0.5-1.0
a Values for Ag in grams per metric ton.
Variable, depending on the furnace charge.
c tr = trace.
Insol ubles include MgO - A10 - Si02 phases.
197
-------
Table 3-18. TYPICAL BLAST FURNACE CHARGE'
Component
Sinter
Coke
Miscellaneous products
(zinc plant residues)
Slag (dross)
Silica
Li me rock
Cadmium residue
Refinery dross
Baghouse product
Weight, kg
1250-1650
125-165
0-90
0-225
0-36
0-27
0-9
0-35
0-35
198
-------
3. Operating Conditions - Temperatures in a blast furnace range from
215°C for the charge mass near the top of the furnace to 1220°C in the
slag zone. Slag temperatures range from 1000° to 1220°C and bullion
temperatures, from 900° to 950° C.
Because of the exhaust gas configuration, the blast furnace oper-
ates at a pressure slightly above atmospheric.
4. Utilities - Air is injected into the charge at a pressure of 0.1 to
0.3 kilograms per square centimeter.3 Consumption of 140 to 175 cubic
meters per hour is required for a charge of 1360 metric tons per day.3
Cooling water is employed to control furnace temperatures through
jacketed shafts.
5. Waste Streams - Particulate emission rates in blast furnace exhaust
gas range from 125 to 180 kilograms per metric ton of bullion produced/'5
Particle sizes of the dust range from 0.03 to 0.3 micron.5 The dust is
composed of oxides, sulfates, and sulfides of the various metals present
in the furnace charge plus chlorides, fluorides, and coke dust.5*5
Undiluted gas temperatures are estimated to be 650° to 750°C,2»5
with theoretical flue gas rates of 170 to 400 standard cubic meters per
minute. After dilution by air and water vapor, however, volume typically
increases from 9 to 12 times the theoretical flow.1
Exhaust gas analysis after air dilution and CO combustion is re-
ported as follows.5'?
Component
co2
°2
C0a
S00
2
N2
Percent by volume
15
15
5
0.05
Remainder
CO concentration estimated to be 25 to 50 volume
percent prior to combustion.
Others have reported that CO and S02 concentrations, although highly
variable, average 2.0 and 0.01 to 0.25 volume percent, respectively.1'5
An estimated 10 to 20 weight percent of total sulfur in the feed concen-
trate is removed in the blast furnace, half emitted as S02 and the rest
retained in the slag or matte.8
199
-------
Table 3-19. ATMOSPHERIC CONTROL SYSTEMS ON PRIMARY
LOCAL BLAST FURNACES7
Plant
Control system
Bunker Hi 11/Kellogg, Idaho
Amax, Inc./Boss, Missouri
St. Joe/Herculaneum, Missouri
ASARCO/E. Helena, Montana
ASARCO/Glover, Missouri
ASARCO/E1 Paso, Texas
Blast furnace gas stream joined
to weak sinter gas stream and
hygiene air, then to baghouse
then to stack.
Blast furnace gases join sinter
weak gases, then to baghouse,
and then out the stack.
Blast furnace gases join sinter
weak gases and other gases pass
thru baghouses and stack.
Blast furnace gases join reverb
and ventilation gases, then pass
thru three baghouses in parallel
with stack for each house.
Blast furnace gases to water
spray, baghouse, and three
stacks.
Blast furnace and dross furnace
gases mix, then pass thru a
spray chamber and a baghouse,
then out six stacks.
200
-------
Slag from the blast furnace can be discharged and granulated with
cooling water instead of sending it through a slag fuming furnace. If
this procedure is followed, the slag fuming process is eliminated.
6. Control Technology - The dilute S02 emissions from the blast fur-
nace are not controlled at lead smelters. The best available control
technology is chemical scrubbing.
Particulates in blast furnace exhaust gases are controlled at all
smelters by means of baghouses. Control efficiency ranges from 95 to
99.9 percent. Table 3-19 describes controls for blast furnace gases.
See Process No. 6 of the copper section for details.
Current practice for slag disposal is to convey it hydraulically
with the granulating water stream to a dump or tailings pond. Recom-
mended technology includes use of concrete settling pits, ground sealing
of disposal area, and diversion of runoff to a water treatment lagoon.
Cost of such technology is estimated at $1.37 per metric ton of lead
produced.
7. EPA Classification Code - 3-03-010-02
8. References -
1. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW, June
1969.
2. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
3. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
4. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA 440/1-
75/032-a. February 1975.
5. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
6. Phillips, A.J. The World's Most Complex Metallurgy (Copper,
Lead, and Zinc). Transactions of the Metallurgical Society of
AIME. Volume 224: pp. 657-668. August 1962.
201
-------
7. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. Department
of HE&W, June 1969.
8. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
202
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 6
Slag Fuming Furnace
1. Function - A slag fuming furnace is used to recover metal values
otherwise lost in the slag. Blast furnace slag contains appreciable
concentrations of zinc, since most lead ores contain zinc and many
smelters add zinc to the blast furnace charge in the form of zinc re-
sidues containing as much as 12 weight percent zinc.
The slag is fed to the fume furnace in a molten state, and pul-
verized coal is added to maintain temperature by combustion. By holding
at an elevated temperature, a fume is generated from the slag that
contains zinc, germanium, lead, cadmium, chlorine, and fluorine. The
fume is condensed and processed to recover primarily zinc and lead. The
lead fraction is recycled to the sintering process. The presence of
chlorine and fluorine in the fume necessitates a leaching step to avoid
accumulation of these elements.
A matte is sometimes separated from the slag in this operation for
recovery of substantial amounts of copper and silver from the dezinced
slag. When fuming has subsided, the slag is dumped and cooled with
water.
2. Input Materials - Composition of the blast furnace slag introduced
to the fume furnace is shown in Table 3-17 of Process No. 5. Two tons
of slag are generated per ton of crude lead bullion produced by the
blast furnace.
Pulverized coal is added to maintain temperature by combustion.
The amount is not specified in the literature.
3. Operating Conditions - The slag temperature range is 1000° to
12000-C. Atmospheric pressure is usedj
4- Utilities - Air is injected into the furnace for combustion of the
coal. Quantities are not cited in the literature.2
Water is used for slag cooling and granulation in amounts ranging
from 200 to 8,200 cubic meters per day,3 the amount depending upon the
cooling water circuit of a given plant. A typical analysis is given in
Table 3-20.
5. Waste Streams - The gas stream emanating from the furnace typically
has a low SO? concentration. The literature? cites a value of 0.02
volume percent for a flow rate of 5,660 standard cubic meters per
minute. Gas stream temperature is about 1200°C.
203
-------
Table 3-20. WASTE EFFLUENTS FROM SLAG GRANULATION
Parameter
PH
Alkalinity
COD
Total solids
Dissolved solids
Suspended solids
Oil and grease
Sulfate (as S)
Chloride
Cyanide
Aluminum
Arsenic
Cadmium
Calcium
Chromium
Copper
Iron
Lead
Magnesium
Mercury
Molybdenum
Nickel
Potassium
Selenium
Silver
Sodium
Tellurium
Zinc
Total'
plant
intake
mg/1
7.6
203
8
-
408
3
-
145
18
-
• -
-
-
70
-
0.02
1.70
0.12
.31
-
-
0.03
-
-
-
-
0.05
Total
plant
discharge
mg/1
8.3
186
8
-
500
36
-
215
-
-
-
-
-
-
0.02
-
0.30
-
-
-
0.04
-
-
-
-
-
0.12
Net
change,
mg/1
-17
0
-
92
33
-
70
-
-
-
-
-
-
-
0
-
0.18
-
-
-
0.02
-
-
-
-
-
0.38
Net loading
kg/ton
NLCa
0
-
0.89
.32
-
0.67
-
-
-
.
•
. - .
'
0
-
0.0018
-
•
-
0.00018
-
•-
'
-
-
0.0037
Process water flow: 6 million liters/day (1.55 million gal/day)
Production: Metric tons/day (630 short tons/day).
Source: This contract and 1971 RAPP data.
Notes:
a NLC = no load calculable.
204
-------
The furnace gas exhaust contains high concentrations of particulate
and fume composed of the volatile components of the blast furnace slag.
The literature does not cite quantities or composition.
The dumped slag and water used for granulation constitute the major
waste stream for the process. The slag is made up of various compounds
of iron, calcium, silicon, aluminum, magnesium, and other elements. The
water-soluble portions are leached by the cooling water. Table 3-20
presents analyses of the intake and outflow streams of water used for
slag granulation.
6. Control Technology - The exhaust gas from the fuming furnace is
cooled by waste heat boilers or cooling chambers before being sent to
baghouses for removal of particulate and condensed volatiles. Baghouse
operation is limited to a maximum temperature of 285°C. Particulate
removal efficiency ranges from 95 to 99 percent.4 It is preferable to
operate the baghouse at the lowest possible temperature to allow removal
of volatile matter contained in the gas stream. The copper section,
Process No. 6, gives additional details.
Slag disposal is the same as described in Process No. 5, involving
conveyance with the granulating water stream to a dump or tailings pond
use of concrete settling pits, ground sealing, and diversion of run-off.
With respect to the slag granulation water, current treatment and
disposal practices are variable, as summarized in Table 3-21. Normally,
it is desirable to recycle the water after cooling and clarification. A
smaller stream is bled off to neighboring surface water to control
buildup of water-soluble components. The best available control tech-
nology for waste treatment is a combination of neutralization and clari-
fication. With this technology, the following effluent concentrations
can be realized:
Component
Cadmium
Lead
Mercury
Zinc
Concentration,
0.5
0.5
0.005
5.
mg/1
These values are currently being met by five of the six lead smelters.
The estimated cost of treatment is $2.68 per metric ton ($2.43 per U.S.
ton) of lead produced. Further information is given in Water Management
(Section 6.0).
%
7. EPA Classification Code - None
205
-------
Table 3-21. . PRIMARY LEAD SLAG GRANULATION
WASTEWATER TREATMENT5
Plant
1
2
3
4
5
6
Treatment
Sent to cooling pond.
Sent to settling pit then to
a cooling pond.
Sent to settling pond and
recycled.
Sent to two settling ponds in
series.
Sent to a slag pile.
No data.
Discharge
8,230 m3/day
(2,200,000 gpd)
273 m3/day
(72,000 gpd)
0
Discharge is present
but no quantities
are available.
No apparent discharge
to surface. Leaching
is not mentioned.
No data
206
-------
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur 6. McKee & Co. for U.S. DHEW, June
1969.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
4. Vandegrift, A.E. (Dr.), L.J. Shannon (Dr.), P.G. Gorman, E.W.
Lawless (Dr.), E.E. Salle, and M. Reichel. Particulate
Pollutant System Study - Mass Emissions, Volumes 1, 2, and 3.
U.S. Environmental Protection Agency, (NTIS), Durham, North
Carolina. PB-203 128, PB-203 522 and PB-203 521. May 1971.
5. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA 440/1-
75/032-a. February 1975.
207
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 7
Drossing
1. Function - Dressing is the initial step in refining lead bullion.
Molten bullion from the blast furnace is placed into dressing kettles of
90 metric-ton capacity where submerged air lances provide agitation and
oxidation. The kettle and molten bullion are cooled to a temperature at
which lead is still a liquid but oxides of the common impurities, espe-
cially copper, and oxides of lead solidify. The term "dross" refers to
any solid scum floating on top of a metal bath. Dross may contain
varying amounts of lead, copper, tin, indium, arsenic, antimony, and
bismuth. Because of the high specific gravity of molten lead, these
solid oxides float and are easily skimmed off the molten lead. Dressing
of the blast furnace bullion always occurs before the lead is sent to
the refinery.
For more complete removal of copper, sulfur is added to the dres-
sing kettle. This sulfur combines with the remaining copper forming
cuprous sul-fide, CU2$, which floats and is skimmed off with the rest of
the dross. By dressing, a bullion with copper content as high as 2
percent can
By dressing, a bullion with copper content as high as
be reduced to approximately 0.005 percent copper.'
The dross is skimmed and sent to a dross reverberatory furnace for
further treatment and recovery of marketable products. Dross may typi-
cally contain 90 percent lead oxide, 2 percent copper, and 2 percent
antimony, with other values such as gold, silver, arsenic, bismuth,
indium, zinc, tellurium, nickel, selenium, and sulfur. The collected
dross amounts to 10 to 35 percent of the blast furnace bullion.1 A
typical assay of drossed bullion to the refinement process is shown in
Table 3-22.
2. Input Materials^ - During the dressing procedures sulfur is added in
the ratio of approximately 1 kilogram per ton of bullion from the blast
and dross reverberatory furnace. Various amounts of coal or co.ce,
ammonium chloride, soda ash (Na2C03), and litharge or baghouse fume
(PbO) are added to the kettles as needed.
3. Operating Conditions - The molten bullion is cooled to a tempera-
ture of 370 to 500°C and maintained within that range. Pressures are
atmospheric.!
4. Utilities - Most of the dressing kettles are heated with natural
gas. About 1.1 million kilogram-calories (4 million Btu) per metric ton
are consumed during this stage, of which 90 percent is allocated to gas
and 10 percent to oil.4
208
-------
Table 3-22. LEAD BULLION ANALYSIS
Basis: As drossed
1,2,3
Element
Ni
Ag
Au
Cu
Fe
Te
As
Sb
Bi
Se
Sn
Wt. percent
tr.
0.13-0.31
1.2-3.19
0.08-0.005
0.7-0.8
0.01-0.03
0.7-1.1
1.0-1.75
0.01-0.03
tr.
tr.
Value of Au in g/metric ton.
tr. = trace.
209
-------
Conveyors, agitators, pumps, and similar equipment are powered by
electricity. No quantities are given for electrical power consumption.
Air is injected by submerged lances for supplementing oxidation and
agitation.1»5>6 Quantities of air are not given in the literature.
5. Waste Streams - The dressing operation generate small amounts of
air pollutants and slag. The air pollutants are S02 and volatile com-
ponents of the lead bullion. Table 3-16 presents a typical analysis of
the bullion. Varying quantities of copper, iron, arsenic, zinc, cad-
mium, antimony, and bismuth are carried off; it is believed that the
quantities are very small because of the low temperatures. One source
quantifies the particulate loading as between 1.0 and 21.7 grams per cu.
meter of off-gas.' Another cites the emission rate as 10 kilograms per
metric ton of lead produced. The SO? content of the off-gas is very
low, usually less than 0.05 percent by volume. Flow rates of exit gases
from a blast furnace are typically 5,100 to 5,500 Nm^ per minute.
Temperature of these gases are low, approximately 200 to 350°C.6
6. Control Technology - No control methods are presently applied to
the weak S02 stream emitted from the dressing operation. The best
available technology is chemical scrubbing. No data were found on the
costs of applying this control to dressing kettles.
Particulates and fumes from the dressing kettles are controlled
with the blast furnace off-gas at all plants. As mentioned earlier,
five plants use baghouses and one plant uses an ESP. Control efficiency
is estimated at 95 to 99.8 percent. Additional information is given the
copper section on reverberatory smelting (Process No. 6).
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA 440/1-
75/032-a. February 1975.
3. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
210
-------
4. Fejer, M.E. and Larson, D.H. Study of Industrial Uses of
Energy Relative to Environmental Effects. U.S. Environmental
Protection Agency. Research Triangle Park, North Carolina.
July 1974.
5. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. Environmental Protection Agency, Research Triangle
Park, North Carolina, EPA-450/2-74-002a. October 1974.
6. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW, June
1969.
7. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
211
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 8
Dross Reverberatory Furnace
1. Function - Dross removed from the lead bullion by dressing requires
further treatment for separation of components. As cited in Process No.
7, the dross is composed of about 90 percent lead oxide, 2 percent
copper, 2 percent antimony, and lesser amounts of other elements.
Prior to smelting, the dross may be treated with soda ash, litharge
or baghouse fumes, coke, and sulfur to produce matte and speiss with
high ratios of copper to lead. Whether treated or not, the dross is
charged into a reverberatory furnace along with pig iron, silica, and
limerock (optional).
When smelting of the charge has been completed, the products sepa-
rate into four layers: slag on top, matte and speiss intermediate, and
molten lead at the bottom. By the use of suitably placed taps on the
furnace, each layer can be removed separately.
The slag, amounting to 2 to 4 weight percent of the dross charge,
is returned to the blast furnace for smelting. The slag typically
assays 6 percent Cu, 38 percent Pb, 11 percent Zn, 11 percent FeO, and
16 percent
The matte and speiss are tapped separately, granulated, and shipped
to copper smelters for recovery of copper and precious metals. The
matte amounts to 10 to 14 weight percent of the dross charged; the
speiss, 20 to 30 percent. The collective assay of these materials is 42
percent Cu, 38 percent Pb, 16 percent S, 1 percent Fe, and 0.6 percent
As, plus small amounts of zinc, rare earths, and precious metal sJ
The lead layer is 94 to 98 percent lead and comprises 50 weight
percent of the dross charged. It is returned to the blast furnace.
2. Input Materials - Along with the dross charged, the process re-
quires the addition of pig iron, silica, and limestone. The amounts of
these materials vary with each charge, depending upon dross composition.
If soda treatment is used, equal amounts of soda ash, litharge,
coke, and sulfur are added. Each is 3 to 5 percent by weight of the
dross charged.
3- Operating Conditions - Smelting temperatures are the same as those
in the blast furnace, ranging from 1000° to 1200°C. Smelting is done at
atmospheric pressure.
212
-------
4. Utilities - Gas or oil fuels are used for heating and maintenance
of temperature. Quantities are not given in the literature.
5. Waste Streams - Atmospheric emissions are the only form of pollu-
tion from the dross reverberatory furnace. All solids are recycled or
marketed. Water used in matte and speiss granulation is evaporated
before transport.
Particulate emission rates are 10 kilograms per ton of lead pro-
duced.^ The reference does not indicate whether this emission includes
condensed fume. Exhaust gases are about 760 to 980°C.2
Sulfur dioxide, carbon dioxide and monoxide, sulfur trioxide, and
nitrogen and its compounds are released to the atmosphere as products of
combustion. The exit gas volume from a dross reverberatory ranges from
30 to 170 Nm^ per minute.2,3 The S02 content of this gas is usually
below 0.05 percent.
6. Control Technology - No control methods are now used for the weak
S02 stream emitted from the dressing reverberatory. The best available
technology is chemical scrubbing.
Particulates and fumes from the dressing reverberatory are combined
with the blast furnace off-gas at all plants. As mentioned earlier,
five plants use baghouses and one plant uses an ESP. Control efficiency
is estimated at 95 to 99.8 percent.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
3. Systems Study for Control of Emissions Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. Department
of HE&W, June 1969.
213
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 9
Cadmium Recovery
1. Function - The flue dusts generated by lead smelting are processed
to recover cadmium values. Since dust from blast furnace exhaust gases
is recycled to the sintering machine, sinter dust becomes enriched in
cadmium, thallium, and zinc. When cadmium content in the dust reaches
12 weight percent or greater, the dust is subjected to a separate
roasting operation for cadmium separation and recovery.!
Roasting, performed in a Godfrey roaster (See Primary Copper
Production, Process No. 16), causes the cadmium, together with thallium,
indium, and selenium, to be volatilized from the dust charge.2 The
fume is cooled and collected for shipment to a zinc smelter for further
processing. The roaster residue, which contains zinc, lead, and antimony,
is returned to the .sintering machine.
2. Input Materials - Flue dust collected from the sintering machine
exhaust gases is the only input material.
3. Operating Conditions - Conditions of temperature and pressure are
not reported.
4. Utilities - Equipment can be oil- or gas-fired. No quantities are
specific.
5. Waste Streams - Fume emissions from the roaster are cooled and
recovered as product. The roaster residue is recycled. Data for fume
capture are not furnished in the literature.
6. Control Technology - The flue dust and fume emitted from the
roaster can be contained by further cooling with water sprays and col-
lection in a baghouse.
7. EPA Source Classification Code - None
8. References -
1. Systems Study for Control of Emission Primary Nonferrous
Smelting Industry. Arthur G. McKee & Co. for U.S. DHEW, June
1969.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
214
-------
PRIMARY LEAD PRODUCTIONS PROCESS NO. 10
Reverberatory Softening
1. Function - The drossed lead bullion is further purified by a pro-
cess termed "softening", which entails removal of the elements that make
lead hard, notably arsenic, antimony, and tin. Several softening
processes can be used (See Process Nos. 11 and 12).1.2
The reverberatory method is similar to the dressing procedure and
is particularly applicable for processing bullion with a wide range of
impurities. The bullion is charged into a reverberatory furnace,
melted, and blown with air to effect oxidation of the arsenic, antimony,
tin, and other impurities. If hardness of the feed bullion is greater
than 0.3 to 0.5 weight percent antimony equivalent, litharge is added to
the charge to increase the rate of impurity oxidationJ
Furnaces with capacities up to 300 metric tons are used for the
process. The oxides rise to the surface to form a slag that is skimmed
off and can be further treated to recover contained values. The softened
lead is tapped from the bottom of the furnace and pumped to the desil-
verizing process. Table 3-23 presents typical analyses of softened
bullion product and the softened slag. Hardness of the softened bullion
is less than 0.03 weight percent antimony equivalent.
2. Input Materials - An analysis of drossed lead was presented in
Table 3-22 (Process No. 7). Litharge is added only when especially hard
bullion is processed. Coke or coal may be added to inhibit the oxida-
tion of lead.
3. Operating Conditions - Temperatures during softening reach 700°C.
Pressures are atmospheric.1
4. Utilities - Electricity is the power source for mechanical agita-
tors, pumps, and conveyors.
Most of the heat is supplied by the exothermic oxidation of im-
purities. Gas or oil is used to begin the reaction and for temperature
maintenance.
Air is injected through lances or pipes into the bath. Air con-
sumption is not reported.
5. Waste Streams - The air blow from the furnace is the only waste
stream for the process. No data are reported for fume emissions.
6. Control Technology - No controls of atmospheric fume emissions are
reported. The exhaust gas could be routed to blast furnace baghouses.
215
-------
Table 3-23. TYPICAL COMPOSITIONS OF SOFTENED LEAD BULLION AND SLAG
(AMOUNTS IN WEIGHT PERCENT)
1
Constituent
Pb
Cu
Se
Te
As
Sb
Sn
Ag
Au
Softened lead
bullion
0.004
0.001
0.025
0.15
1.25a
Softened slag
(liquid dross)
75.
0.005
tr
tr
1.7
12.0
tr
tr
tr
Value for gold in grams per metric ton.
tr - trace.
216
-------
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA 440/1-
75/032-a. February 1975.
217
-------
PRIMARY LEAD PRODUCTION , PROCESS NO. 11
Kettle Softening
1. Function - Arsenic, antimony, and tin may be removed from the
drossed lead bullion by kettle softening. Other softening methods are
reverberatory (Process No. 10) and Harris (Process No. 12). Unlike the
reverberatory method, in which air is blown through molten bullion, the
kettle method entails addition of oxidizing agents to remove impurities.
Application is usually limited to treating bullion with a relatively low
impurity content, 0.3 weight percent or less antimony equivalent.1*'2
Drossed bullion is charged to a kettle and melted. Oxidizing
fluxes such as caustic soda (NaOH) and niter (NaNOs) are then added
while the charge is agitated. The fluxes react with the impurities to
form a series of salts such as sodium antimonate (NaSb03)J>3 A slag
containing the oxidized impurities results from the reactions. When the
reactions are complete, agitation of the kettle is stopped and the slag
rises to the top of the kettle, where it is skimmed off. The purified
lead bullion is sent to the desilvering process. Composition of the
softened bullion is similar to that shown in Table 3-23 (Process No.
10); residual hardness is less than 0.03 weight percent antimony equiva-
lent.
2. Input Materials - In addition to the drossed lead bullion, caustic
soda and sodium nitrate (niter) are required for fluxing, in amounts
continguent on impurity content of the feed bullion. A slight excess
over stoichiometric requirements is desirable for effective removal of
impurities.
3. Operating Conditions - A kettle temperature of 700°C is required.
Pressure is atmosphericJ
4. Utilities - Electricity is used to power the process equipment,
such as agitators and conveyors.
Gas or oil are used to heat the kettle and maintain temperature.
5. Waste Streams - The slag containing the oxidized impurities is
discarded.This material contains lead as well as water-soluble sodium
oxide salts of arsenic, tin, and antimony. The amount of slag generated
is not reported, but it represents a hazardous waste from this industry.
6. Control Technology - Slag is dumped with that generated in either
the blast furnace or fuming furnace.2 There is no recognized control
technology for disposition of this slag. Substitution of Harris soft-
ening (Process No. 12) is recommended.
218
-------
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
3. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/032-a. February 1975.
219
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 12
Harris Softening
1. Function - Removal of arsenic, antimony, and tin from drossed lead
bullion can be accomplished by the Harris softening process. Alternate
softening processes are reverberatory (Process No. 10) and kettle
(Process No. 11). As with kettle softening, the process is most appli-
cable in treating bullions containing 0.3 percent or less antimony.
The process consists of two operations. The initial pyrometal-
lurgical step is the same as the kettle softening process. The drossed
bullion is charged to a kettle, melted, and agitated. Sodium nitrate
and sodium hydroxide are added to react with the impurities and form
sodium oxide salts. A slag containing these impurities is skimmed off,
and the purified lead is sent to the desilvering processJ
The second operation is hydrometallurgical treatment of the cooled
slag. The slag is crushed and leached with hot water to dissolve the
sodium salts. After filtration, the depleted slag filter cake is dis-
carded, and the solution is cooled to precipitate sodium antimonate
preferentially. After separation by filtration, the antimony-rich
filter cake is subjected to further processing (Process No. 13), and the
filtrate is mixed with lime to precipitate calcium salts of arsenic and
tin in separate operations.
Upon removal from solution, the calcium arsenate is reported to be
sold to insecticide manufacturers, and the calcium stannate is shipped
to tin producers. The residual sodium hydroxide solution is evaporated
to produce dry sodium hydroxide, which is recycled to the softening
process.
2. Input Materials - Aside from the drossed lead bullion, sodium
hydroxide and sodium nitrate are required in slightly more than stoichi-
ometric amounts for fluxing.
Process water is fed to the operation for slag leaching, in un-
specified amounts.
Lime is required in quantities sufficient to precipitate salts of
arsenic and tin.
3. Operating Conditions - Maximum temperatures are 700°C for the
pyrometallurgical operation, 100°C for hydrometallurgical processing,
and more than 200°C for sodium hydroxide evaporation.2 Atmospheric
pressure is employed for all operations.
220
-------
4. Utilities - Process equipment, such as agitators, pumps, and con-
veyors are powered by electricity.
Gas or oil is utilized for kettle heating, temperature maintenance,
and sodium hydroxide recovery.
5. Waste Streams - The leached slag constitutes the only waste stream
for the process. It is not water soluble.! Composition and quantity of
slag are not reported.
6. Control Technology - The leached slag is dumped together with slag
from the blast furnace or fuming furnace.
7. EPA Source Classification Code - None
8. References -
1. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
221
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 13
Antimony Recovery
1. Function - Softening slags are processed to separate and recover
the contained mineral values, notably arsenic, tins, and antimony.
Depending upon the product desired, two methods are commonly used to
recover antimony.
If antimonial lead or "hard lead" is desired, the softening slag is
subjected to a reduction process. The slag is charged to a furnace and
heated with a reducing agent and slagging fluxes. Since the oxidation
potential of the other contained minerals is higher, the oxides of lead
and antimony are preferentially reduced. Slag is formed and skimmed off
while the remaining lead and antimony metal is tapped as a marketable
product. If the slag is rich in tin, it may be sold to a tin producer;
otherwise, it is recycled to the sintering process or blast furnace.'
If the desired product is antimony trioxide (Sb203), the .softening
slag feed is treated by a volatilization process. The slag is fed to a
furnace, where it is heated to volatilize arsenic trioxide and antimony
trioxide. Since arsenic trioxide is more volatile, it is driven off
first and is separated from the antimony trioxide by selective con-
densation. 2 Collection of the oxides consists of condensing the vola-
tilized fume and capturing it in an electrostatic precipitator or a
filter baghouse. The recovered antimony trioxide is sent to antimony
refining plants, usually located nearby. Recovered arsenic trioxide may
be sold to arsenic processors. The nonvolatilized furnace residue, con-
taining appreciable lead values, is returned to the blast furnace or
sintering process.
2. Input Materials - Slag is obtained from the softening processes.
The slag contains primarily lead, arsenic, antimony, and tin. A typical
analysis is presented in Table 3-23 (Process No. 10).
For hard lead production, coke or charcoal is used as a reductant,
the quantity dependent on feed slag composition. Soda ash or silica is
used as a flux. No quantitative data are reported.
3. Operating Conditions - Temperatures range from 800° to 900°C.
Pressures are atmospheric.2
4. Utilities - Gas or oil is used to maintain furnace temperatures.
Electricity is required for operation of conveyors.
In antimony trioxide production, cooling water is required for fume
condensation.
222
-------
No quantitative data are furnished in the literature.
5. Waste Streams - In the volatilization process, the air stream
carrying the oxide fume is exhausted to the atmosphere after condensa-
tion and collection.
The arsenic trioxide, if it cannot be sold, represents the only
solid waste from the process.
6. Control Technology - In the volatilization and condensation pro-
cess ,~Th^~Tuliie~Ttr^ampasses through an electrostatic precipitator or
filter baghouse.3 Details on these collection devices are given in
Process No. 3 in the copper section, Section 2.
No established control for excess arsenic trioxide has been de-
veloped.
7. EPA Source Classification Code - None
8. References -
1. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-72-274a. September 1973.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/032-a. February 1975.
223
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 14
Parkes Desilverizing
1. Function - Gold and silver are removed from the softened lead
bullion by the Parkes desilverizing process. Since these metals do not
oxidize easily, they are not removed by any of the previous refining
steps. The Parkes desilverizing process is based on the fact that gold
and silver have a greater affinity for zinc than for lead. Therefore,
zinc is added to the molten lead bullion to form alloys with the copper,
gold, and silver contained in the bullion. These alloys are insoluble
in the lead and rise to the surface to form a crust that is skimmed off.
To simplify subsequent recovery processes, the gold and silver are
often recovered in separate steps. Since gold and most of the copper
are first to combine, zinc is added in two steps. The initial addition
results in a crust rich in gold values.U2 Upon removal of this crust,
the second addition of zinc is made to allow the removal of silver.
Although these steps are not totally exclusive for either gold or
silver, they do effect a good degree of segregation.
Because other metallic impurities, notably arsenic, can interfere
with this process, they must be removed prior to this operation. Ar-
senic content in the bullion must be less than 0.10 weight percent.'
2. Input Materials - Softened lead bullion is required for the pro-
cess. Hardness equivalent to less than 0.03 combined weight percent of
arsenic, antimony, and tin is desirable. A typical analysis for a
softened bullion is presented in Table 3-23, Process No. 10.
Zinc is the only additive. The amount is 1 to 2 percent in excess
of the amount required to saturate the lead bullion, i.e., 0.55 weight
percent of bullion weight.
3. Operating Conditions - The bullion charge is heated to 54C°C and
then cooled to 40° to 93°C.1>3 Pressure is atmospheric.
4. Utilities - Gas or oil is used to heat the charge. Electricity is
used to operate pump and agitators. Utility quantities are not given in
the literature.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
224
-------
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
3. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/032-a. February 1975.
225
-------
PRIMARY LEAD PRODUCTION
PROCESS NO. 15
Retorting
1. Function - Crusts from the Parkes process are retorted to recover
the zinc for reuse in the desilvering process and to form a purified
dore" metal. Dross is placed in graphite crucibles of 600 to 640 kilo-
gram capacity, which are heated in a special Faber du Faur type furnace.
Zinc is vaporized to be condensed in a cooling chamber, then tapped to
bars or blocks and recycled. The retort metal remaining assays 140 to
400 kilograms dore" per ton of dross. 1 Table 3-24 presents a typical
analysis of this retort metal.
Table 3-24. TYPICAL RETORT ANALYSIS
1
Constituent
Do re*
Zinc
Arsenic
Antimony
Amount, %
156-438a
1.5-2.5
0.4
1.0
Constituent
Copper
Tellurium
Bismuth
Lead
Amount, %
1.5-4.0
0.2
0.25
balance
Values for dore" are in kg/metric ton.
2. Input Materials - Crusts from the desilverizing process are the
only input material for this process. This dross is basically a gold-
silver-zinc compound with small amounts of impurities such as antimony,
copper, tellurium, bismuth, and lead.
3. Operating Conditions - Operating temperatures during retorting are
between 1260 and 1320°C.l Pressure is atmospheric.
4. Utilities - The retort furnaces are gas- or oil-fired. Electricity
is consumed by transport apparatus. No quantities are cited.
5. Waste Streams - The only waste stream consists of small quantities
of metallic fume escaping the condensing chamber. This fume is believed
to be composed predominately of zinc, arsenic, antimony, and lead. No
data were found on emission factors or constituents.
6. Control Technology - Several smelters control fume and particulate
emissions with baghouses. Because of the temperature limits on these
baghouses, the gases are previously cooled in tubular cooling chambers.
Control efficiencies of more than 98 percent are claimed. The collected
flue dusts are recycled to the sintering machine. Other smelters use no
control devices on retorts.
226
-------
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division on John Wiley and Sons, Inc., 1967.
227
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 16
Cupelling
1. Function - Retort metal is purified in the process called cupel-
ling. In a furnace called a cupel, the molten metal charge is suc-
cessively blown with air and slagged to remove impurities. The dif-
ference in oxidation potentials of the impurities allows sequential
removal of slags with distinct characteristics.
Zinc, arsenic, and antimony are oxidized first and removed, and
most of the lead content oxidizes next, forming a product called "good
litharge". Upon removal, it is recycled to the softening process for
use as an oxidizer. Bismuth, copper, and tellurium accumulate in the
dore1 until the final stages of cupelling, when they oxidize to a slag
called "coppery litharge" since copper content may be as high as 10
weight percent. Oxidizing agents are required to remove the last traces
of copper and tellurium from the dore*. These latter slags are returned
to smelting for further processing.
Cupels are rated according to dore" output. Usual capacities range
from 2,850 to 11,300 kg per charge. When impurity removal is complete,
the remaining gold-silver alloy is cast into bars for marketing. Purity
is 99.9 percent.
2. Input Materials - Retort metal is the basic feed to the process. A
typical analysis of the metal is given in Table 3-24, Process No. 15.
"" s
Sodium nitrate and silica flour are added to remove the last traces
of copper and tellurium from the dore1. Amounts depend upon residual
levels of the impurities.
3. Operating Conditions - Temperature of the cupel is raised to
1150°C7lPressure is atmospheric.
4. Utilities - Gas or oil is used to heat the furnace. Pumps and
agitators are electrically powered. Cooling water is pumped through
jacketed furnace sections. Compressed air is injected into the charge
for oxidation of impurities. A cold air stream is also blown across the
face of the bath at a pressure of 70 to 87 g/cm2 to cause rippling,
which hastens oxidation. No quantitative data are given.
5. Waste Streams - Exhaust gases from the process are at a temperature
of 1000° to 1100°C and contain metallic vapors (fume) as well as parti-
culates. Zinc, lead, arsenic, and antimony comprise the fume. Particu-
lates may contain any of the components listed in Table 3-24 (Process
No. 15). Emission data are not given in the literature.
228
-------
6. Control Technology - Several smelters control exhaust gases,
cooling them by passage through tubular cooling chambers before routing
them to fabric filter baghouses. Collection efficiency greater than 98
weight percent is claimed. Collected dust is recycled to the smelter
for further processing. Other smelters do not control cupel emissions.
7. EPA Source Classification Code - None '
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
229
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 17
Vacuum Dezincing
1. Function - Zinc added to the bullion in the desilverizing process
is removed by dezincing. The vacuum distillation method is used most
widely in the industry because the recovered metallic zinc can be
directly recycled to the desilvering process. Alternate methods are
chlorine dezincing (Process No. 18) and Harris dezincing (Process No.
19).
Desilverized lead is charged into a kettle and heated. An inverted
bell is placed on top of the kettle with its skirt projecting into the
molten lead to form a vacuum seal. A vacuum is drawn in the bell and is
held for about 2.5 hours, during which the bath is agitated to bring the
zinc to the surface. The zinc vaporizes and is condensed on the water
cooled dome of the bell. On completion of the process, the vacuum is
broken, the bell removed, and the solidified zinc peeled from the sur-
face of the bell.
The product bullion typically analyses less than 0.001 weight
percent (10 ppm) zinc, 0.0003 weight percent (3 ppm) antimony, and 0.15
weight percent bismuthJ The bullion is sent for debismuthizing or for
casting if bismuth content is low.2
2. Input Materials - Desilverized lead bullion typically containing
0.5 to 1.0 weight percent zinc, is the only feed material.3
3. Operating Conditions - The molten lead bath is maintained at 580 to
595°cU£ with an operating pressure of 50 to 60 microns absolute.'
4. Utilities - Gas or oil is used to maintain the kettle temperature.
Pumps, agitators, and conveyors are electrically powered. Cooling water
is used to remove heat from the jacketed bell surface. Although no
quantitative data are given in the literature, energy consumption is
higher than in other dezincing processes because of the higher tempera-
ture requirements.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
230
-------
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
3. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/175-032-a. February 1975.
231
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 18
Chlorine Dezincing
1. Function - Desilverized lead bullion requires removal of the zinc
added during the desilverizing process. A process known as chlorine
dezincing, or the Betteston process, can be used, as well as vacuum
dezincing (Process No. 17) and Harris dezincing (Process No. 19).
i
Molten lead is recirculated with a pump from a heated kettle
through a reaction chamber into which gaseous chlorine is injected from
a chlorine tank. Since reactivity of zinc with chlorine is greater than
that of lead, zinc chloride is formed in the reactor and subsequently
collects on the surface of the molten lead. The material skimmed from
the lead contains small amounts of lead chloride and mechanically
entrained lead prills. After treatment with metallic zinc for lead
removal and recovery, a marketable by-product analyzing 99 percent zinc
chloride is obtained. The dezinced lead bullion is sent for debis-
muthizing or casting.
A 180-metric-ton kettle with an overall cycle of about 8 hours
typically produces 16,300 metric tons of dezinced bullion per month.
The bullion contains 0.005 weight percent (50 ppm) zincJ
2. Input Materials - The desilverized lead bullion feed contains 0.5
to 1.0 weight percent zinc.
Chlorine is injected at a rate of 180 to 225 kg per hour into
molten lead recirculated at a rate of 7 to 11 metric tons per minute.1
3.
a mol
Operating Conditions - A temperature of 370°C maintains the lead in
ten condition. Pressure is atmosphericJ
4. Utilities - The kettle is heated with oil or gas. Electricity is
used for pumps, agitators, and conveyors. No data regarding consumption
are cited.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
232
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 19
Harris Dezincing
1. Function - The zinc added to the lead bullion for desilverizing
requires removal by a dezincing process. The Harris dezincing process,
consists of a pyrometallurgical step followed by a hydrometallurgical
procedure. Alternate dezincing processes are vacuum (Process No. 17)
and chlorine (Process No. 18).
For the pyrometallurqy, the equipment is the same as for Harris
softening (Process No. 12), i.e., a charging kettle, a reaction cyl-
inder, and a molten lead pump. In a typical cycle, desilverized lead
bullion is charged to the kettle and then pumped in a molten state
through the reaction cylinder, which contains a small amount of caustic
soda saturated with lead oxide. The saturated caustic remains from the
previous dezincing cycle. Upon contact with the molten lead, the lead
oxide in the caustic reacts with the zinc to form zinc oxide, which
reacts with caustic to form sodium zincate. After 30 minutes of lead
recirculation, pumping is stopped, fresh caustic is added to the cyl-
inder to maintain salt fluidity, and the cylinder is dumped to a granu-
lator tank. Fresh molten caustic is again pumped to the cylinder and
pumped recirculation of the lead bullion is resumed. The final caustic
addition will become saturated with zinc and lead oxide and is held over
for the next cycle. When dezincing is completed, the product contains
less than 0.001 weight percent (10 ppm) zinc and 3 ppm antimony. The
dezinced lead is pumped from the kettle for debismuthizing or casting.
The so-called spent salts from the granulation tank are treated
hydrometallurgically. After granulation and solution in hot water,
sodium zincate hydroloyzes to zinc oxide and sodium hydroxide. The zinc
oxide precipitates from the solution and is recovered by filtration. It
is subsequently dried and sold. The sodium hydroxide solution is evap-
orated and the resulting anhydrous caustic recycledJ
2. Input Materials - Desilverized lead containing 0.5 to 1.0 weight
percent lead is charged to the process.
Anhydrous sodium hydroxide is required, the amount dependent upon
the zinc content of the feed bullion.
Water is used for the hydroextraction of by-products.
No data as to quantities are given.
233
-------
3. Operating Conditions - A temperature of 540°C is required for the
pyrometallurgical operation. The temperature for hydroextraction is
100°C. Evaporation of the sodium hydroxide solution requires tempera- •,
tures above 200°C. All operations are performed at atmospheric pressure.
4. Utilities - Kettle heating and soda evaporation of caustic are done
by burning gas or oil. Electricity is used to operate pumps, agitators,
and conveyors.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
234
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 20
Debismuthizing
1. Function - When the dezinced lead bullion contains 0.15 weight per-
cent or more bismuth, it must be debismuthized before the final refining
and casting process. The debismuthizing procedure is called the Betterton-
Knoll process.
Calcium and magnesium are added to the molten lead to form ternary
compounds (e.g., CaMg2Bi2) with the bismuthJ The compounds have a
higher melting point than lead, and a lower density. Therefore, when
the temperature of the mixture is reduced to a point just above the
melting point of lead, the metallic compounds solidify to form a dross
that can be skimmed from the lead. To enhance the physical separation,
antimony or organic agents are somtimes added, further reducing the
bismuth content.
The purified lead is pumped to the casting operation, and the
skimmed metallic compound is sent to bismuth recovery.
Cupel slags rich in bismuth may be similarly treated, the residual
lead being returned to smelting.
2. Input Materials - Dezinced lead bullion fed to the process typi-
cal lylispyT~TO~ppiTrzinc, 3 ppm antimony, and 0.15 weight percent
bismuth.2
The amounts of calcium and magnesium added depend on the amount of
bismuth to be removed. Twice as much calcium is added as magnesium.
Cupel slags are added when bismuth content is high enough (20 to 25
weight percent) to warrant recovery.
Antimony or organic compounds are added as needed to improve bis-
muth separation. The literature does not specify the compounds or the
amounts.
3. Operating Conditions - The molten lead bath is maintained at 500°
to 550bcl»3 for calcium-magnesium addition and is then cooled to 350°C
for dross separation. Pressure is atmospheric.
4. Utilities - A small amount of gas or oil is required to maintain
the lead bath temperature. Pumps and agitators are electrically pow-
ered.
5. Waste Streams - None
235
-------
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/0'32-a. February 1975.
2. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
236
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 21
Bismuth Refining
1. Function - The dross generated by debismuthizing is treated to
recover bismuth. The material is placed in a furnace, where it is
melted. Chlorine gas is injected, and the calcium, magnesium, and
entrained lead combine with the chlorine to form chlorides, while the
bismuth does not. The chlorides form a solid slag that is skimmed from
the surface of the molten bismuth. Air is then blown through the bis-
muth and a caustic soda flux is added to oxidize any residual impuri-
ties. After slag removal, the metal, which is now 99.99 percent bis-
muth, is cast into marketable shapes and soldJ
2. Input Materials - The slag from the debismuthizing process is
chiefly composed of ternary compounds of calcium, magnesium, and bis-
muth.
Chlorine constitutes about 25 percent of the weight of slag charged
to the furnace.
Caustic soda flux is used in variable amounts for oxidation of
impurities.
Charcoal is used as a cover during casting to maintain the bismuth
in a reduced state.
3. Operating Conditions - Bath temperature during chlorination is
500°C. Subsequent temperatures for oxidation and casting are lower.
Pressure is atmosphericJ
4. Utilities - Gas or oil is used for heating and maintance of tem-
perature^Electricity is used to run pumps and agitators. Compressed
air is furnished to oxidize impurities. The literature does not state
the quantities required.
5. Waste Streams - Exhaust gases to the atmosphere contain chlorine
and fume. No emission quantities were found.
Slag composed of calcium, magnesium, and lead chlorides is 40
weight percent of the dross fed to the process. Final oxidation gen-
erates a soda slag, the amount of which is not specified. These slags
contain toxic water-soluble materials.
6. Control Technology - Control of atmospheric emissions is not
practiced.
237
-------
Slag wastes are discarded with the slag generated in smelting.
Further information is given in Process Nos. 5 and 6. The chloride
salts contained in the slag are very water-soluble and hence are easily
leached into adjacent water supplies. Existing practice does not re-
present good control technology.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
238
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 22
Filial Refining and Casting
!• Function - Refined lead bullion from dezincing or debismuthizing is
given a final purification and cast into ingots. The refined lead is
fluxed with oxidizing agents to remove remaining impurities such as lead
oxide and magnesium or calcium residues. After slag removal by skim-
ming, the purified lead, assaying 99.999 percent purity, is reheated and
sent to the casting operation, where it is formed into ingots or pigs.
Most casting is perfomed by fully automated machines. The slag is
recycled to the blast furnaceJ
2. Input Materials - Refined bullion containing a small amount of
impurities is fed to the process. Caustic soda and sodium nitrate are
used as oxidizing flux in amounts varying with the degree of lead
contaminationJ »*
Water is used to cool the cast lead ingots by direct contact, at
rates ranging from 300 to 1,500 liters per minute.3
3. Operating Conditions - Temperatures for final purification range
from 370° to 500°C; casting is at 540°cJ»4 Pressure is atmospheric.
4. Utilities - Gas or oil is required for heating and maintenance of
temperature.
Pumps and agitators are electrically operated. The literature
gives no data on quantities.
5. Waste Streams - Water used to cool the lead castings by direct
contact becomes contaminated with particulate matter, including lead and
lead oxides.
6. Control Technology - Several methods are used to handle the con-
taminated cooling water. The water is either recycled for use in slag
granulation or is sent to a tailings pond for settling of suspended
solids. A variation of the latter method, liming the effluent for
precipitaton of solids, is also practiced.3
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. New York, Interscience
Publishers, a division of John Wiley and Sons, Inc., 1967.
239
-------
2. Development Document for Interim Final Effluent Limitations,
Guidelines and Proposed New Source Performance Standards for
the Lead Segment of the Nonferrous Metals Manufacturing Point
Source Category. Environmental Protection Agency, EPA
440/1-75/0'32-a. February 1975.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. Environmental Protection Agency, Washington,
D.C., EPA-R2-73-274a. September 1973.
4. Trace Pollutant Emissions from the Processing of Metallic
Ores. PEDCo-Environmental Specialists, Inc. August 1974.
240
-------
4.0 ZINC INDUSTRY
INDUSTRY DESCRIPTION
The major product of the primary zinc industry is metallic zinc;
the industry also produces zinc oxide, sulfuric acid, cadmium, and
occasionally other chemicals such as zinc sulfate. For the purpose of
analysis, the zinc industry is considered in segments: pyrometallur-
gical zinc production, electrolytic zinc production, zinc oxide produc-
tion, and cadmium recovery. Production of other by-product compounds
such as germanium, thallium, gallium, and indium is not considered as
part of this industry because this is not done at primary zinc smelters.
Generally, ore is mined and concentrated at one location and then
transferred to smelters for the production of zinc, zinc oxide, or both.
Cadmium is normally recovered at smelters from collected dusts and slags
with sufficient cadmium content. Direct zinc oxide production uses the
same ore concentrate as metallic zinc production.
In 1975, approximately 6,700 people were employed in lead-zinc
mining and milling operations, including underground and surface mines
and mills. Zinc smelters, including secondary smelters, employed 4,100
people in 1975, including workers involved in by-product processes.
Table 4-1 shows mining, production, and consumption totals for zinc and
cadmium in the years 1971 to 1975. These data indicate that zinc and
cadmium production have remained fairly constant over the past few
years. Consumption has exceeded production, and imports and shipments
from government stockpiles have made up the difference. The reduced
rate of economic activity led to a 26 percent drop in slab zinc consump-
tion in 1974-1975.l Imports of cadmium metal have been increasing.
Mining production was reported in 18 states scattered throughout
the U.S. Major producing states in 1975 were Tennessee, 18 percent;
Missouri and New York, 16 percent each; Colorado, 10 percent; Idaho, 8
percent^and New Jersey, 7 percent.' Mining capacities ranged from
10,000 to 10,000,000 metric tons per year of lead-zinc ore. The 25
largest U.S. mines accounted for over 90 percent of the zinc ore mined
in 1975, the five largest alone accounting for 40 percent.! Available
information on these mines is given in Table 4-2. A total of 43 zinc
mines operated in 1975.3
241
-------
Table 4-1. MINING, PRODUCTION, AND CONSUMPTION1 OF ZINC AND CADMIUM
(metric tons)
1975 (est)
1974
1973
1972
1971
Mine production
430
454
434
434
456
Primary slab
zinc production
404
503
529
574
695
Slab zinc
consumption
862
1168
1364
1286
1137
Cadmium production
(includes secondary*)
2.3
3.0
3.4
3.7
3.6
Cadmi urn
consumption
4.2
5.4
5.7
5.7
4.9
There is no significant recycling of cadmium metal. All uses other than nickel-cadmium
batteries and alloys are dissipative.
-------
Table 4-2. TWENTY-FIVE LEADING ZINC MINES IN THE UNITED STATES2'3
NJ
*»
co
Mine
Balmat
Bulck
Sterling
Bunker Hill
Eagle
Z1nc Mine Works
Blue Hill
Friedensville
New Market
Burgin
Austinville &
Ivanhoe
Star Unit
Leadville
Ground Hog
Jefferson City
Edwards
Shullsburg
Location
St. Lawrence. N.Y
Iron, Mo.
Sussex, N.J.
Shoshone, Id.
Eagle, Co.
Jefferson, Tn.
Hancock, Maine
Lehigh, Pa.
Jefferson, Tn.
Utah, Utah
Wythe, Va.
Shoshone, Id.
Lake, Co.
Grant, N.M.
Jefferson, Tn.
St. Lawrence, N.Y.
LaFayette. Wise.
Company
St. Joe Minerals Corp
Amax Lead Co. of Mo.
New Jersey Zinc Co.
Bunker Hill Co.
New Jersey Zinc Co.
U.S. Steel Corp.
Kerramerica Inc.
New Jersey Zinc Co.
ASARCO
Kennecott Copper
New Jersey Zinc Co.
Bunker Hill Co. &
Hecla Mining Co.
ASARCO
ASARCO
New Jersey Zinc Co.
St. Joe Minerals Corp.
Eagle-Picher Ind. Inc.
Opera-
tions
B
B
B
C
B
B
B
B
B
-
B
B
B
A
B
B
-
Type of
mining
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Underground
Underground
-
Underground
Underground
Underground
Underground
Underground
Underground
-
Ore
grade,
percent
-
1.7
-
7-9 Pb-Zn
-
-
-
-
3
-
-
-
16
16
-
-
-
Annual
ore tonnage
1 - 10 million
1 - 10 million
198,000
691 ,000
206.000
100,000-500,000
-
400,000
500,000-1 million
-
596,000
282,000
100,000-500,000
500,000-1 million
390,000
-
-
Products
Lead, zinc
Lead, zinc cone.
refined lead,
sulfuric acid
Zinc ore.iron-
niangenese ore
Lead, zinc, silver
Zinc cone. .lead
cone, copper-
silver ore
Zinc
Zinc, copper
Zinc cone. ,
agricultural
limestone
Zinc cone.
•
Zinc cone. .lead
cone.
Zinc, lead, silver
Lead, zinc, si Tver
Zinc, lead, copper
silver
Zinc conc.,agric.
limestone
Lead zinc
-
Mineralization
_
Galena, Sphalerite.
Chalcopyrite
Zinc oxide, and
silicates
Lead-zinc-silver
Marmatite, Galena
_
_
Sphalerite
•
_
Sphalerite, Gelena
-
Sul fides
Sphalerite. Gelena
Sphalerite
_
-
-------
Table 4-2 (continued). TWENTY-FIVE LEADING ZINC MINES IN THE UNITED STATES
2 3
£.,0
ro
Mine
Idarado
Bruce
Sunnyside
Rend Oreille
Ozark
Young
Magmont
Viburnum 129
Location
San Miguel, Co.
Yavapai, Az.
San Juan, Co.
Pend Oreille, Wa.
Reynolds, Mo.
Jefferson, Tn.
Iron, Ho.
Washington, Mo.
Company
Idarado Mining Co.
Cypress Mines Corp.
Standards Metals Corp
Pend Oreille Mines &
' Metals Co.
Ozark Lead Co.
ASARCO
Cominco American Inc
St. Joe Minerals Corp.
Opera-
tions
B
B
B
B
e
"•
Type of
mining
Underground
Underground
/
Underground
Underground
Underground
—
Ore
grade,
percent
6.34
-
4-5 Pb-Zn
8
-
Annual
ore tonnage
388,000
192,000
100,000-500,000
1-10 million
1,042,000
-
Products
Zinc, lead, copper
gold, silver,
cadmium
Gold, lead, zinc,
silver, copper,
cadmium
Zinc, lead, silver
Lead & zinc
cone.
Zinc conc.icop-
per cone, silver
-
Mineralization
Zinc, copper, lead.
gold, silver,
cadmium
Galena, Sphalerite
Lead, zinc
Galena, Sphalerite
Galena, Sphalerite
Chalcopyrite
-
-------
Direct zinc oxide production is considered part of the primary zinc
industry since the process involves the reduction of zinc concentrates
followed by oxidation. Indirect zinc oxide production is not discussed,
since slab zinc is used as the raw material.
All of the zinc smelters produce cadmium. Collected flue dusts
from roasting and sintering operations and cadmium-containing materials
from refining and precipitation processes are treated hydrometallurgi-
cally for cadmium recovery at the zinc smelter.
Other metals recovered as by-products from zinc ore are germanium,
thallium, indium, and gallium. This processing, however, is not done at
any primary zinc smelters in the U.S. The waste materials containing
these by-product metals are shipped as intermediate products or are
disposed of as waste if the by-product content is not sufficient for
recovery.
In 1970 the U.S. Bureau of Mines predicted that primary zinc demand
by the year 2000 will be between 1.90 and 3.63 million metric tons.
However, the decline in primary zinc production, beginning in 1969,
indicates that the future demand may not be as high as expected. Much
of this decline is related to sale of some of the domestic zinc stock-
piles and to the increasing use of foreign zinc supplies.
Primary domestic smelting capacity has declined 47 percent since
1968, with the closing of eight smelters due to outdated equipment and
environmental problems.4
During 1975, one company indefinitely postponed the construction of
a 163,000-metric ton-per-year electrolytic refinery, whereas another
company planned to proceed with construction of an 82,000-metric ton-
per-year electrolytic refinery in partnership with a Belgian firm.
Raw Materials
Zinc is usually found in nature as the sulfide called sphalerite,
which has a cubic lattice structure and is commonly referred to as zinc
blende, blende, or jack. Zinc contents can be 67.1 percent in the pure
phase. A polymorph of sphalerite, wurtzite, has a hexagonal structure
and is more stable at elevated temperatures. Almost all other zinc
minerals have been formed as oxidation products of these sulfides. A
list of the most common zinc minerals is presented in Table 4-3. Most
of these oxidized minerals are minor sources of zinc, although frank-
linite and zincite are mined for their zinc content at the New Jersey
Zinc Co. mine.5
Iron is the most common impurity or associated metal of zinc, owing
to the chemical similarities and relative ease of substitution in their
respective lattices.
245
-------
Table 4-3. COMMON ORES MINED FOR THEIR ZINC CONTENT^
ZnO
ZnS0 • 7H0
ZnC0
H 2V
3
Zn4Si207(OH)2 •
(Zn2Mn)0 • Fe203
2ZnC03 • 3Zn(OH)2
ZnS
(Fe,Zn)S
Zincite
Goslarite
Smithsonite (or calamine in Europe)
Hemimorphite (or calamine in America,
called electric calamine in Europe)
Willemite
Franklinite
Hydrozincite
Sphalerite, wurtzite
Marmatite
246
-------
A sulfide zinc ore with a ratio of Fe: Zn above 1:8 is known as
marmatite.
Cadmium is the second most abundant impurity of zinc. It is always
associated with zinc, and is usually present as greenockite (CdS).
Complete solid solutions exist between zinc and cadmium sul fides,
but the cadmium content rarely exceeds 1 or 2 percent.
In zinc ores, commonly associated non-zinc minerals are calcite
(CaC(h), dolomite (Ca,Mg)C03, pyrite and marcasite (Fe$2), quartz
chalcopyrite (CuFeS2), and barite (83804).
Zinc ores are processed at the mine to form concentrates containing
typically 52 to 60 percent zinc, 30 to 33 percent sulfur, and 4 to 11
percent iron. 6 Roasting at the plant lowers the sulfur content to about
2 percent. Other raw materials are required at the smelter for pro-
ducing metallic zinc. Coke or coal and sand along with inert materials
are required during pyrometallurgical sintering, in quantities depending
upon the specific concentrate properties and the desired characteristics
of the sinter. Coal or coke must also be added during reduction. Exact
quantities are variable, depending on the properties of the feed and
type of reduction furnace. In hydrometallurgical processing, sintering
or pyrometallurgical reduction is not required but sulfuric acid is
required for leaching the calcine.
The energy required for production of one metric ton of slab zinc
is 8.8 to 16.4 million kg-cal, depending on the process. In 1972, 9.6 x
109 kg-cal were used in the manufacture of primary zinc slabs, an average
of 15.3 million per metric ton. 7 This represents an 9 percent increase
in efficiency of energy utilization over 1967, primarily due to the
closing of inefficient horizontal retorts.
Products
The products of the primary zinc industry are metallic zinc, zinc
oxide, and cadmium. Uses for these products are widely varied. Metal-
lic zinc is used for galvanizing, for making pigments and zinc com-
pounds, for alloying, and for grinding into zinc dust. Table 4-4 shows
U.S. consumption of slab zinc for 1975. Usage patterns in the U.S.
differ from those in the rest of the world in the heavy emphasis on
zinc-base alloy castings, mainly for the automotive industry. The
primary product of most zinc companies is slab zinc, which is produced
in five grades and classified by its purity. These grades are presented
in Table 4-5.
247
-------
Table 4-4. U.S. SLAB ZINC CONSUMPTION - (1975)°
Galvanizing
Brass products
Zinc-base alloy castings
Rolled zinc products
Zinc oxide
Other
Metric tons
341,906
104,622
303,173
24,773
35,398
29,572
839,444
Percent
41
12
36
3
4
4
100
Table 4-5. GRADES OF COMMERCIAL ZINC
Composition, percent
Special high grade
High grade
Intermediate
Brass special
Prime western
Zinc
99.990
99.90
99.5
99.0
98.0
Lead
0.003
0.07
0.20
0.6
1.6
Iron
0.003
0.02
0.03
0.03
0.05
Cadmium
0.003
0.03
0.40
0.50
0.50
248
-------
Zinc oxide is used in rubber, emollients, ceramics, and fluorescent
pigments, and in the manufacture of other chemicals. Metallic cadmium
is used in production of alloys, in corrosion-resistant plating for
hardware, as a counter electrode metal for selenium rectifiers, as
neutron shielding rods in nuclear reactors, in nickel-cadmium batteries,
and in plastics and cadmium compounds. Cadmium metal accounts for 60 to
70 percent of consumption, and cadmium sulfide used for pigments for
another 12 to 15 percent.*
Companies
The six primary zinc smelters and one zinc oxide plant that use
primary concentrate feed are listed in Table 4-6 along with their capa-
cities. The three pyrometallurgical plants range in capacity from
57,000 to 227,000 metric tons per year. The smallest of the pyrometal-
lurgical plants is currently being replaced by an electrolytic facility
of 45,000 metric tons per year capacity. Completion is expected in
1977. Upon its completion, U.S. primary zinc capacity will be about
equally divided between pyrometallurgical and electrolytic processes.
The single largest plant accounts for 35 percent of the total domestic
primary capacity. All of the primary smelters produce sulfuric acid as
a by-product. In addition, three of these were responsible for over 75
percent of primary cadmium production in 1975.1
Environmental Impact
Uncontrolled atmospheric emissions from the primary zinc industry
have decreased in recent years because of the increase in electrolytic
zinc recovery operations and the closing of several retort zinc smelt-
ers. In 1969, zinc emissions to the atmosphere totalled 65 metric tons
from mining and milling, and 45,350 metric tons from metallurgical
processing.^ Sulfur dioxide is emitted from roasting and sintering
operations, although the roaster S02 emissions are normally collected to
produce sulfuric acid.
Smelter solid wastes are usually recycled for by-product metal
recovery. Some sludges may require several months' storage before
recycling, and others may be disposed of as landfill. Sludges and solid
wastes generated during mining and concentration processes are disposed
of at the mining site in tailings ponds or as mine backfill.
Liquid effluents can be classified as non-contact and contact.
Noncontact water is used for cooling in heat exchangers. Contact, or
process, wastewater is produced in such operations as scrubbing of
roaster gas and reduction furnace gas, cooling of metal castings,
cadmium production, and auxiliary air pollution controls. The pollu-
tants of concern are primarily zinc and sulfates, accompanied normally
by such elements as cadmium and lead, and small amounts of arsenic and
selenium.°
249
-------
Table 4-6. PRIMARY ZINC PROCESSING PLANTS IN THE UNITED STATES10'11'12
Company
AMAX, Inc.
ASARCO, Inc.
ASARCO, Inc.
St. Joe Minerals
Corporation
The Bunker Hill Co.
The National Zinc
Co.
The New Jersey Zinc
Co.
Location
Sauget, Illinois
Columbus, Ohio
Corpus Christi,
Texas
Monaca,
Pennsylvania
Wallace, Idaho
Bartlesville,
Oklahoma
Palmerton,
Pennsylvania
Type of operation
Electrolytic
Grate furnace
Electrolytic
Electric retort
Electrolytic
Horizontal retort
Vertical retort
Annual zinc
producing capacity
metric tons/yr
76,000
20,000
98,000
227,000
93,000
50,000
103,000
Date plant
built
1929
1967
1942
1930
1928
1907
1899
-------
Limitations on liquid effluent and atmospheric emissions from new
sources have been promulgated.8»10 The solid waste problem is currently
under investigation.
Bibliography
1. Commodity Data Summaries, 1976. U.S. Department of interior.
Bureau of Mines. Washington, D.C., 1976.
2. McMahon, A.D. et. al. Bureau of Mines Minerals Yearbook,
1973. U.S. Department of the Interior. U.S. Government
Printing Office. 1973.
3. 1975 E/MJ International Directory of Mining and Mineral Pro-
cessing Operations. Mining Informational Services of the
McGraw-Hill Mining Publications. New York. 1975.
4. Deane, G.L. et. al. Cadmium: Control Strategy Analysis. GCA
Corporation Report No. GCA-TR-75-36-G. U.S. Environmental
Protection Agency, Research Triangle Park, North Carolina.
April 1976. p. 157.
5. Smelting and Refining of Nonferrous Metals and Alloys. 1972
Census of Manufacturers. Publication MC72(2)-33C Bureau of
the Census. U.S. Department of Commerce. 1972.
6. Schlechter, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22, pp. 555-603.
7. Fejer, M.E., and D.H. Larson. Study of Industrial Uses of
Energy Relation to Environmental Effects. EPA-450/3-74-044.
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. July 1974. pp. VII-1 - XII-16.
8. U.S. Department of Interior, Bureau of Mines. Mineral Industry
Surveys. Zinc Industry in May 1976. August 2, 1976.
9. Jones, H.R. Pollution Control in the Nonferrous Metals In-
dustry 1972. Noyes Data Corporation, Park Ridge, New Jersey.
10. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA 450/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Pro-
tection Agency. Research Triangle Park, North Carolina.
October 1974 pp. 3-125 to 3-169, 5-18, 5-27, and 6-66 to
6-105.
251
-------
11. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. U.S. Environmental
Protection Agency. Washington, D.C. November 1974.
12. American Bureau of Metal Statistics. Nonferrous Metal Data
1975. New York, New York. 1976. pp. 65-87.
13. National Inventory of Sources and Emissions - Section V, Zinc.
W.E. Doris and Associates. Report No. APTD-1139. National
Technical Information Service, U.S. Department of Commerce.
Springfield, Virginia. May 1972. p. 77.
252
-------
INDUSTRY SEGMENT ANALYSIS
As with the copper and lead segments, the environmental impacts of
the zinc mining and smelting industries have received considerable
attention in recent years from both private and governmental organiza-
tions. This industry segment analysis examines each production opera-
tion, to define its industrial purpose and its potential and practice in
affectting the quality of the environment. Each process is examined as
follows:
1. Function
2. Input Materials
3. Operating Conditions
4. Utilities
5. Waste Streams
6. Control Technology
7. EPA Classification Code
8. References
The only processes included in this section are those that are now
operating in the United States or are under construction. Figure 4-1 is
a flowsheet that shows these processes, their interrelationships, and
their major waste streams.
253
-------
COHTUCT
SULFUIIIC
«cio PLJUT
tuTILITII!
UflOJKt
«X
2
k UTIIIIIIS
x;
• FIOTATIO*
U01IIVU
f \
STONE
Zltt OHM
nooucis
THICUKEI
Ct/UllOt
VII11TICS
voa CLECTitn.nE
Figure 4-1. Zinc industry flow sheet.
254
-------
IT-KOOUCT ucoton
t*
fUtlFTIIK
11
h zinc OUST
V
9
13
h UTILITIES
: COLLOID
SOOtUH CARtOIIATC OH
Wild HTOaUlDC
b UTILITIES
aoio
<£
CADHIUH
LEACH ,j
D
• UTILITIES
- HlSO,
- SOOIUH CHLORATE
- SPEKT ELECTROL
•CAUSTIC
L f«EC!PITATO*S
/ K )
5i*
y 5(
V
TE
V*3
CAOHIUH
FKECIFITATIOD
l<
h UTILITIES
E mcIMTATDU
•
r i
If
v
OT
CADKIIM
PU»IFICAT!0»
ABO CASTIKE 17
- LEAD SKELTER
IAGHOUSE OUST
- UTILITIES
- COAL
-------
PRIMARY ZINC PRODUCTION PROCESS NO. 1
Mining
1. Function - Mining involves work directed at the severance and
treatment of zinc ore deposits. Most zinc is mined underground using
open shrinkage, cut-and-fill, or square-set stoping methods. A few
mines, particularly in early stages of operation, use open pit methods,
which closely follow those used in copper mining. In underground mines,
walls and pillars are usually left behind to support the overlying rock
structure, unless the width of the ore body is such that it can be
unsupported and the entire ore body extracted.
The cycle of operations in mining consists of drilling, blasting,
and removing the broken rock. After removal from the mine, the ore is
loaded onto trucks for transportation to concentrating facilities.
Mining ranges from small operations handling several hundred metric tons
of ore per day to complexes capable of processing about.six thousand
metric tons per day. Some large, single-level, open-stage operations
employ "trackless" mining, with equipment mounted on crawler-type tread
or pneumatic-tired vehicles. This equipment facilitates movement and
increases speed of operations, thus increasing total output and reducing
costs; with such equipment it is sometimes feasible to mine ore con-
taining as little as 2 percent zinc.
2. Input Materials - Inputs to the mining process include the ore
deposits and the explosives used to remove them.
3. Operating Conditions - Ores are mined at atmospheric conditions.
Conditions depend upon type of mining, i.e., open pit or underground.
4. Utilities - Fuels and electricity are used for operating mining
equipment and transporting ores to concentrating facilities. One esti-,
mate for electrical usage'for mining ore in 1973 was 2.7 x IQll kg-cal.
With the 1973 mine production level of 434,000 metric tons,2 eiectrical
usage was about 620,000 kg-cal per metric ton of zinc moved. Unspeci-
fied amounts of water are pumped into mines for machinery and hydraulic
backfill operations.
5. Waste Streams - Mining operations give rise to residuals in the
form of air and water pollutants and solid wastes.
Zinc ores typically contain 3 to 11 percent zinc;3 thus 10 to 40
tons of ore must be mined for every ton of zinc produced. Most of this
ore is mined for other metals as well. Fugitive dust emissions are
similar to those of other mining operations, amounting to about 0.1 kg
per metric ton of zinc mined.4 Cadmium emissions also occur during the
mining of zinc ore. Emissions due to wind loss from tailings are esti-
256
-------
mated at 0.1 kg per metric ton of ore mined.5 Total emissions of
cadmium to the atmosphere were thus 240 metric tons for 19685 and 220
metric tons for 1973.6
The major solid wastes from mining operations result from removal
of rock to get to the ore. The discarded waste is of essentially the
same composition as raw ore, with lower metallic content. No quantita-
tive or qualitative estimate was found concerning these spoils.
Mine waste water results from infiltration of ground-water, water
pumped into the mine for machines, hydraulic backfill operations, and
infiltration of surface water. Quantities of effluent have no necessary
relation to the quantity of ore mined. The water required to maintain
operations may range from thousands of liters per day to 160 million
liters per day. this water contains such impurities as blasting agents,
fuel, oil, and hydraulic fluid. The dissolved solids found in the
wastewater are generally lead, zinc, and associated minerals.7
Conditions compatible with solubilization of certain metals--
particularly zinc--are associated with heavily fissured ore bodies.
Although the minerals being recovered are sulfides, a fissured ore body
allows oxidation of the ore, which increases the solubility of the
minerals.
6. Control Technology - The solid wastes from mining operations are
disposed of in a spoil pile or pond. These wastes can be used as mine
backfill.
Control of water from this process is described in the section on
sulfide ore mining and concentrating, covering all three industries. In
the zinc industry, mine water generated from natural drainage is reused
in mining and milling operations whenever possible. Discharge may
result because of an excess of precipitation, lack of a nearby milling
facility, or inability to reuse all of the mine waste water at a parti-
cular mill.
Low quantities of water are usually needed in the zinc flotation
process; mine-water effluent is used at many facilities as mill process
makeup water. The mine water may pass through the process first, or it
may be conveyed to a tailings pond from which it is conveyed for mill
flotation with recycled process water. The practice of combining mine
water with mill water can disturb the overall water balance unless the
mill circuit is capable of handling the water volumes generated without
a resulting discharge.
Acid mine water is neutralized by the addition of lime and lime-
stone. Acid mine water containing solubilized metals may be effectively
treated by combining the mine water with the mill tails in the mill
257
-------
tailings pond. The water may be further treated by lime-clarification
and aeration.
Fugitive dusts from drilling, conveying, and crushing can be re-
duced by wetting and control systems.8
7. EPA Souce Classification Code - None
8. References -
1. Dayton, S. The Quiet Revolution in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975.
2. Commodity Data Summaries 1976. U.S. Department of Interior.
Bureau of Mines. Washington, D.C. 1976.
3. McMahon, A.D. et al. In 1973 Bureau of Mines Minerals
Yearbook. U.S. Department of the Interior. U.S. Government
Printing Office. 1973.
4. National Inventory of Sources and Emissions - Section V, Zinc.
W.E. Doris and Associates. Report No. APTD - 1139. National
Technical Information Service, U.S. Department of Commerce.
Springfield, Virginia. May 1972.
5. National Inventory of Sources and Emissions - Cadmium, Nickel,
and Asbestos - 1968. Cadmium, Section I. W.E. Davis and
Associates. Report No. APTD - 1968. National Technical
Information Service. Springfield, Virginia. February 1970.
6. Deane, G.L. et al. Cadmium: Control Strategy Analysis. GCA
Corporation Report No. 6CA-TR-75-36-G. U.S. Environmental
Protection Agency, Research Triangle Park, North Carolina.
April 1976.
7. Development Document for Interim Final and Proposed Effluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source Cate-
gory Vol . 1 . EPA 440/1-75/061. Effluent Guidelines Division
Office of Water and Hazardous Materials, U.S. Environmental
Protection Agency. Washington, D.C. October 1975.
8. An Investigation of the Best Systems of Emission Reduction for
Quarrying and Plant Process Facilities in the Crushed and
Broken Stone Industry. United States Environmental Protection
Agency, Office of Air Quality Planning and Standards, Emission
Standards and Engineering Division, Research Triangle Park,
North Carolina 27711. Draft. April 1976.
258
-------
PRIMARY ZINC PRODUCTION PROCESS NO. 2
Concentrating
1. Function - Concentrating consists of separating the desirable
mineral constituents in the ore from the unwanted impurities by various
mechanical processes. The ore from mining must be concentrated because
mined sphalerite is seldom pure enough to be reduced directly for zinc
smelting.
The zinc-bearing ore is first crushed by standard jaw, gyratory,
and cone crushers to a size based on an economic balance between the
recoverable metal values and the cost of grinding.1 Size separation is
accomplished by vibrating or trommel screens and classifiers. Heavy-
medium cones, jigs, and tables separate the zinc minerals from a low
specific-gravity gangue. Classification and recycling between stages
reduces the material to a particle size appropriate for milling. The
final milled product is typically 60 percent smaller than 325 mesh.
After being transported to large bins for blending and storing, the
ore is pumped as an aqueous slurry to flotation cells, where it is con-
ditioned by additives and separated by froth flotation to recover the
zinc sulfide and sometimes lead or copper sulfides. In some cases, the
ore is reconcentrated by mechanical separation based on specific gravity
principles prior to froth flotation. Large mixers stir the solution,
and the zinc-bearing minerals separate and float to the surface where
they are skimmed off. Generally, zinc sulfide flotations are run at a
basic pH (usually 8.5 to 11), and the slurry is periodically adjusted
with hydrated lime, Ca(OH)2.' After flotation, the underflow (tailings
or gangue materials) is sent to a tailings pond for treatment.
Once separated, the metal concentrates are thickened in settling
tanks and the slurry is fed to vacuum drum filters, which reduce the
moisture content. At completion of the concentration process, the zinc
content is about 55 to 60 percent. Thermal drying in direct-fired
rotary dryers may further reduce the moisture content of the concentra-
tes, which are then transported to a storage site. Concentrate enters
the dryer with about 11 percent moisture and leaves with about 3 percent
moisture.2
In certain western ores, notably those from Idaho, the lead and
zinc are too finely divided for satisfactory separation even by flota-
tion. For such ores, final separation involves sulfuric acid leaching
at an electrolytic zinc plant. Some foreign producers use the Imperial
furnace to treat these ores, but is is not used at any U.S. zinc smel-
ters. Treatment of such ores was the primary reason for development of
electrolytic zinc recovery methods in North America.
259
-------
The quantity of zinc concentrate produced is about 10 to 15 percent
of the zinc ore by weight. A typical analysis would be 52 to 60 percent
zinc, 30 to 33 percent sulfur, 4 to 11 percent iron, and lesser quan-
tities of lead, cadmium, copper, and other elements.2 Table 4-7 lists
some of these elements of ore concentrate.
Additional details regarding concentrating equipment and procedures
are given in Primary Copper Production, Process No. 2. The concentrating
techniques and equipment are similar to those used for copper ores.
2. Input Materials - The quantity of ore required per ton of zinc
concentrate produced varies with the zinc content of the ore. A common
range is 5 to 10 tons of ore per ton of zinc concentrate.'
Hydrated lime is used to adjust the pH. Promoters or collectors
are added to the zinc sulfide to provide a coating that repels water and
encourages absorption of air. Frothers are added to produce a layer of
foam on the top of the flotation machine, and depressants are added to
stop unwanted minerals from floating. Table 4-8 lists commonly used
reagents.
Additives are an estimated 1.9 kg of lime, 0.4 kg of copper sul-
fate, 0.04 kq of Z-5 zanthate, and 0.02 kg of pine oil per metric ton of
concentrate.6
3. Operating Conditions - Flotation machines are operated at ambient
temperatures and pressures. Flotations are generally run at elevated pH
values of 8.5 to II.1
4. Utilities - Electrical energy is consumed during milling opera-
tions']One estimate for electrical usage in the milling of zinc ores in
1973 is 1.8 x 1011 kg-cal.7 At the 1973 mine production level of 434,000
metric tons,8 electrical usage was about 410,000 kg-cal/metric ton of
zinc produced.
The water requirement ranges from 330 to 1,100 m3/metric ton of ore
processed per day.' Feed water for the mills is usually taken from
available mine waters.
5. Waste Streams - It is estimated that 1 kg of particulate is emitted
per metric ton of ore processed during crushing and grinding operations.
The composition is that of the raw ore fed to the process. After water
is added to form an ore-water slurry, particulate emissions are negli-
gible. In plants that incorporate an ore drying operation following
concentration, dust emissions occur as the hot air passes over the
moving bed of concentrate. Operating factors affecting emissions are
the process feed rate and moisture content.
260
-------
Table 4-7. RANGE OF COMPOSITIONS OF ZINC CONCENTRATES
3,4,5
Constituent
Pb
Zn
Au
Ag
Cu
As
Sb
Fe
Insolubles
CaO
S
Bi
Cd
Percent
0.85-2.4
49.0-53.6
not iden.
not iden.
0.35
0.105-0.15
not iden.
5.5-13.0
3.4
not iden.
30.7-32.0
not iden.
0.24
261
-------
Table 4-8. TYPICAL FLOTATION REAGENTS USED FOR ZINC
CONCENTRATION UNITS1
Reagent
Purpose
Methyl isobutyl-carbinol
Propylene glycol methyl ether
Long-chain aliphatic alcohols
Pine oil
Potassium amyl xanthate
Sodium isopropol xanthate
Sodium ethyl xanthate
Dixanthogen
Isopropyl ethyl thionocarbonate
Sodium diethyl-dithiophosphate
Zinc sulfate
Sodium cyanide
Copper sulfate
Sodium dichromate
Sulfur dioxide
Starch
Lime
Frother
Frother
Frother
Frother
Collector
Collector
Collector
Collector
Collectors
Collectors
Zinc depressant
Zinc depressant
Zinc activant
Lead depressant
Lead depressant
Lead depressant
pH adjustment
262
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Liquid waste streams from zinc mills vary in volume from 1000 to
16,000 cubic meters per day. In terms of volume of ore processed,
liquid waste streams from milling operations range from 330 to 1,100
m3 per metric tonj Tailings slurry discharge is about 4 m3 per metric
ton of ore processed.
The raw wastewater from a lead/zinc flotation mill consists prin-
cipally of the water used in the flotation circuit, along with any
housecleaning water. The waste streams consist of the tailings streams
(usually the underflow of the zinc rougher flotation cell), the overflow
from the concentrate thickeners, and the filtrate from concentrate
dewatering.
The principal characteristics of the waste stream from mill opera-
tions are as follows:
(1) Solids loadings of 25 to 50 percent (tailings).
(2) Unseparated minerals associated with the tails.
(3) Fine particles of minerals, particularly if the thickener
overflow is not recirculated.
(4) Excess flotation reagents not associated with the mineral
concentrates.
(5) Any spills of reagents that occur in the mill.
It is very difficult analytically to detect the presence of excess
flotation reagents, particularly those that are organic. The surfactant
parameters may give some indication of the presence of organic reagents,
but provide no definitive information.
A typical quantity of solid wastes is 0.9 to 1 metric ton per metric
ton of ore milled. Based on a 25 to 50 percent solids loading in the
liquid waste stream, solid wastes from flotation could range from 80 to
550 m3 per metric ton.' The main component of the waste is dolomite.
Table 4-9 gives raw and treated waste characteristics of five
mills. This summary does not include information for a mill using total
recycle and one at which mill wastes are mixed with metal refining
wastes in the tailings pond. Feed water for the mills is usually drawn
from available mine waters; however, one mill uses water from a nearby
lake. These data illustrate the wide variations caused by ore miner-
alogy, grinding practices, and reagents.
263
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Table 4-9. RANGES OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE LOADS FOR FIVE SELECTED MILLS
1
to
(Pi
Parameter
PH3
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and grease
MBAS surfactants
P
Ammonia
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Cyanide
Sulfate
Chloride
Fluoride
Range of
^ concentration
in wastewater mg/1
lower limit
7.9
26
310
<2
670
71.4
11
0
0.18
0.042
<0.05
<0.0001
<0.1
0.12
<0.02
0.005
<0.02
<0.02
0.05
<0.01
295
21
0.13
upper limit
8.8
609
1,760
108
2,834
1,535
35
8
3.7
0.150
14
0.1
1.9
0.46
0.36
0.011
0.67
0.08
0.53
0.03
1,825
395
0.26
Range of raw waste load
Per unit ore milled,
kq/1000 metric tons
lower limit
410
460
7
940
6
6.35
5
0.236
0.108 .
0.064
<0.00013
<0.127
0.089
<0.026
0.008
<0.026
<0.026
0.064
<0.013
130
20
0.370
upper limit
1 ,600
4,700
285
8,500
4,800
130
21
13
0.876
26.4
0.0026
6.9
17.2
0.158
0.018
1.77
0.290
1.16
0.109
4,800
870
0.944
Per unit concentrate produced,
kg/ 1000 metric tons
lower limit
1,450
2,290
30
4,800
30
30
30
2.05
0.54
0.32
<0.00168
<0.900
0.62
<0.18
<0.18
<0.18
<0.45
0.012
0.091
1 ,260
210
203
upper 1 imf t
10,200
32,500
2,000
50,900
50,000
580
130
60.7
2.54
185
0.130
32.2
86.0
1.96
8.85
1.36
10.0
0.198
0.509
33,700
4,070
5.45
a Value In pH units.
-------
6. Control Technology - Particulate abatement equipment at the dryer
can capture dust and recirculate it into a storage bin for further use.
This can be accomplished with low-energy scrubbers and multicyclones.
Cyclones, operating on a dry principle, could remove many of the large
particles for reuse directly in the roaster without further processing.
Particulates caught in the scrubbers must be dried before reuse in
roasters or sintering plants.
Lime precipitation is often used for the removal of heavy metals
from wastewater. This treatment yields reductions for several heavy
metals including copper, zinc, iron, manganese, and cadmium.
Various techniques are employed to augment lime neutralization.
Among these are the secondary settling ponds, clarifier tanks, or the
addition of flocculating agents (such as polyelectrolytes) to enhance
removal of solids and sludge before discharge. Readjustment of the pH
after lime treatment can be accomplished either by addition of sulfuric
acid or by recarbonation. Sulfide precipitation may be necessary for
further removal of metals such as cadmium and mercury.
Water separated from the concentrates is often recycled in the
mill, but it may be pumped to the tailings pond, where primary separa-
tion of solids occurs. Usually, surface drainage from the area around
the mill is also collected and sent to the tailings pond for treatment,
as is process water from froth flotation.
Tailings pond water may be decanted after sufficient retention
time. One alternative to discharge, which reduces the output of efflu-
ent, is reuse of the water in other facilities as either make-up water
or process water. Usually, some treatment is required before reuse of
this decanted water. Treatments include secondary settling, phosphate
or lime addition, pH adjustment, flocculation, clarification, and
filtration.
The most frequently used control technology is use of a settling or
sedimentation pond system with a primary tailings pond and a secondary
settling or "polishing" pond, with pH adjustment prior to discharge.
Six lead and zinc ore processing facilities now use this technology.
Effluent concentrations are limited to the following average values (in
milligrams per liter): copper - 0.05, mercury - 0.001, lead - 0.02, and
zinc - 0.5.'
Control of wastewater is further discussed in the Section 6.0,
Water Management, which presents data typical for most of the zinc
industry.
265
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The tailings from mine concentrator operations may present a
serious water pollution problem if adequate precautions are not taken.
Coarse tailings may be removed with a cyclone separator and pumped to
the mine for backfill ingJ The most effective means of control is to
isolate disposal sites from surface flows and impoundment. Other tech-
n.iques include the fol 1 owing:1
(1) Construction of a clay or other type of liner beneath the
planned waste disposal area to prevent infiltration of surface
water (precipitation) or water contained in the waste into the
groundwater system.
(2) Compaction of waste material to reduce infiltration.
(3) Maintenance of uniformly sized refuse to enhance good compac-
tion (which may require additional crushing).
(4) Construction of a clay liner over the material to minimize
infiltration. This is usually followed by placement of top
soil and seeding to establish a vegetative cover for control
of erosion and runoff.
(5) Excavation of diversion ditches surrounding the refuse dis-
posal site to exclude surface runoff. These ditches can also
be used to collect seepage from refuse piles, with subsequent
treatment if necessary.
No data were found on the extent of application of these methods or
the relative costs.
7. EPA Source Classification Code - None
8. References -
1. Development Document for Interim Final and Proposed Affluent
Limitations Guidelines and New Source Performance Standards
for the Ore Mining and Dressing Industry. Point Source Cate-
gory Vol. 1. Publication EPA 440/1-75/061.. Effluent Guide-
lines Division Office of Water and Hazardous Materials, U.S.
Environmental Protection Agency. Washington, D.C. October
1975.
2. Field Surveillance and Enforcement Guide for Primary Metal-
lurgical Industries. EPA 450/3-73-002. Office of Air and
Water Programs, U.S. Environmental Protection Agency. Re-
search Triangle Park, North Carolina. December 1973. pp.
269-309.
266
-------
3. Water Pollution Control in the Primary Nonferrous - Metals
Industries -Vol. 1 - Copper, Zinc, and Lead Industries. EPA-
R2-73-247a. Office of Research and Development, U.S. Environ-
mental Protection Agency. Washington, D.C. September 1973.
4. Burgess, Robert P., Jr., and Donald H. Sargent. Technical and
Microeconomic Analysis of Arsenic and Its Compounds. EPA
560/5-76-016. Office of Toxic Substances, U.S. Environmental
Protection Agency. Washington, D.C. April 12, 1976.
5. Katari, V. et al. Trace Pollutant Emissions from the Pro-
cessing of Metallic Ores. EPA 650/2-74-115. U.S. Environ-
mental Protection Agency, Office of Research and Development.
Research Triangle Park, North Carolina. October 1974.
6. Mineral Processing Flowsheets. Denver Equipment Company.
Denver, Colorado. 1962.
7. Dayton, S. The Quiet Revoltuion in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975.
8. Commodity Data Summaries 1976. U.S. Department of the In-
terior, Bureau of Mines. Washington, D.C. 1976.
267
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PRIMARY ZINC PRODUCTION PROCESS NO. 3
Multiple-Hearth Roasting
1. Function - Multiple-hearth roasting is a high-temperture process
that removes sulfur from the concentrate and converts zinc to an impure
zinc oxide called calcine. Roasting may also be accomplished by suspen-
sion (Process No. 4) or fluidized-bed (Process No. 5) units. In a
multiple-hearth roaster, the concentrates drop successively from hearth
to hearth. As much as 20 percent of the cadmium present in the zinc
concentrate may be vaporized.! Any mercury present is volatilized and
goes with the gas stream. Multiple-hearth roasters are the oldest type
in use in the United States.
At present, only one domestic primary zinc smelter uses multiple-
hearth roasters, and this is in conjunction with fluidized-bed roasters.
The deleaded product, termed "partially desulfurized concentrate",
typically contains about 22 percent sulfur and is used as feed to the
fluidized-bed roasters.2
In roasting, if enough sulfur is originally present as sulfide, the
operation becomes autogenous. For pyrometallurgical refining, zinc
sulfate must be removed. The following reactions occur during roasting:
2ZnS + 302 + 2ZnO + S02
ZnO + S03 + ZnS04
2S02 + 02 -> 2S03
In pyrometallurgical reduction only the oxide state is desired, whereas
in electrolytic reduction, small amounts of the sulfate state are ac-
ceptable.3
Elimination of the remaining small percentages of sulfur requires a
long residence time. Hence a multiple-hearth-type roaster that eli-
minates all but 6 to 8 percent sulfur from up to 350 tons of copper
concentrates per day can roast only about 50 to 60 tons of zinc con-
centrate per day to about 2 percent sulfur.
Because concentrations of S02 produced during roasting are high
enough to allow recovery of sulfuric acid in an acid plant, this process
has become a normal part of zinc production.
Multiple-hearth roasters in use today are vertical cylindrical
types, modifications of the Nichols-Herreshoff design. They are versa-
tile and flexible, permitting different conditions on different hearths,
and are used effectively in deleading.
268
-------
The roaster consists.of a brick-lined cylindrical steel column 6 to
7.5 meters in diameter with 7 to 16 hearths. Modern furnaces are
larger, having 9 or more hearths. A motor-driven central shaft has two
rabble arms attached for each hearth, as well as cooling pipes. The
concentrates enter at the top of the roaster and are first dried in an
upper hearth. The central shaft rotates slowly, raking the concentrates
over the hearth with the rabble arms, gradually moving them to the
center and a drop hole to the next hearth. They move across this second
hearth to a slot near the outer edge, where they drop to the next
hearth. The concentrates continue down through the roaster in this
successive spiral fashion and are discharged at the bottom. Additional
fuel must be added to maintain combustion. Off-gases contain up to 6 to
7 percent sulfur, depending on the sulfur content of the feed.
The low production rates are a major disadvantage of multiple-
hearth roasting. However, since less dust is carried away in the gas
stream more volatile sulfides such as lead and cadmium are removed
preferentially. This is helpful when lead and cadmium are to be re-
covered from the flue dust, since there is less zinc dust contamina-
tion. 4
Total residual sulfur in the calcine produced in multiple-hearth
roasters is 2.4 percent; 0.5 to 1.0 percent is present as sulfide and
the balance as sulfate.3
Feed capacities for the multiple-hearth roasters average 180 metric
tons per day.
2. Input Materials - Zinc concentrate is the input material for
multiple-hearth roasters. Approximately 2.4 tons of pure ZnS is re-
quired to produce 1 ton of pure ZnO. In practice, the quantity of raw
materials required to produce one unit of calcine varies depending on
impurities in the concentrate, efficiency of the dust-collecting devices
on the roaster, and the type of roaster used.
Sodium or zinc chloride may be added to combine with cadmium dust
in the roaster and facilitate removal of cadmium as a by-product after
sintering. Specific quantities have not been reported.
3. Operating Conditions - Multiple-hearth roasters are unpressurized.
Average operating temperature is about 690°C;3 the lower hearths (sixth
through tenth from the top) are maintained at 950° to 980°C.2 Operating
time depends upon the type of roaster, composition of concentrate, and
amount of sulfur removal required.
4. Utilities - The reaction converting ZnS to ZnO is exothermic and is
self-sustaining after ignition. Gas, coal, or oil must be added ini-
tially to bring the charge up to reaction temperature. About 280,000
kg-cal/metric ton of ore are required for ignition. The primary fuel is
269
-------
natural gas. In some multiple-hearth furnaces, when concentrations of
less than 1.0 percent of sulfide sulfur are required, about 1.1 million
kg-calories are required per metric ton of feed.5. Some additional fuel
is added to the lower hearths to reduce the zinc sulfide content to as
low as 0.5 to 1.0 percent.
Cooling water and air are also used to cool the furnace shaft.
5. Waste Streams - In a zinc smelter the roasting process is typically
responsible for more than 90 percent of the potential SO? emissions; 93
to 97 percent of the sulfur in the feed is emitted as sulfur oxides.
Concentrations of S02 in the off-gas vary with the type of roaster
operation. The volume of off-gases ranges from 140 to 170 cubic meters
per minute^ for furnaces currently in use (approximately 180 metric tons
per day). Typical S02 concentrations range from 4.5 to 6.5 percent.J
Oxygen, nitrogen, carbon dioxide, and water vapor are other components
of the gas.
Particulate matter is also emitted. The amount and composition
vary depending on such operating conditions as air flow rate, particle
size distribution, and equipment configuration. Particulate emissions
consist of fumes and dusts composed of the zinc concentrate elements in
various combinations. One plant has reported the following compositions
of flue dusts from multiple-hearth, suspension, and fluid bed roasters:
zinc - 54.0 percent, lead - 1.4 percent, sulfur - 7.0 percent, cadmium -
0.41 percent, iron - 7.0 percent, copper - 0.40 percent, manganese -
0.21 percent, tin - 0.01 percent, and mercury 0.03 percent.2 Typical
particulate emissions in the off-gas range from 5 to 15 percent of the
feed.5 Composition of the distilled metal fumes, which constitute an
appreciable portion of the waste gas particulate carryover, depends
primarily on the concentrate composition and operating conditions. The
single zinc smelter using multiple-hearth roasters, which also uses
suspension and fluidized-bed roasters, recovers mercury eliminated
during roasting from the gas purification stream.2 At that smelter, the
roasted material is treated by flotation to separate a fraction of waste
rock prior to further processing. Solid and liquid wastes are produced;
details in the literature are fragmentary.
6. Control Technology - Gases leave the roaster at a temperature of
only 200 to 220°C.They are routed directly to "hot" Cottrells2 and
then to the acid plant.
7. EPA Source Classification Code - Roasting/multiple-hearth
3-03-030-02.
270
-------
8. References -
1. Development Document for Interim Final Efflunet Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. Effluent Guidelines Divi-
sion Office of Water and Hazardous Materials. U.S. Environ-
mental Protection Agency, Washington. November 1974.
2. Lund, R.E. et al. Josephtown Electrothermic Zinc Smelter of
St. Joe Minerals Corporation. AIME Symposium on Lead and
Zinc, Vol. II. 1970.
3. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA-450/2-74-002a.
Office of Air and Waste Management. U.S. Environmental Pro-
tection Agency. Research Triangle Park, North Carolina.
October 1974. pp. 3-125 to 3-169, 5-18, 5-27, 6-66 to 6-105.
4. Schlechter, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22, pp. 555-603.
5. Fejer, M.E., and D.H. Larson. Study of Industrial Uses of
Energy Relative to Environmental Effects. EPA-450/3-74-044.
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. July 1974. pp. XII-1 to XII-16.
6. Field Surveillance and Enforcement Guide for Primary Metal-
lurgical Industries. EPA-450/3-73-002. Office of Air and
Water Programs, U.S. Environmental Protection Agency. Re-
search Triangle Park, North Carolina. December 1973. pp.
269-309.
271
-------
PRIMARY ZINC PRODUCTION PROCESS NO. 4
Suspension Roasting
1. Function - Suspension, or flash, roasting is a process for rapid
removal of sulfur and conversion of zinc to calcine by allowing the
concentrates to fall through a heated oxidizing atmosphere or blowing
them into a combustion chamber. Alternate roasting processes are the
multiple-hearth (Process No. 3) and the fluidized-bed (Process No. 5).
Roasting in suspension promotes better heat transfer than multiple-
hearth roasting and thereby increases reaction rates for desulfuriza-
tion. The chemical reactions occurring in the two processes are the
same, and the S02 stream produced during conversion of the zinc sulfides
to calcine is strong enough for sulfuric acid production. Removal of
mercury and cadmium is also similar.
Suspension roasting is similar to the burning of pulverized coal,
in that finely ground concentrates are suspended in a stream of air and
sprayed into a hot combustion chamber where they undergo instantaneous
desulfurization. In practice, a stream of hot air passes through the
finely ground concentrate to temporarily suspend the particles. The
reaction usually proceeds without the addition of fuel unless the sul-
fide content of the ore is too low.
The roaster consists of a refractory-lined cylindrical steel shell
with a large combustion space in the top and two to four hearths in the
lower portion, similar to those of a multiple hearth furnace. In more
recent models of flash furnaces, the feed concentrate is introduced into
the lower one or two hearths to dry before final grinding in an auxil-
iary ball mill and introduction into the combustion chamber. Because
the feed must be carefully sized, additional grinding may be needed for
proper preparation.
About 40 percent of the roasted product settles out on a collecting
hearth at the bottom of the combustion chamber. This coarser rnaterial
is likely to contain the most sulfur, so it is further desulfurized by
being rabbled across this hearth and another hearth immediately below
before being discharged from the roaster. Particulate collected in
ducts and control devices can be fed to these hearths to achieve further
oxidation and sulfate decomposition or to obtain a more homogeneous pro-
duct.1
The remaining 60 percent of the product leaves the furnace with the
gas stream, passing first through a waste-heat boiler and then to
cyclones and an electrostatic precipitator, where it is recovered.
About 20 percent of the suspended dust drops out in the boiler; the
cyclones and precipitator remove about 99.5 percent of the remainderJ
272
-------
Total residual sulfur in the calcine is 2.6 percent, with 0.1 to
5.0 percent as sulfide and the balance as sulfate.2
Feed capacities of older suspension roasters are about 90 metric
tons per day, whereas newer, roasters can handle about 320 metric tons of
concentrate per dayJ
2. Input Materials - Zinc concentrate is the input material for sus-
pension roasters. Sodium or zinc chloride may be added to facilitate
later removal of cadmium. Specific quantities for these inputs are not
available.
3. Operating Conditions - Suspension roasters are unpressurized and
operate at an average temperature of 980°C.l Off gases exit the roaster
at about 1000°C.3 Operating times vary, depending upon the same factors
as in multiple-hearth roasting.
4. Utilities - Natural gas, oil, or coal is used to bring the roaster
feed to ignition, after which exothermic oxidation of the sulfur main-
tains operating temperatures. About 280,000 kg-cal/metric ton of ore
are required for ignition.4 Air is also added to the suspension roaster.
Water is supplied to the waste heat boiler system, and relatively small
amounts of electricity are required for fans and pumps.
5. Waste Streams - The S0£ concentration in the off gas from suspen-
sion roasting is higher than that in multiple- hearth processes, averag-
ing 10 to 13 percent.5 It also contains oxygen, nitrogen, carbon dio-
xide, and water vapor. This higher S02 content increases the efficiency
of sulfuric acid production. Emission of particulates depends on oper-
ating conditions, averaging about 6 percent of the feed. These emis-
sions consist of dust and metal fumes, depending on the concentrate
composition and operating conditions. The volume of off gases ranges
from 280 to 420 cubic meters per minute.5
There are no solid or process water wastes. Solids are recycled to
recover by-product metals. There is a boiler blowdown from the waste
heat boiler.
6. Control Technology - The S02 stream is concentrated enough to allow
sulfuric acid production, discussed in detail in Section 2 of this
report. The roaster off gases are cooled to about 400°C by heat trans-
fer with waste heat boilers. 3 The gas is then typically diluted with
air and humidified with water sprays before cleaning in an ESP. After
conditioning, gas volume is 475-700 cubic meters per minute, and the S02
concentration is 6 to 8 percent. 6
7. EPA Source Classification Code - None
'•• 273
-------
8. References -
1. Schlechten, A.M., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22, pp. 555-603.
2. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA450/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Pro-
tection Agency. Research Triangle Park, N.C. October 1974.
pp. 3-125 to 3-169, 5-18 to 5-27, 6-66 to 6-105.
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA440/1-75/032. Effluent Guidelines
Division Office of Water and Hazardous Materials. U.S. EPA,
Washington. November 1974.
4. Study of Industrial Uses of Energy Relative to Environmental
Effects. EPA-450/3-74-044. U.S. Environmental Protection
Agency. Research Triangle Park, N.C. July, 1974. pp. XII-1
to XII-16.
5. Field Surveillance and Enforcement Guide for Primary Metal-
lurgical Industries. EPA-450/3-73-002. Office of Air and
Water Programs, U.S. Environmental Protection Agency. Re-
search Triangle Park, N.C. December 1973. pp. 269-309.
6. Jones, H.R. Pollution Control in The Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
274
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PRIMARY ZINC PRODUCTION PROCESS NO. 5
Fluidized-Bed Roasting
1. Function - The fluidized-bed roaster is the newest method for
removing sulfur and converting zinc to calcine. Multiple-hearth (Pro-
cess No. 3) and suspension (Process No. 4) roasters are also available.
In this roaster, finely ground sulfide concentrates are suspended and
oxidized in a bed supported on an air column. As in the suspension
roaster, the reaction rates for desulfurization are more rapid than in
the older multiple-hearth processes. The chemistry of this process is
the same as that in other roasters. This process also produces enough
S02 for manufacture of sulfuric acid. Removal of mercury and cadmium is
similar to removal of other wastes.
The fluidized-bed roaster was originally designed for calcining
arsenopyrite gold ores; several North American zinc smelters have
adopted it in several forms for use in pyrometallurgical and electro-
lytic processes. Designs differ primarily in whether the roasters are
charged with a wet slurry or a dry charge. One variation is fluid
column roasting, which was developed in this country by the New Jersey
Zinc Company after being used abroad. In this process, the feed to the
roasters is pelletized. The calcined product is also pelletized,
eliminating the need for further agglomeration by sintering.
In the fluidized-bed process, no extraneous fuel is required after
ignition has been achieved. Operation of the system is continuous. The
feed enters the furnace and becomes fluidized, or suspended, in a bed
supported on an air column. Temperature control is achieved via water
injection, manually or automatically. Relatively low, uniform operating
temperatures appear to lessen the formation of ferrite. The tempera-
tures in the roaster are high enough to warrant the use of waste-heat
boilers to cool the off gases.
The sulfur content of the charge is reduced from about 32 percent
to 0.3 percent. The efficiency of the operation is maximum when 20 to
30 percent excess air is supplied over the stoichiometric requirements
for oxidation of both sulfur and metals of the charge.l
Dust carryover into the dust collecting system is somewhat less
than in flash roasting. In some cases it has amounted to 20 to 40
percent of the calcine, although in others it has risen as high as 77
percent. The amount varies with the feed rate (and consequently with
the air rate), and with the size analysis of the material being roasted.
The lower figures for dust carryover are found in the fluid column type
of roasters, which also operate at slightly higher temperatures.
275
-------
The S02 content of roaster gas is reported to be 7 to 12 percent.
If higher, it is diluted to about 7 percent before reaching the contact
acid plant. The theoretical maximum SOg concentration achievable is
14.6 percent when roasting 100 percent zinc sulfide concentrates to
completion, unless the air supply is enriched with oxygenj
Although the preferred method of regulating temperature within the
bed is by water injection with automatic thermocouple-operated control,
water injection and slurry feeding can be eliminated when it is desir-
able to minimize the water content of the gas.
The major advantage of the fluidization roaster is the processing
of higher tonnages per furnace per unit time, because of the increased
reaction rates for desulfurization. Fewer man-hours are required.
Also, like the suspension roaster, the fluidized-bed roaster can produce
a calcine with lower total sulfur content than the multiple-hearth pro-
cesses; the exact percentage elimination of sulfur during roasting is a
function of the initial sulfur content of the concentrate.
Total residual sulfur is 2.6 percent, with 0.1 percent as sulfide
and the balance as sulfate.2
Feed capacities can range up to 320 metric tons of concentrate per
day.2
2. Input Materials - Zinc concentrate is the primary input material
for fluidized-bed roasting. As with other processes, sodium or zinc
chloride may be added to combine with cadmium dust, facilitating the
later removal of cadmium as a by-product. The feed is finely ground, at
one plant to 90 percent minus 0.044 millimeter.
3. Operating Conditions - Fluidized-bed roasters operate under a
pressure slightly lower than atmospheric through as much of the system
as possible. Operating temperatures average 1000°C. The temperature in
fluid column roasters is 1050°C. Operating times are variable, depending
upon the same factors as in other roasting processes.
4- Utilities - Natural gas, oil, or coal is used to bring the roaster
feed up to reaction temperatures, after which exothermic oxidation of
the sulfur maintains the temperature and operation is continuous. About
280,000 kg-cal/metric ton of ore is required for ignition.4 Cooling
water is added, as well as low-pressure air (19 to 21 kg/cm2) which is
introduced into the windbox for combustion and fluidization of the mix.
5. Waste Streams - Typical S02 concentrations in the off gas from
fluidized-bed roasters range from 7 to 12 percent, although the higher
figure is more common. Fluid column roasters average 11 to 12 percent.
Exit gases also contain oxygen, nitrogen, carbon dioxide, and water
vapor.
276
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Particulates emitted depend on the feed and average 50 to 85 per-
cent for fluidized-bed and only 17 to 18 percent for fluid column.5 As
with the other processes, the amount and composition of these dusts and
metal fumes depend on the concentrate composition and operating condi-
tions. The volume of off gas produced ranges from 170 to 280 cubic
meters per minute.5 Temperatures are approximately 950°C. As with
other roasting processes, solids are recycled to recover by-product
metals and there are no solid or process-water wastes.
6. Control Technology - Emissions controls are the same as for the
flash roasting process. The S02 stream is used for sulfuric acid pro-
duction; a detailed discussion of this process is presented in Section 2
of this report. Since the S02 control systems normally require clean
gas streams, particulates are captured by cyclones and ESP's before the
gas stream enters the acid plant and thus present no air pollution
problems. Waste heat boilers cool the gases to 400°C.
In one operation, roasting 127 metric tons of dry concentrates per
day, 30 percent of the calcine left the roaster via the overflow pipe,
23 percent was deposited in the waste-heat boiler, 44 percent was cap-
tured by the cyclones, and 3 percent entered the hot Cottrell electro-
static precipitator with the flue gasesJ Flue dusts are not recircula-
ted to the reactor, being sufficiently low in sulfide sulfur. Roasters
with pelletized feed yield about 80 percent to the overflow and 20
percent carryover as dust.
7. EPA Source Classification Code - None
8. References -
1. Schlechten, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology. Inter-
science Division of John Wiley and Sons, Inc. New York, 1967.
Volume 22, pp. 555-603.
2. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA-450/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Pro-
tection Agency. Research Triangle Park, N.C. October 1974.
pp. 3-125 to 3-169, 5-18 to 5-27, 6-66 to 6-105.
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. Battelle Columbus Laboratories. EPA Con-
tract Environmental Protection Agency. Research Triangle
Park, N.C. December 1973. pp. 269-309.
277
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4. Study of Industrial Uses of Energy Relative to Environmental
Effects. EPA-450/3-74-044. U.S. Environmental Protection
Agency. Research Triangle Park, N.C. July 1974. pp. XII-1
to XII-16.
5. Field Surveillance and Enforcement Guide for Primary Metallur-
gical Industries. EPA-450/3-73-002. Office of Air and Water
Programs, U.S. Environmental Protection Agency. Research
Triangle Park, N.C. December 1973. pp. 269-309.
278
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PRIMARY ZINC PRODUCTION PROCESS NO. 6
Sintering
1. Function - Sintering has two purposes: first, to volatilize lead
and cadmium impurities and discharge them into the effluent stream where
they can be captured; and second, to agglomerate the charge into a hard,
permeable mass suitable for feed to a pyrometallurgical reduction
system.
Dwight-Lloyd-type sinter machines are typically used in the zinc
industry. These are downdraft units in which grated pallets are joined
to form a continuous conveyor system. The feed is normally a mixture of
calcine or concentrates, recycled ground sinter, and the required amount
of carbonaceous fuel, which is pelletized and sized to assure a uniform,
permeable bed for sintering before-it is fed to the machines and ignited.
Different smelting methods demand different properties of the sintered
feed. Purity of the final zinc product dictates the type of sinter
needed.
The feed is dumped on one end of a moving conveyor and is ignited
as it enters the natural-gas-fired ignition box. Combustion is sustain-
ed by supplying air to the pellets. Temperature control is achieved by
limiting the coke and coal content and the sulfur content of the sinter
mix. Once oxidation is started, it becomes self-sustaining. Air flow
regulation provides additional temperature control. A rotating scalper
shaves off the top layer of the sinter bed just before the discharge end
of the machine. This top layer is the sinter product, with composition
at one plant as shown in Table 4-10. Estimates are that 80 to 90 per-
cent of the cadmium and 70 to 80 percent of the lead are removed from
the sinter feed.2,3 The dust collected from this circuit is greatly
enriched in these impurities, and by recycling, levels are built up high
enough in the flue dust to permit economic recovery of the cadmium. The
lower portion of the bed, which was not removed by the scalper, is
discharged to a set of crushers. The coarser material may be separated
on a vibrating screen. Oversized particles are returned to the sinter
machine, while undersized material is incorporated with the feed mix.
Usually 5 to 10 percent of the total feed appears as dust in the gas
that is discharged.3 These dusts become the input material to the
cadmium recovery process.
Structural strength of the sinter must be considered, especially in
vertical reduction furnaces, since it must be able to support high
degrees of confining pressure from the overburdening charge. Mechanical
collapse is prevented by close chemical control of the nonvolatile
ingredients and addition of silica to increase hardness and strength of
the sinter mass. In horizontal-reduction-type retorts, a soft, friable
sinter is usually desirable. Structural strength is not needed because
of the lack of heavy overburdening pressures. The cadmium content of
the dust is usually 1 percent, allowing profitable cadmium recovery.
279
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to
00
o
Table 4-10. PRODUCT SINTER COMPOSITION
(percent)
1
Prime western
Intermediate
High-grade soft
High-grade hard
Zn
55.5
57.1
63.6
58.6
Pb
0.33
0.037
0.005
0.006
Fe
7.9
8.9
6.8
7.9
r Cd
0.017
0.015
0.012
0.006
Si02
9.4
8.8
5.2
8.9
S
0.15
0.15
0.36
0.10
Other dusts
5-10
5-10
5-10
5-10
-------
A process known as "sinter slicing" may be utilized in Dwight-Lloyd
machines where higher grades of zinc are sought or recovery of impuri-
ties is profitable, the process is based on the fact that sintering or
ignition is not homogenous throughout the charge, but migrates downward,
eventually resulting in concentration of the lead and cadmium sulfides
at the bottom of the sinter cake. The finished sinter is sliced off for
further refining in zinc-reduction retorts. The high-cadmium lower
section of the sinter cake passes on to the discharge end of the ma-
chine. After crushing it is returned to the charge.
The dust collected in a baghouse still contains incompletely
oxidized materials such as sub-micron sized particles of cadmium metal.
Temperatures are controlled by CO gas burners and tempering air. The
burned dust usually contains 10 to 20 percent cadmium, 12 to 15 percent
zinc, and 35 to 45 percent or more lead.4 The dust is sent to a lead
smelter for further cadmium concentration and then to an electrolytic
plant for recovery of pure cadmium.
In sintering preroasted charges, it is reported that substitution
of fluosolid calcines for those made by older types of roasters has
meant handling a more finely divided charge consisting of about 2.8
percent sulfur, of which about 2.4 percent is sulfate. These calcines,
considerably finer than hearth-roaster calcines, are reported to sinter
very rapidly when mixed with 4 to 5 percent fine coal. The product
averages approximately 0.3 percent sulfur and 0.05 percent cadmium.5
In one pyrometallurgical zinc plant, a briquetting step is added in
preparing the charge for reduction. The sinter is ground, then mixed
with pulverized coal, clay, moisture, and a binder. The mixture is
pressed into small briquettes (about 0.7 kg) which are fed into a step-
grade autogenous coking furnace. The briquettes attain a strong struc-
ture, which resists disintegration as well as keeping the reductant and
zinc oxide in close contact. Heat is generated by burning the volatile
constituents of the charge produced inside the furnace.6
2. Input Materials - Specific quantities of input materials to a
sinter machine are dependent on properties of the concentrate and the
type of reduction (i.e. retort or electric). The sinter mix typically
contains calcine, recycled sinter, coke or oil, and sand or other inert
ingredients. One plant has reported using 22 kg of sand and 80 kg of
breeze coke (not including carbon in furnace residue and bag filter
dust) per metric ton of sinter produced.6 Moisture is added when the
constituents are mixed, although where available, zinc sulfate solutions
from in-plant leaching operations are used to moisten the feed for
pelletizing, since this conserves water and enhances zinc recovery.
281
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In typical zinc sintering operations, charging capacities range
from 220 to 550 metric tons per day and from 35 to 80 percent of the new
feed is recycled.7
3. Operating Conditions - Typical operating parameters for zinc
sintering are 1040°C temperature and atmospheric pressure. The wind box
fan operates at a high negative pressure required to pull combustion air
through the bed.
4. Utilities - The sinter mix usually contains 4 to 5 percent coke or
coal to supply enough heat for sintering.3 Natural gas is used to
ignite the mix. Electrical energy is required for operating the sinter
machine. Fuel consumption has been estimated to be about 280,000 kg-
cal/metric ton of concentrate processed, although one plant reported
use of 51,000 kg-cal/metric ton of sinter produced. Water (or zinc
sulfate solution, if available) is added to facilitate pelletizing. Air
is supplied to maintain combustion.6
5. Waste Streams - The sintering operation is a source of air pollu-
tants in the form of particulates and SO?. The SO;? concentration in the
off-gas from the sinter machine is very Tow, 0.1 to 2.4 percent by
volume, which represents only 1 to 5 percent of the sulfur originally
present in the feed.3 When zinc calcines are sintered, the sulfur
emissions from a sinter machine are primarily determined by the sulfur
content of the input calcine, although some emissions result from the
zinc sulfate liquor added to the sinter mix. Typically, the resulting
weak effluent stream contains approximately 1000 ppm S02*, the concen-
tration, however, can range from 400 to 3000 ppm S02, depending upon the
total sulfur content of the feed stock.7
Gas rates vary widely. With a calcined feed, exhaust gas rates
vary from 42 to 72 cubic meters per square meter of grate. With a
concentrate feed the gas rate is 5.4 to 6.5 cubic meters per square
meter.9 Temperature of the combined exit sinter gases varies from 160°C
to 380°C in different plants.9 In addition to the sulfur oxides, the
gases contain air, water vapor, carbon dioxide, and traces of other
gases.
All of the particulate matter is less than 10 microns. Solids
loadings range from 9 to TOO g/cu m.9 The emissions consist of dusts
and metallic fumes of composition similar to that of the calcine.
Typical chemical composition of the particulates is 5 to 25 percent
zinc, 30 to 55 percent lead, 2 to 15 percent cadmium, and 8 to 13 per-
cent sulfur.9 Other hazardous constituents include copper, arsenic,
antimony, bismuth, selenium, tellurium, and tin. The cadmium and lead
contents are especially high because about 90 percent of the cadmium and
70 to 80 percent of the lead is eliminated in the sinter machine. The
fumes condense and are collected with the dust.
282
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The sintgr process entails no sources of process liquid or solid
wastes. The water that is added to the mix is vaporized and emitted
through the stack, while all solids left unsintered are recycled.
6. Control Technology - Currently no control methods are applied to
the weak S02 stream from the sintering operation. The best available
control technology is application of chemical scrubbing techniques.
These are described in Process No. 6, Primary Copper Production, which
discusses scrubbers used for control of the S02 from reverberatory
smelters and similar processes. The exit gases may be cooled by air
dilution and water sprays in preparation for gas cleaning.
The ideal control for S02 emissions from sintering would be to
eliminate as much sulfur as technically possibly during roasting. Based
on the capability of fluid bed and suspension roasters, a calcine aver-
aging 1.5 percent total sulfur could be produced, rather than the cur-
rent typical calcine with approximately 3 percent total sulfur. Roast-
ing operations that reduce the residual sulfur content in the calcine
produce a corresponding decrease in sulfur emissions to the atmosphere
by sintering. Since additional coke or coal is then needed to accom-
plish sintering, the elimination of sulfur in the calcine would increase
the cost.
Various combinations of settling flues, ESP's, and baghouses are
used on the sintering machines. Efficiencies range from 94 to 99
percent.
The sinter crushing and screening operations have enormous poten-
tial for particulate emissions. These operations are hooded and ducted
to control devices. The exhaust gas is further cooled by dilution air
and water sprays to condition it for cleaning in dust collectors.
7. EPA Source Classification Code - 3-03-030-03
8. References -
1. Lund, R.E., et al. Josephtown Electrothermic Zinc Smelter of
St. Joe Minerals Corp. AIME Symposium on Lead and Zinc, Vol.
II. 1970.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. Effluent Guidelines
Division Office of Water and Hazardous Materials. U.S. EPA,
Washington. November 1974.
283
-------
3. Field Surveillance and Enforcement Guide for Primary Metal-
lurgical Industries. EPA-450/3-73-002. Office of Air and
Water Programs, U.S. Environmental Protection Agency. Re-
search Triangle Park, N.C. December 1973. pp. 269-309.
4. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer.
Encyclopedia of Chemical Technology. Interscience Division of
John Wiley and Sons, Inc. New York, 1967. Volume 3, pp. 884-
899.
5. Schlechten, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22, pp. 555-603.
6. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. Battelle Columbus Laboratories. EPA Con-
tract No. 68-01-1518. Draft Data.
7. McMahon, A.D., et al. The U.S. Zinc Industry: A Historical
Perspective. Bureau of Mines Information Circular 8629. U.S.
Department of Interior. 1974.
8. Fejer, M.E. and D.H. Larson. Study of Industrial Uses of
Energy Relative to Environmental Effects. EPA-450/3-74-044.
U.S. Environmental Protection Agency. Research Triangle Park,
N.C. July 1974. pp. XII-1 to XII-16.
9. Jones, H.R. Pollution Control in The Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
284
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PRIMARY ZINC PRODUCTION PROCESS NO. 7
Horizontal Retorting
1. Function - The horizontal, or Belgian, retort is a batch reduction/
volatilization process in which zinc oxide is reduced to zinc at ele-
vated temperatures, volatilized, and condensed to metallic zinc. Only
one primary zinc smelter uses horizontal retorts, and it is currently
being replaced with an electrolytic plant. The changeover is expected
to be completed in 1977. This process is discussed in some detail,
however, since many of the basic principles are involved in other
retorting systems (Process Nos. 8 and 9).
Because of the relatively low boiling point of zinc (906°C), re-
duction and purification of zinc-bearing minerals can be accomplished to
a greater extent than with most minerals. Immediate separation from
nonvolatile impurities is possible. In fact, if the material is treated
pyrometallurgically, there is virtually no alternative. Even at 857°C
(the lowest temperature at which the oxide can be continuously reduced),
the vapor pressure of zinc is high enough to cause it to vaporize
immediately upon reduction. Balanced against the easy separation from
nonvolatiles, however, are the difficulties of condensing the vapor and
the high volatility of several of the most common impurities, especially
cadmium and lead.
The substance most responsible for direct reduction of zinc com-
mercially is carbon monoxide. The zinc-reduction cycle consists of the
following reactions:
ZnO + CO -> Zn(vapor) + C02 (actual reduction step)
C02 + C •*• 2 CO (regeneration of CO)
Both reactions are reversible, and since the second is the slower of the
two at temperatures below about 1100°C, it controls the rate of reduc-
tion in most commercial situations. Above 1100°C the rates of diffusion
and heat transfer predominate as rate-controlling factors. Both of the
above reactions are highly endothermic.l The following additional
reactions occur during retorting:
Fe203 + 3C •*• 2Fe + SCO
Fe203 + C * 2FeO + CO
ZnS + Fe -»• Zn + FeS
Zn + C02 •*• ZnO + CO (blue powder formation)
285
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+ 2C ->• 2Zn + 2CO
FeO + SiO* •*• xFeO-Si02 (slag formation).
Carbon for reduction is normally provided by coal or coke, the
choice determined by the need for structural strength within the smelt-
ing column or, if zinc oxide is the product, by the danger of contamina-
ting the product with soot. Carbon in excess of stoichiometric reduc-
tion requirements is normally provided to furnish extra reaction surface
to compensate for the slowness of reduction of carbon dioxide by carbon.
The horizontal furnace consists of a series of externally fired
tubular refractory receptacles with a ratio of length to diameter
between 4 and 7.5. This shape is necessary for prolonged exposure of
the charge and reduction fuel at elevated temperatures. Because of the
high ratio of surface area to volume, gaseous combustion products are
passed around the retortsJ
The capacity of these retorts is about 18 to 27 kilograms of metal
per day. A typical retort is about 0.20 meter in internal diameter and
1.5 meters long, although the length of each is dictated by the hot
strength of the refractory materials. The tubes are made of approxi-
mately equal mixtures of grog and clay. Silica or high-refractory
materials such as silicon carbide are sometimes added. A 2 percent
increase in zinc recovery is claimed for the silicon carbide retort
because of its higher heat conductivity than clay bodies.2 Other ad-
vantages are longer life and the ability to handle a larger charge
because of the greater width of the retort.
Heat economy of the horizontal, batch-type retort is much less than
that of a continuous vertical retort. The necessity of reducing the
firing rate toward the end of the run is a major factor in the added
fuel consumption, since more energy is required to heat up than to
maintain temperature while producing continuously. Reasons for lowering
temperatures of the retort are (1) to prevent damage where the charge is
insufficient for absorbing the heat from the endothermic reaction, and
(2) to allow maneuvers for draining or removing residue from the re-
torts, replacing broken retorts, recharging retorts, etc.
A large number of these retorts are placed horizontally within an
enclosure with the retort faces exposed to the atmosphere. To conserve
energy, the furnaces are positioned back-to-back and are arranged in
banks stacked side by side. The refractories are fired externally by
passing combustion products through the areas between the retorts.
'&
Prior to a new distillation cycle, the residue is removed and
replaced with a mixture of roasted concentrate or sinter, .about 60
percent of its weight being semianthracite coal or coke. The apparent
density of the mixture is about 900 kilograms per cubic meter.1 An-
286
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thracite ordinarily is not used alone, partly because of its high cost
and partly because it is not quite as effective a reducing agent as
slightly softer grades of coal. Grades such as semianthracite tend to
form soot during pyrolysis, leaving the solid portion with a rather
porous structure. These materials exhibit a much higher reaction rate
in promoting the re-reduction of C02 to CO (the rate-controlling reaction
in the zinc-reduction cycle), thus speeding the reduction of zinc and
minimizing the oxidation of zinc vapor already formed. On the other
hand, coals so high in volatile hydrocarbons that they form a large
fraction of the retort atmosphere for an appreciable part of the reduc-
tion period cause dilution of the zinc vapor and lower condensation
efficiency.
The zinc vapors are passed into condensing chambers located over
the ends of the retorts and open at both ends. The mouths of the con-
densers are stuffed with charge or other high-zinc material, almost
completely sealing the outlet. Gas is allowed to escape through one
small hole in the stuffing after heavy firing is begun. About 6 hours
after starting, the first zinc begins to come off the charge and the
condensers must perform their functions. As the zinc condenses, it is
collected in a zinc well which is a depression on the bottom of the
condenser. Most but not all of the zinc vapor is condensed.
The reaction between zinc vapor and carbon dioxide (Zn + C02 •*• ZnO
+ CO) is continous in the condensers. Some zinc oxide is formed by
direct oxidation of zinc with external air. This reaction from left to
right is the more stable direction below 1100°C. The zinc oxide coats
the condensed fine globules of zinc metal, creating an undesirable by-
product called "blue powder". Rapid cooling of the zinc forms less blue
powder. Recovery of zinc from the sintered charge is about 95 percent,!
including the blue powder charge.
A temperature low enough to prevent the escape of any uncondensed
zinc will result in excess formation of blue powder; alternatively,
however, a temperature high enough to minimize blue powder formation
will cause appreciable loss of vapor. For this reason sheet metal tubes
called "prolongs" are attached to the ends of the condensers. These
were developed to collect escaping zinc vapor for recycling while
minimizing the percentage of blue powder in the condenser zinc. Pro-
longs increase zinc recovery by about 2 or 3 percent, but in the domes-
tic industry the cost of using them is considered to be greater than the
value of the extra recovery.
Gases leaving the retort section contain zinc vapor, carbon mon-
oxide, and particulates. The gases pass out the top of the column to a
condenser, where the zinc is condensed from the gas stream. Next, the
gas is scrubbed and recycled to the combustion zone.
287
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A full cycle for reduction is usually 24 or 48 hours, with three or
four draws of metal each day. Further refining is necessary to remove
unwanted impurities that were not eliminated during roasting and reduc-
tion. The two methods employed are redistillation and liquation.
Fractional distillation can produce zinc of higher purity than
liquation. The condensed vapors from this process are typically high in
cadmium content and can be used for cadmium recovery. The boiling point
of lead is about 1620°C and of cadmium, 767°C. The boiling point of
zinc lies between the two. Since .none of these elements forms stable
compounds with each other that vaporize without dissociation, they are
amenable to separation by fractional distillation. The most common
method entails dual fractionating columns of silicon carbide, heated
externally. Cadmium and zinc are largely volatilized from the first
column, leaving lead, iron, and other high-boiling-point constituents
that can be removed from the base. The condensate passes to a second,
or cadmium, column, where cadmium and low-boiling-point impurities are
removed by reflex condensation. Purified zinc flows out the bottom.of
the column. The cadmium is collected as cadmium-zinc metal or as cad-
mium dust depending on how the process is operated.
The purity of the zinc produced by fractional distillation averages
from 98.0 to 98.5 percent. Depending upon the metal analysis and the
specifications to be met, metal from each draw may be cast separately,
or metal from several draws may be combined in a holding furnace and
cast together. The melt is retained just above its melting point, at a
temperature at which impurities are least soluble, until a high-lead
layer and a high-iron layer are formed at the bottom. These are removed
from below.
Liquation removes lead and iron using the principle of mutual
solubilities to precipitate these impurities. By cooling the liquid
metal to just above the melting point of zinc, some iron and lead will
precipitate out forming a layer containing these impurities on the
bottom while any oxides will float on top. The purified zinc Jrawn from
the middle typically contains 1.2 percent lead and 0.025 percent iron.
The partially purified zinc can then be cast as slab zinc into prime
western grade or higher, depending on purity.
The external firing and batch nature of the horizontal retort are
only two of the unattractive features which are adding to the demise of
the system. Low productivity and excessive operating costs are addi-
tional problems. Since each retort produces only about 45 kilograms of
zinc each 48 hours, a furnace must incorporate many retorts, requiring
large and expensive work crews, to produce significant quantities of
zinc. Horizontal retorts are also not amenable to particulate control.
288
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2. Input-Materials - The feed input consists of sinter and coal or
coke as well as recycled material in the blue powder charge. Quantities
of coal or coke range from 0.5 to 0.8 ton per ton of slab zinc output.2
The horizontal retort places few restrictions on the input sinter
properties. Because it requires only a soft, friable, and poor-ap-
pearing sinter, sinter production need not be tied to a particular
roasting or sintering process.
3. Operating_C6riditibns - Horizontal reduction furnaces are not
pressurized. Temperatures outside the retorts reach 1300 to 1400°C;
while temperatures inside are about 1100 to 1200°C.l
For further purification steps, exact operating temperatures vary.
Separation is based on differences in boiling points. By-products of
zinc (boiling point 907°C) such as lead, copper, aluminum, tin, and iron
have significantly higher boiler points, and cadmium has a much lower
boiling point (767°C).3
Liquation typically requires 16 to 24 hours of settling time at
temperatures just above the melting point (approximately 420°C) for
proper separation.4
4. Utilities - Horizontal retorts are fired with natural gas. The
amount of heat theoretically required to produce 1 kilogram of zinc
metal by reducing ZnO with carbon to zinc and carbon monoxide is 1390
kg-cal."
Distillation of zinc requires from 2,500 to 3,500 kg-cal/kg of zinc
produced.4 Typically, a horizontal retort operates at fuel efficiencies
of 5 percent or less because most of the fuel energy is lost up the
flues. Energy consumption for this process is about 16.7 million kg-
cal /metric ton of zinc produced. Typical energy requirements for the
subsequent refining steps are approximately 834,000 kg-cal/metric ton of
zinc produced.5
5. Waste Streams - Emissions from retorting are minor compared with
those from other operations such as roasting and sintering. Only about
0.2 to 0.3 percent of the total sulfur in the raw concentrate is re-
leased during retorting operations. Particulate emissions consist
primarily of metal and metal oxide fumes.
Particulate emissions for horizontal retorts (blue powder) average
about 8 kilograms per metric ton of zinc produced. The principal
component of these emissions is zinc oxide, a fine, highly-visible white
particulate. Cadmium, copper, chromium, lead and iron are also released.
Solids loading ranges from 0.9 to 3.0 grams per m3; the chemical composi-
tion of the particulates is 50 to 70 percent zinc and 0 to 3 percent
289
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lead. Particle size has been reported as ranging from micron to sub-
micron diameters. Particles in fumes from one horizontal retort plant
were reported to be 34 percent less than 2.5 microns, 35 percent between
2.5 and 5.0 microns, and 31 percent larger that 5 microns.6 The flow
rate of the condenser waste gas including extraneous air is 95,000 to
140,000 m3 per minute. Flow rate of the flue gas is 10,700 to 16,250 m3
per metric ton of zinc product, with 12 to 17 percent carbon dioxide.6
The important process variables are the zinc and coke content of
the feed and the air flow rates. Proper air flow is important, since
too much air might result in nearly complete combustion of the coal.
The most important process parameter is temperature, which is regulated
by fuel rate adjustment.
Reduction operations are also sources of solid residue amounting to
around 1050 kilograms per metric ton of zinc produced. The residue may
frequently amount to as much as 20 to 30 percent of the charge (assaying
13 to 15 percent zinc and 45 percent carbon)J The residues contain
lead, copper, silver, gold, nickel, germanium, gallium, arsenic, anti-
mony, cadmium, zinc, indium, silicon, iron, calcium, aluminum, magne-
sium, and manganese. The residue can be sent to lead smelters to re-
cover the metallic values. The germanium and gallium contents origi-
nally found in the blends are in general concentrated in the residues
from the retorts. These residues can be treated by dissolution in
caustic soda, followed by treatment by various methods, most important
probably being the extraction of gallium chloride with an organic
solvent. There are many variations on these methods to circumvent
various impurity problems. This processing is not done at any of the
primary zinc smelters. Two companies, one in Arkansas and one in
Oklahoma, account for .the total domestic production of gallium, using
residues from zinc and aluminum production. One refinery in Oklahoma
produced all the primary domestic germanium from zinc smelter residues
in 1975.7
In liquation or redistillation steps, the vaporizing and condensing
systems are closed circuits, so no zinc vapors escape. Because all
solid residues are reprocessed to recover the remaining zinc and by-
product metals, there is no solid waste. Waste gases are produced by
the combustion gases used in the final purification steps. Typically
0.35 to 0.5 metric ton of coal or the equivalent is required per metric
ton of high-grade zinc produced, which is approximately 3,000 to 4,500
Mm3 per metric ton of gaseous zinc.3 The pollutants in these gases
depend upon the fuel used.
The water used to wash the gas stream to permit use of carbon
monoxide as a fuel contains zinc and metal oxides, possibly hydrocar-
bons, various particulates (as suspended solids), and the corresponding
products of hydrolysis.
290
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6. Control Technology - The principal control method for reducing
participate emissions at the present time is use of a wet scrubber.
Zinc oxide is recovered as "blue powder" in slurry form and is settled
in concrete pits. Eventually the zinc oxide is dredged and recycled
through the process. No data were found on control efficiency or. costs.
The cleaned gases rich in carbon monoxide are recycled and used as
auxiliary fuel for the retorts. Seldom are they directly vented.
Cyclone removal of large particles followed by fabric filtration after
cooling is another control possibility.
Stringent requirements for reducing particulate emissions could be
prohibitive for horizontal retorting. Systems for collecting the many
small gas streams for cleaning, or for cleaning each individual stream,
have not yet been devised.
Residues from the reduction operations are disposed of mostly on
open slag dumps. Since the retort residue can amount to 20 to 30
percent of the charge, it is frequently not included in the blue powder
charge but is treated on magnetic separators. One plant processes the
residue to recover metallic values. It is thought that these slags do
not leach, but if this is not true further control techniques are
necessary. The best control technology consists of sealing the soil at
the slag dump and routing runoff to a waste treatment lagoon.
7. EPA Source Classification Code - 3-03-030-04
8. References -
1. Schlechten, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22. pp. 555-603.
2. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II, Preimary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft-
April 1975). pp. 81-127, 239-250.
3. Restricting Emission of Dust and Sulfur Dioxide in Zinc
Smelters. Association of the Metal Smelting and Refining
Trade and Committee on Zinc of the German Ore Smelting and
Mining Society VDI No. 2284. September 1961.
4. McMahon, A.D., et al. The U.S. Zinc Industry: A Historical
Perspective. Bureau of Mines Information Circular 8629. U.S.
Department of Interior. 1974.
291
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5. Study of Industrial Uses of Energy Relative to Environmental
Effects. EPA-450/3-74-044. U.S. Environmental Protection
Agency. Research Triangle Park, N.C. July 1974. pp. XII-1
to XII - 16.
6. Jones, H.R. Pollution Control in The Nonferrous Metals
Industry. Noyes Data Corporation, Park Ridge, New Jersey.
1972.
7. Commodity Data Summaries 1976. U.S. Department of Interior,
Bureau of Mines. Washington, D.C., 1976.
292
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PRIMARY ZINC PRODUCTION PROCESS NO. 8
Vertical Retorting
1. Function - The vertical retort process is a continuous reduction/
volatilization method for producing high-purity zinc from zinc oxide by
reduction with carbon at elevated temperatures in vertical silicon-
carbide retorts. Alternate retorting processes are the horizontal
(Process No. 7) and the electric (Process No. 9). The vertical retort
overcomes the obvious disadvantages of batchwise production, although
like the horizontal retort it is externally fired. Operation can be
automated to reduce costs. Since vertical retorting eliminates many of
the major faults of the horizontal system, this system should remain in
use for the foreseeable future.
Carbon monoxide is used for direct reduction of zinc in vertical
retorts. The reactions are the same as for horizontal retorts, and also
similarly, the carbon for reduction is provided by coal or coke.
The charge to a vertical retort cannot be a granular, loose, and
soft mixture as in the horizontal retort, but must be in the form of a
hard sinter or briquette. The reduction fuel is mixed with the ore and
a temporary binder. The briquettes are then coked in an autogeneous
coking furnace in an operation that also drives off volatiles which
serve as fuel. Each unit has a capacity of 90 to 108 metric tons of
briquettes averaging about 40 percent zincJ
Although briquetting is an expensive operation, the ability to use
briquetted sinter feed could be turned into an advantage since it per-
mits the use of a soft sinter produced either by coke-sintering or
roast-sintering. Such a process is being used in England by the Impe-
rial Smelting Corporation, Limited, but it has not been adopted by
domestic producers.
Vertical retorts are large, refractory-lined vessels with external
gas chambers. The furnaces consist of three major sections - the charge
column, the reflux section, and the combustion-heating chambers. The
retort is rectangular in general cross-section, about 0.3 meter wide,
1.8 to 2.4 meters long, and 10.6 meters high, giving a capacity of about
7.25 metric tons of zinc per retort per day. Walls are of silicon
carbide to facilitate heat transfer and minimize penetration by zinc
vapor. Joints in the end walls are packed with silicon carbide and
graphite to permit differential expansion upon heating. Production
rates are reported at an average 195 to 215 kilograms of metallic zinc
per day for each square meter of long wall surface when heated to
1300°C. Reported life of retorts is about 3 years.'
293
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Without intermediate cooling, coked briquettes are fed to the
charging extension at the top of each vertical retort. The charge is
heated by gas in chambers surrounding the retort sidewalIs. Gases from
the combustion chambers are used to preheat incoming air for combustion
by means of recuperators. The briquettes maintain their shape through-
out the reduction operation. The furnace has a vertical retort shaft
which allows the charge, with the aid of gravity, to pass downward
through the combustion or heating zone of the column; heat produced in
the combustion chamber is transferred through the refractory walls of
the column to the charge. As the charge moves down through the retort,
the zinc oxide decomposes to form zinc vapors and carbon monoxide.
Approximately 95 percent of the zinc vapor leaving the retort is con-
densed to liquid zinc.2 The residue, containing approximately 10 per-
cent zinc,3 is removed at the bottom through an automatically controlled
roll discharge mechanism into a quenching compartment, from which it is
removed for further treatment.
During the passage of the briquettes through the retort, sufficient
air or exhaust combustion products are introduced at the base of the
charge to ensure that no zinc vapor moves concurrently with the charge
and eventually condenses on spent residue. The gaseous-reaction pro-
ducts formed in the retort, which rise up through the charge column, are
approximately 40 percent zinc vapor, 45 percent carbon monoxide, 8
percent hydrogen, 7 percent nitrogen, and some carbon dioxide.4 These
gases exit near the top of the charge column, pass first through a zinc
condenser, and then to a venturi scrubber. By means of a splash system,
whereby a mechanically driven device fills the condenser chamber with a
rain of zinc droplets that fall back into a batch of molten zinc, the
zinc vapor from the retorts is condensed and collected with excellent
efficiency. Over 95 percent of the zinc vapor leaving the retort is
condensed to liquid zinc.5 The carbon monoxide is recycled to the
combustion zone.
A back oxidation reaction in the condensation process is respon-
sible for the production of 3 to 5 percent partially oxidized .-inc, or
blue powder, which floats on top of the zinc bath and is periodically
skimmed off. Scrubbers may be used to remove the entrained blue powder
from the flue gases, and the cleaned gas may be either used as supple-
mentary fuel or flared. The zinc produced may be further refined.
About one-half of the zinc produced in vertical retorts is refined to
99.99 percent purity by means of continuous fractioned distillation.
This purification process is the same as that described for horizontal
retorts, and the condensed vapors produced are similarly high in re-
coverable cadmium.
294
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2. Input Materials - The feed input is the same as for a horizontal
retort, consisting of sinter, coal or coke, and recycled material in the
blue powder charge. In addition, since the charge must be either a hard
sinter or briquette, a temporary binder such as sulfite waste liquor,
tar, or pitch is added. A typical composition is 60 percent sinter, 25
percent bituminous coal, 5 percent anthracite fines, 10 percent plastic
refractory clay, and 1 percent sulfite liquor.4 With zinc recovery
typically 90 percent, about 2.8 metric tons of feed is required per
metric ton of zinc collected in the condenser for a feed containing 40
percent by weight zinc. Quantities of coal or coke required range from
0.5 to 0.8 metric ton per metric ton of slab zinc produced.3
3. Operating Conditions - Vertical reduction furnaces generally oper-
ate from 1300 to 1400°CJ Internal pressures are slightly below ambi-
ent. Exact operating temperatures for fractional distillation or other
purification steps vary, since separation is based on differences in
boiling points.
4. Utilities - Heating is done in vertical retort furnaces by com-
bustion of gas surrounding the retort side walls. One source estimates
that 4,000 to 5,000 kg-cal/kg of zinc are required.6 Another source
reports energy efficiency to be about 10 percent, with typical energy
consumption 5.0 million kg-cal/metric ton of zinc produced. One company
using vertical retorts finds that 9.1 million kg-cal of coal and coke
per metric ton of zinc produced is required for the briquetting process.
However, this company claims an energy consumption by the retorts of
only 2.8 million kg-cal/metric ton of distilled zinc, for a total energy
consumption of 12.8 million kg-cal/metric ton.? Energy requirements for
refining are the same as for horizontal retorting, about 834,000 kg-
cal/metric ton of zinc produced.8
5. Waste Streams - As with horizontal retorting, emissions are minor
when compared with other steps in the smelting process such as roasting
and sintering. $03 emissions average less than 50 ppm.9 Flow rate for
the carrier gas is 23,000 cu m/metric ton of product, with 2.5 to 3.0
percent carbon dioxide.3
Particulate emissions are evident only during charging for approxi-
mately one minute. High-efficiency metal recovery is possible from
these metal and metal oxide fumes. Particulate emissions are greater
than with horizontal retorting, averaging about 50 kg/metric ton of zinc
produced. Zinc oxide is the principal constituent of these emissions
along with cadmium, copper, chromium, lead, and iron.
The zinc and coke content of the feed and the air flow rates are
the important process variables. Temperature is the most important
process parameter.
295
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The gas washing water contains zinc and metal oxides, possibly
hydrocarbons, various participates (as suspended solids), and the corre-
sponding products of hydrolysis.
Residues are also generated in vertical retorting operations.
Amounts produced are similar to horizontal retorting, around 1050 kilo-
grams per metric ton of zinc produced.3 The residues contain a variety
of metals such as lead, copper, silver, gold, nickel, germanium, gallium,
arsenic, antimony, cadmium, zinc, indium, silicon, iron, calcium,
aluminum, magnesium, and manganese. Processing of the residues to
remove elements such as germanium and gallium is as described for hori-
zontal retorts, and is not done at primary zinc smelters.
An analysis of the potentially hazardous constituents in the resi-
due from one vertical retort furnace is given below:3
Constituent
Cadmium
Chromium
Copper
Lead
Zinc
Concentration, ppm
850
46
4,600
2,400
107,000
Blue powder production amounts to only about 3 percent of the zinc
charged, considerably less than in horizontal retorting.10 The residue
also contains less zinc.
In the redistillation system, no zinc vapors can escape since it is
a closed circuit. Solid residues can be reprocessed to recover zinc and
other metals. Waste gases are produced by the combustion, their composi-
tion depending on the type of fuel used.
6. Control Technology - Wet scrubbers are the available control method
for particulate emissions. All gases are exhausted from the furnace by
means of a venturi scrubber. The carbon monoxide from the zinc conden-
sation chamber is scrubbed with water sprays to remove entrained solids.
The gas is then used as part of the fuel for heating the retorts.
Metallic zinc and zinc oxide is recovered as blue powder residue from
the scrubbing system and from the condenser during periodic cleaning.
The blue powder is recycled.
Residues are either disposed of on open slag dumps or processed to
recover their metallic values. As with horizontal retorting, the best
control technology for the slag dump is sealing the soil and routing
runoff to a waste treatment lagoon.
296
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7. EPA Source Classification Code - 3-03-030-05
8. References -
1. Schlechter, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and.Sons, Inc. New York
1967. Volume 22. pp. 555-603.
2. McMahon, A.D. et al. The U.S. Zinc Industry: A Historical
Perspective. Bureau of Mines Information Circular 9629.
United States Department of the Interior. 1974.
3. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II, Primary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft -
April 1975). pp. 81-127, 239-250.
4. Field Surveillance and Enforcement Guide for Primary Metal-
lurgical Industries. EPA-450/3-73-002. Office of Air and
Water Programs, U.S. Environmental Protection Agency. Re-
search Triangle Park, North Carolina. December 1973. pp.
269-309.
5. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. Effluent Guidelines
Division Office of Water and Hazardous Materials. U.S.
Environmental Protection Agency, Washington. November 1974.
6. Restricting Emission of Dust and Sulfur Dioxide in Zinc Smelt-
ers. Association of the Metal Smelting and Refining Trade and
Committee on Zinc of the German Ore Smelting and Mining Soci-
ety VDI No. 2284. September 1961.
7. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. Battelle Columbus Laboratories. EPA 68-01-
1518. Draft Data.
8. Fejer, M.E., and Larson, D.H. Study of Industrial Uses of
Energy Relative to Environmental Effects. EPA-450/3-74-044.
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina July 1974. pp XII-1 to XII-16.
297
-------
9. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA-450/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Protec-
tion Agency. Research Triangle Park, North Carolina. October
1974. pp. 3-125 to 3-169, 5-18 to 5-27, 6-66 to 6-105.
10. Water Pollution Control in the Primary Nonferrous Metals
Industry - Vol. 1 -Copper, Zinc, and Lead Industries. EPA-R2-
73-247a. Office of Research and Development, U.S. Environ-
mental Protection Agency. Washington, D.C. September 1973.
298
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PRIMARY ZINC PRODUCTION PROCESS NO. 9
Electric Retorting
1. Function - Electric retorting is a continuous reduction/volatiliza-
tion process in which electricity supplies the energy needed to produce
high-purity zinc from zinc oxide by reduction with carbon at elevated
temperatures in a vertical cylindrical retort. Horizontal (Process No.
7) and vertical (Process No. 8) retorts are other reduction methods in
use at zinc smelters. This newest purely zinc-smelting furnace was
designed to overcome the difficulties involving in heating the charge
externally. As in the other retorting processes, carbon monoxide is
used for direct reduction, producing zinc vapor. The carbon dioxide
also produced in this reaction is regenerated with carbon. The carbon
for reduction is provided by coke.
There are several variations on the resistance-type electric
furnace available, including the electrothermic arc furnace (Sterling
Process) and the Imperial Smelting Furnace. However, the only electric
retorts in use in the United States are the electrothermic furnaces
developed by the St. Joe Minerals Corporation, which began commercial
operation in 1930.'
The St. Joe electrothermic furnaces are basically vertical, refrac-
tory-lined cylinders. The largest furnaces now in use have an inside
diameter of 1.5 meters and are 15 meters high, with a production capa-
city of about 90 metric tons per day.2 Graphite electrodes protrude
into the shaft, and the reaction heat is generated from the resistance
of the furnace charges to the current flow between the electrodes.
Eight pairs of electrodes introduce power into the furnace. Each top
electrode has a mate near the bottom.
Preheated coke and sinter, along with miscellaneous minor zinc-
bearing products such as blue powder, are fed continuously into the top
of the furnace from a rotary feeder. As in a vertical retort, gravity
moves the charge downward through the shaft. Unlike other retorting
processes, an unusually hard sinter is required to maintain strength and
porosity in the tall columns, even after most of the zinc content has
been removed. Silica is usually added to the sinter mix to increase its
structural strength. The coke serves as the principal electrical con-
ductor, carrying the alternating current between each top electrode and
the bottom electrode on the opposite side. The heat developed provides
the energy required for smelting. The zinc vapor and carbon monoxide
produced pass from the main furnace to a vapor ring, which provides a
free space around the periphery of the charge for removal of the gaseous
mixture. The gas then goes to a condenser, where zinc is recovered by
bubbling through a molten zinc bath. It was the development of this
Weaton-Najarian vacuum condenser that first made possible the production
of over 90 metric tons per day from a single unit. If necessary, further
299
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refining, such as the liquation and redistillation steps described for
the other retorting processes, may be used.
The electrothermic furnace has a number of advantages over other
processes. First, the increased thermal efficiency (compared with
external heating methods) results in cost savings in fuel consumption.
Larger quantities of charge can be treated, and the continuous operation
is amenable to automation. The furnace can readily process secondary
zinc scrap and zinc residues. In addition, because of special deleading
by heat treatment in multiple-hearth roasters followed by desulfuriza-
tion in fluidized-bed roasters, the electrothermic furnaces emit prac-
tically no S02 or particulates.
2. Input Materials - The feed input consists of sinter, coke, and
recycled zinc-bearing products, such as blue powder. In addition,
silica is usually added to increase structural strength. Particle size
is controlled so as to provide coke particles larger than sinter, there-
by concentrating larger coke at the axis of the furnace. In this way
the maximum fraction of electric current is directed along the axis,
which becomes the region of maximum temperature, minimizing both damage
to the refractory walls by slagging and heat loss through the walls.
Quantities of coke required range from 0.5 to 0.8 ton per ton of slab
zinc produced.3
3. Operating Conditions - The St. Joe electrothermic furnaces operate
at atmospheric pressure. Internal temperatures are 1400°C and higher at
the axis of the furnace, 1200°C in the main body of the charge, and
900°C near the wall. A vacuum of 15 to 25 centimeters Hg is applied to
the outlet of the condenser, causing the vapor-gas mixture to be drawn
through it in large bubbles.4 Water-cooled hairpin loops at the con- ?
denser cooling well maintain a constant batch temperature of 480-500°C.
Temperatures for purification steps vary, since separation is based on
differences in boiling points.
4. Utilities - Electrothermic furnaces use electricity to supply the
energy for reduction. Current through each of the 30-cm-diameter
graphite electrodes may range as high as 800 amperes. A furnace con-
tains eight individual single-phase circuits, each typically carrying
11,000 to 17,000 kg-cal/min. The working range for such a circuit is
250 to 160 volts. The overall power factor of transformers, bus system,
and furnaces is from 90 to slightly over 95 percent. One plant using
electrothermic furnaces reported energy consumption averaging 2.4
million kg-cal/metric ton of metal. Power input per electrode circuit
ranging up to 18,000 kg-cal/min corresponding to a maximum of 143,000
kg-cal/min per furnace, has been reported.5 Energy consumption for
subsequent refining steps is the same as that used in other retorting
processes.
300
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Another source reports 25 to 30 percent energy efficiency for
electrothermic retorting, with the process consuming about 2.8 million
kg-cal/metric ton of zinc. With an additional 5.8 million kg-cal/metric
ton of zinc consumed as coke, a total of 8.6 million kg-cal/metric ton
of zinc is used. If the energy for electric generation is also con-
sidered (assuming 33 percent efficiency), total energy consumption
increases to 14.2 million kg-cal/metric ton.6
5. Waste Streams - As with other types of retorting, emissions are
minor relative to those from roasting or sintering.
Particulate emissions consist primarily of metal and metal oxide
fumes. Particulate emission levels for electrothermic retorting average
about 10 kg/metric ton of zinc produced. As in the other retorting
processes, zinc oxide is the principal constituent of these emissions,
along with cadmium, copper, chromium, lead, and iron. Also as with the
other processes, the important process variables are the zinc and coke
content of the feed and the air flow rates. Temperature, which is
controlled by regulating current flow to the electrodes, is the most
important process parameter.
Temperature of the gases vented from the furnace vapor ring aver-
ages 850°C. Gas composition is approximately 45 percent zinc vapor and
45 percent carbon monoxide; the balance is nitrogren, carbon dioxide,
and hydrogen.5
The gas washing water contains the same impurities as in horizontal
and vertical retorting.
The residues generated during electrothermic reduction are similar
in composition and quantity to those from the other retorting processes.
Germanium and gallium are extracted from the residues by various methods
at other facilities.
Waste streams from further purification steps such as liquation and
redistillation are the same as described for horizontal retorting.
6. Control Technology - High-velocity impingment type scrubbers are
used to clean gases from the condenser. The clean gas, containing 80
percent carbon monoxide and having a heating value of 2200 kg-cal/ra3,
furnishes fuel for smelter use.5 Some blue powder or uncondensed zinc
and zinc oxide is later recovered by settling the scrubber slurry in
ponds. The solids (75 to 80 percent zinc) are dried and briquetted for
furnace feed.
Residue is removed from the furnace, preferably as discrete solid
particles. It goes to a reclamation plant, where residual coke and some
unreacted zinc are recovered and recyled. Besides permitting recovery
of residue containing enough zinc and carbon to make retreatment worth-
301
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while, a minimum of power is consumed in unproductive melting of re-
sidue. At the reclamation plant, where sand may be added to make a hard
sinter, sufficient ferrosilicon is present in some residues to warrant
recovery as a by-product.
7. EPA Source Classification Code - None
8. References -
1. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA-450/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Pro-
tection Agency. Research Triangle Park, North Carolina.
October 1974. pp. 3-125 to 3-169, 5-18 to 5-27, 6-66 to
6-105.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. Effluent Guidelines
Division Office of Water and Hazardous Materials. U.S.
Environmental Protection Agency, Washington. November 1974.
3. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II, Primary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft •
April 1975). pp. 81-127, 239-250.
4. Schlechter, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York
1967. Volume 22. pp. 555-603.
5. Lund, R.E. et al. Josephtown Electrothermic Zinc Smelter of
St. Joe Minerals Corporation. AIME Symposium on Leao and
Zinc, Vol. II. 1970.
6. Fejer, M.E., and Larson, D.H. Study of Industrial Uses of
Energy Relative to Environmental Effects. EPA-450/3-74-044.
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. July, 1974. pp XII-1 to XII-16.
302
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PRIMARY ZINC PRODUCTION PROCESS NO. 10
Oxidizing Furnace
1. Function - In the direct, or American, process for zinc oxide pro-
duction, zinc vapor from sintering is immediately oxidized without being
condensed. As with zinc metal production, zinc must first be produced
in vapor form. Only the direct method of producing zinc oxide is dis-
cussed because other production methods, such as the French process,
start with slab zinc, the major product of the primary zinc industry.
The three types of furnaces used for this purpose in the U.S. are
the grate-type furnace, the rotary or Waelz kiln, and the electrothermic
furnace. In all three, the feed is reduced to form zinc vapor and sub-
sequently the vapor is oxidized and the product collected. Two of the
main difficulties in producing zinc metal, dilution of the zinc vapor
and reoxidation by carbon dioxide, are desirable in the production of
zinc oxide.
In the grate-type furnaces, the coal-sinter feed (usually as bri-
quettes) is spread over grates (traveling or stationary). Coal is
spread first and ignited, then the sinter is deposited on top of the
fuel layer. Air is forced through the bed to support combustion and
furnish a reducing atmosphere to liberate the zinc vapors. The zinc
vapors are then ducted to a combustion chamber, where oxidation occurs
and the zinc oxide product is formed.
The Waelz kiln is a large diameter, long rotary kiln that can be
used for production of pure zinc oxide, although it is more commonly
used for pyrometallurgical concentration of residues of a rather mixed
character. Zinc-bearing material and solid fuel are continuously fed to
the kiln, which typically rotates 1 to 1.5 rpmj The additional heat
for reduction of the contained zinc is supplied by the flow of gaseous
fuel through the kiln. The vaporized zinc then is ducted to a combus-
tion chamber, where air is admitted and the vapor burned to form zinc
oxide. Temperature control is very important, since intimate contact
must be maintained between the solid, zinc-bearing part of the charge,
the solid fuel in the charge, and the reducing atmosphere.
The St. Joe vertical electrothermic furnace may be modified for use
as either a metal or an oxide producer.2 Instead of a large "vapor
ring" or bulge in the furnace barrel midway between the upper and lower
electrodes, the oxide furnace has openings at four levels between the
electrodes, through which the evolved zinc vapor and carbon monoxide
exit the charge. Preheated coke and zinc-bearing sinter are continously
fed to the furnace. The coke serves as the principal electrical con-
ductor. Electricity introduced through the electrodes develops the heat
energy required for smelting. Large furnaces can produce as much as 90
metric tons of zinc per day. Further details on the electrothermic
furnace are given in Process No. 9.
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Analysis of an American process zinc oxide from one plant is as
follows:2
Lead as PbO 0.009%
Cadmium as CdO 0.0101
Iron as Fe20s 0.015%
Manganese as MnO 0.002%
+325 mesh screen residue <0.03%
Brightness; Hunter D-40 93.0 for large sizes
91.0 for fine sizes
The zinc oxide is filtered from the carrier gases in bag collectors.
Soft zinc oxide pellets may be formed by squeezing the oxide between two
rubber rolls to form pellet nuclei.
2. Input Materials - The mix for a grate-type furnace is typically 50
percent sinter briquettes and 50 percent coal.1 The preferred type of
coal is a fine-sized anthracite, used to minimize contamination of the
vapor stream with soot and to prevent caking and slagging.
For the rotary kiln, the mix is usually 65 to 75 percent zinc-
bearing material and the remainder crushed coal or anthracite (to give
about 25 percent carbon in the charge) plus a small amount of sand to
stiffen the bed, if desiredj
The principal feed to the electrothermic furnace also consists of
sinter and coke, but as much as 25 percent of the total zinc input is
other zinc-bearing materials. Nominal coke rate is 44 percent of the
weight of sinter (roughly equal volumes of coke and sinter) and approxi-
mately 300 percent of the stoichiometric carbon relative to the zinc in
the sinter. Other zinc-bearing materials are fed in the form of almond-
shaped briquettes, granules, 8 by 25 millimeter metallic screenings, and
slab dross.2 High product purity is achieved by using low-volatility
cokes and sinter that is low in impurities.
3. Operating Conditions - The grate-type furnace and rotary kiln
operate at atmospheric pressures. Temperatures in the grate-type bed
range from 1000 to 1950°C;l no data are given on operating temperatures
for the rotary kiln.
Electrothermic furnaces also operate at atmospheric pressures.
Operating conditions are the same as in electrothermic furnaces for zinc
metal production, with temperatures at the vapor ring elevation 1200
1400°C in the main smelting zone, 900°C near the walls, and 1300°C at
the bottom electrode elevation.! Further details on operating condi-
tions of electrothermic furnaces are given in Process No. 9.
4- Utilities - No information is available regarding heat input per
unit of product output for grate-type or rotary kiln processes. The bed
304
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is ignited by residual heat from the previous charge in grate-type
furnaces, and gaseous fuel supplies the heat for reduction in rotary
kilns. Air is added in all three furnaces, in unspecified quantities.
The main factor influencing production in electrothemic furnaces is
the quantity of electricity introduced. Josephtown's largest furnace
operates at 143,000 kg-cal/min. Energy consumption averages 2.4 million
kg-cal/metric ton of metal.2
5. Waste Streams - Combustion products and some zinc oxide are emitted
from all three furnaces; again, quantities are unspecified.
A solid residue is produced, estimated at 350 kg/metric ton of zinc
oxide. This residue contains approximately 6.2 percent zinc and 0.09
percent of other potentially hazardous materials such as cadmium,
chromium, copper, and lead.3 it may also contain slag, coke, and
globules of ferrosilicon. One analysis of waste samples from oxide
furnace residue revealed concentrations of potentially hazardous consti-
tutents as follows:3
Constituent
Cadmium
Chromium
Copper
Lead
Zinc
Concentration,
10
17
810
68
62,000
ppm
Further details on waste streams from electrothermic furnaces are given
in Process No. 9.
6. Control Technology - Control techniques for direct zinc oxide
processes are similar to those used in other reduction operations.
Further details are given in Process Nos. 7, 8, and 9. Uncondensed zinc
(blue powder) in gases from the condenser is recovered by settling the
water slurry in ponds. Solids can be dried and briquetted for furnace
feed.
7. EPA Source Classification Code - None
8. References -
1. Schlechten, A.M., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22. pp. 555-603.
305
-------
2. Lund, R.E., et al. Josephtown Electrothermic Zinc Smelter of
St. Joe Minerals Corp. AIME Symposium on Lead and Zinc, Vol.
II. 1970.
3. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II, Primary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft -
April 1975). pp. 81-127, 239-250.
306
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PRIMARY ZINC PRODUCTION PROCESS NO. 11
Leaching
1. Function - The purpose of leaching, the first step in electrolytic
zinc processing following roasting, is to selectively dissolve as much
of the zinc as possible, precipitating iron and accompanying impurities
without precipitating any of the dissolved zinc. Basically the zinc
oxide within the calcine reacts with sulfuric acid to form zinc sulfate
(ZnSOd) in solution. Because other elements also go into solution,
purification must follow the leaching step. The dilute sulfuric acid
used in the process is provided by spent electrolyte recycled from the
electrolysis process.
Although the leaching step varies somewhat depending on concentrate
impurities, the basic process of selectively precipitating the impuri-
ties from the leach solution remains the same. The chemical reaction
converts the metal oxides to the respective metal sulfates. The soluble
metal sulfates, among them zinc, copper, cadmium, arsenic, antimony,
cobalt, and nickel, go into solution. The insolubles suspended in this
aqueous mixture include lead, silver, gold, iron, silica, and calcium.
They are first separated by filtration and the dewatered residue from
the settling tank is sent to the smelter, where the lead, gold, and
silica are recovered.
Two general types of leaching are practiced, single and double. In
single leaching all of the calcine is added to sufficient spent electro-
lyte. The acidic solution must then be neutralized with lime or lime-
stone to cause the precipitation of iron, silica, antimony, and arsenic.
A thickener is generally added to coagulate the solids, simplifying the
overflow operation. Since zinc will also precipitate after these im-
purities reach a minimal concentration, the amount of neutralizer must
be determined carefully, depending on the composition of the calcine.
Calcium and sulfuric acid will form calcium sulfate beyond the neutral
point, resulting in unnecessary loss of acid. Single leaching is usu-
ally justified by savings in capital investment and in operating cost,
but this must be balanced against the risk of lower recoveries, higher
acid requirements, and other factorsJ
In double leaching, the first operation is a neutral leach, which
both extracts the easily soluble zinc and precipitates such impurities
as silica, iron, and alumina. The solution becomes alkaline with the
addition of sufficient calcine to produce a large excess of zinc oxide.
This residue is collected and sent to the acid tanks for a second leach.
An excess of electrolyte is used to ensure dissolution of all soluble
zinc and thus maximize zinc recovery. The liquors of first and second
leaches are combined.
307
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Effective extraction of the impurities depends largely on the iron
content of the calcine. Ferric hydroxide is formed during the neutral
leach, resulting in a heavy floe that effectively traps arsenic and
antimony. In iron-deficient calcine, ferric sulfate solution and scrap
iron may be added for compensation. When considerable zinc has been
made relatively insoluble during roasting by the formation of zinc
ferrites, the practice is to leach with comparatively strong spent
electrolyte to dissolve both zinc and iron, and then precipitate the
iron as jarosite (kFe3(OH)g(SO^)2).
The primary chemical reaction occuring in leaching the calcine is:
ZnO + H2S04 -»• ZnS04 + H20
Typical secondary reactions include oxidation of ferrous iron to ferric,
followed by precipitation of ferric hydroxide as the addition of more
calcine raises the pH of the solution:
FES0 + 2 HS0 -> Zn0 + Fe(S0) + ZnS0
4
4
Fe2(S04)3 + 3 ZnQ + 3 H20 + 2 Fe(OH)3 + 3 ZnS04
Cyanide is added to remove precious metals. Air is then pumped
through the tanks to achieve violent agitation combined with excellent
aeration. The aeration serves a twofold purpose: (1) cyanidation of any
precious metals present, and (2) oxidation of any ferrous iron to the
ferric state.
The leach should remove silica, iron, and alumina. It should also
remove antimony, arsenic, and germanium, although they may not be
removed completely if originally present in appreciable quantity. The
major detrimental impurities remaining are copper, cadmium, and cobalt.'
The trend in the U.S. is to leach continuously or semicontinuously
rather than batchwise, since this requires less equipment, spa^e, and
labor and is better adapted to automatic control.
2. Input Materials - Input materials to the leaching process are the
substantially sulfide-free, finely calcined feed from roasting, other
zinc oxide products, milk of lime or finely ground limestone, and sul-
furic acid. The acid is normally spent electolyte of sufficient con-
centration to yield a 0.3 to 0.5 percent excess of sulfuric acid after
all soluble zinc has been dissolved J
The electrolyte may contain around 200 grams per liter of sulfuric
acid.^ A thickener may be added to coagulate the solids, and a ferric
sulfate solution or scrap iron may be added to iron-deficient calcine.
In some plants, a prel caching step with acid is performed to reduce the
308
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magnesium content of the concentrates and prevent introduction of mag-
nesium sulfate into the electrolytic plant. Cyanide in unspecified
quantities is added for recovery of precious metals.
3. Operating Conditions - Batch leaching is done at ambient pressure,
whereas continuous leaching takes place at 1.4 to 2.5 kg/sq cm.' Both
proceed at ambient temperatures, although since the reaction of the ZnO
with HgSCty is exothermic, the solution temperature will increase. The
magnitude of this temperature increase is not stated in the literature.
The spent electrolyte and calcine are added to the leaching tanks with
acidity control both to avoid dissolving an excess of iron and then to
precipitate the iron that is dissolved.
4. Utilities - Electricity is required for pumping the solution and
conveying the calcine and final residue. No estimates of energy con-
sumption for the leaching process are given. In batch-leaching, air at
about 6.3 kg/sq cm must be available to clear out accumulating deposits
of coarse material at the bottom of the tanks. In continuous leaching,
it is seldom necessary to use air at pressures higher than the normal
operating pressure of 1.4 to 2.5 kg/sq cmJ Continuous leaching requires
a larger total volume of air per tank, because agitation must be con-
tinuous. During agitation in either process, consumption ranges from
2.8 to 4.2 cubic meters of air per minute.'
5. Waste Streams - After leaching, the neutral or pregnant solution is
filtered and a solid residue remains. This is in the form of a filter
cake, high in metals value. About 360 kilograms of residue is generated
per metric ton of zinc produced. When a preparatory acid leach is used
to remove magnesium from the concentrates, a high amount of magnesium
sulfate is introduced into the process wastewater circuit.
6. Control Technology - The solid residue remaining after leaching is
usually shipped to a lead smelter for further smelting and therefore
presents no disposal problem. The leached residue may also be treated
in the same manner as a calcine in which zinc has been made relatively
insoluble in roasting by the formation of zinc ferrites; it can be
leached with hot, comparatively strong spent electrolyte.
Part of the indium is concentrated in leach residues and sent to
lead smelters for further treatment. Indium may be reclaimed from
bullion dross or zinc fume derived from lead slag by some combination of
leaching, precipitation, and electrodeposition.
Storage of the residues in unlined lagoons before being dredged for
shipment to the lead smelters could result in a possible leachate pro-
blem. The best available control technology is the use of lined la-
goons, with immediate shipment of dredged sludge to the smelters. Cost
of this technology, which would also handle precipitates from the
electrolysis process, is estimated at $0.15 per metric ton of zinc
produced/
309
-------
7. EPA Source Classification Code - None
8. References -
1. Schlechten, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York
1967. Volume 22. pp. 555-603.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75-032. Effluent Guidelines
Division Office of Water and Hazardous Materials. U.S. EPA,
Washington. November 1974.
310
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PRIMARY ZINC PRODUCTION PROCESS NO. 12
Purifying
1. Function - Purification removes unwanted elements that remain after
the leaching process - mainly copper, cadmium, and cobalt - by filtra-
tion and precipitation. The method of treating the solution depends
partly upon the extent to which these impurities are present.
The insoluble materials are first .separated from the solution by
filtration. This is usually done in large drums or filters. The
dewatered residue from the settling tank contains lead, gold, and
silver.
Purification is accomplished largely by additions of zinc dust,
which precipitates copper, cadmium, cobalt, nickel, and other residue by
replacement. The filtered zinc sulfate solution is piped to an agitator
tank, and enough zinc dust is added to precipitate the copper. The
copper precipitate is removed from solution by filtration. Addition of
zinc dust in multiple stages allows rough separations such as a high-
copper precipitate and a high-cadmium precipitate, each carrying down
some of the other impurities. Other remaining metals such as cobalt,
nickel, antimony, and arsenic are then precipitated out of the solution.
If the cobalt content of the solution is low, it can be precipitated
with zinc dust and later separated from the copper-cadmium cake. Solu-
tions with higher cobalt contents are precipitated with l-nitroso-2-
naphthol.'
Copper and arsenic may be added for the purpose of providing addi-
tional seed precipitate to bring down other minor impurities with the
copper or arsenic when zinc dust is added.
2. Input Materials - The principal inputs to the purification process
are the pregnant leach solution and multiple stages of zinc dust.
Copper sulfate and arsenic may be added, and l-nitroso-2-napthol may be
added to precipitate cobaltJ
3. Operating Conditions - Purification is carried out at ambient
temperatures and pressures.
4. Utilities - Electricity is required to pump the solution. No
process-specific estimates of energy consumption were found.
5. Waste Streams - A copper cake residue is formed, which contains
cadmium and sometimes such elements as indium, thallium, gallium, and
germanium. The dewatered residue from early filtration steps contains
lead and precious metals. As with the residue from pyrometallurgical
retorts, this iron mud can be processed by various methods to remove the
desired elements. Such processing takes place at facilities other than
the primary zinc smelters.
311
-------
6. Control Technology - The filtration residue containing lead, gold,
and silver is usually sent to a lead smelter.
After having been treated to recover copper cake and a cadmium pro-
duct, the remaining residue is usually recycled to leaching. The copper
cake is sent to a copper smelter for processing, and the cadmium and
other by-products are recovered. Residues containing sufficient concen-
trations of other elements are shipped to other locations for recovery.
Two companies are reported to have refined indium from these residues in
1975.2
7. EPA Source Classification Code - None
8. References -
1. Schlechten, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York
1967. Volume 22. pp. 555-603.
2. Commodity Data Summaries 1976. U.S. Department of Interior,
Bureau of Mines. Washington, D.C., 1976.
312
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PRIMARY ZINC PRODUCTION PROCESS NO. 13
Electrolysis
1. Function - In electrolysis, metallic zinc is recovered from the
purified solution by passing current through an electrolyte solution,
causing zinc metal to deposit on a cathode. The discovery of the com-
plex zinc-lead-silver ores of British Columbia and Montana promoted
research on electrolytic processing of zinc. These ores yielded a 30 to
40 percent zinc product and were not suitable as charge to retorting
processes.
Electrolytic zinc extraction represents approximately 56 percent of
the free world's zinc production capacity and about 42 percent of U.S.
primary zinc capacity.'.2,3 The advantages of electrolytic extraction
are its amenability to automation and its ability to treat lower-grade
concentrates and produce a high-grade metal without subsequent refining.
In addition, the process alows ready recovery of by-products and elmi-
nates dust and heat from furnaces other than roasters. Perhaps the
greatest advantage is the relative ease with which air pollutants such
as S02 and particulates can be controlled. S02 emissions are generated
only in the roasting operations.
In the past, two major classes of electrolytic plants were oper-
ated: (1) low-current-density, low-acid strength circuits, and (2) high-
current-density, high-acid strength circuits. In the low form, the
current density averaged 1070 amp/sq m of cathode area in solution,
whereas in the high form current density was normally above 3570 amp/sq
m. Acid strength was about 6 percent ^804 for the former and 25 per-
cent for the latter. Modern plants are typically modifications of the
original high-density plants with intermediate values of current density
at about 2140-2500 amp/sq m.4
The electrolytic cell room is a large area containing tanks through
which the zinc-containing solution, or electrolyte, slowly flows. Each
tank contains a number of alternate anodes and cathodes. Anodes are
rectangular, commonly made of cast lead containing 0.75 to 1.0 percent
silver. Cathodes are aluminum, with slightly smaller dimensions. All
anode and cathodes of a given cell are connected electrically in paral-
lel. Spacing of the electrodes is as close as possible, since a drop in
voltage is directly proportional to the separation.
Zinc is deposited from solution onto the aluminum cathodes at a
rate governed mostly by the current density. The considerable heat that
develops in the electrolyte is reduced by internal cooling coils or
external cooling systems. The current efficiency of less than 100
percent causes release of oxygen at the anodes and hydrogen at the
cathodes. Bursting of these bubbles causes an acid spray which corrodes
313
-------
equipment. Sodium carbonate (or barium hydroxide) is added to the
electrolyte in some plants to reduce lead contamination of the deposited
zinc.
Any irregularity in surface or shape of the electrodes will cause
slight variations in voltage gradient at localized points; deposition of
zinc tends to start where the voltage gradient is greatest. Once such
an uneven deposition starts, the tendency becomes self-accelerating
because of the locally lowered resistance; the final result is a short
circuit. To minimize this effect, a protective colloid such as glue,
goulac, gum arabic, or agar agar plus sodium silicate and cresylic acid
is added to provide a smoother deposit and reduce interference by
impurities. Typical amounts are 0.42 kg/metric ton of zinc deposited.
The colloid possesses an electric charge and apparently concentrates in
areas of high current density. It increases resistance locally, thus
tending to equalize deposition over the entire cathode.
When the deposit attains a desired thickness, for example in 24
hours, the cathodes are removed and zinc is stripped. Stripping is
usually done manually, which is expensive, although various mechanical
means and air blasting have been used experimentally. The cathode zinc
is washed and sent on to casting. After passing through a series of
cells, the electrolyte is recirculated. Some portion of the spent
electrolyte is sent back to the leaching plant continuously.
Deposition of zinc is more rapid in high-strength than in low-strength
systems. To prevent short circuits, stripping periods are much shorter
in the high-strength system, and labor expenses are higher. The high-
strength system also requires a greater investment per unit of capacity.
An advantage is that the stronger acid dissolves zinc ferrite and leads
to a deposition of higher-purity zinc. Other advantages of the high-
strength system include improved filtration and production of more zinc
per unit of cathode area.
Although most zinc concentrates can be processed in electrolytic
zinc smelters, certain impurities can cause extraction problems.
Concentrates containing germanium, cobalt, tellurium, iron, and other
minor elements require special treatment and are, therefore, less
desirable for electrolytic extraction. Germanium can,make commercial
extraction difficult. Cobalt is the most troublesome impurity to remove
and will magnify the effects of any germanium in the same solution.
Tellurium deposits with the zinc and reduces its purity. Processing of
concentrates containing these impurities increases cost of production,
since special variations of the leaching and purification process plus
strict processing controls are required.
An electrolytic plant requires high capital costs and also is not
suited for all locations and production requirements. Feasibility in a
specific application depends on both the availability and cost of
electrical power. Electrolytic zinc requires debasing to fill consumer
requirements for lower quality prime western or intermediate grade zinc,
314
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the initial product from pyrometallurgical processes. Finally, electro-
lytic zinc extraction is basically limited to the processing of concen-
trates. Processing of non-oxide residues, metallic residues, and re-
cycled scrap generally is not possible.
2. Input Materials - The required materials are zinc electrolyte,
anodes, and cathodes. Anodes are typically cast lead containing 0.75 to
1 percent silver; cathodes are usually aluminum.2 Quantities of electro-
lyte required per ton of zinc recovered vary with concentrations of zinc
ion in the solution. Specific data are not available. Sodium carbonate
or barium hydroxide may be added to the electrolyte to reduce lead
contamination. A colloid such as glue, goulac, gum arabic, or agar agar
plus sodium silicate and cresylic acid may be added.
3. Operating Conditions - The reaction is carried out at room tempera-
ture and pressure.Since the electrolysis is exothermic, cooling coils
or external cooling systems are used to prevent excessive temperatures.
4. Utilities - Electricity is required for pumping the solution and
for electrolysis. The rate at which zinc deposits on the cathode de-
pends on the current density. Typical density is 750 amp/sq m of
cathode area; density may be as high as 1130 amp/sq m.4
The basic electrolytic process is reasonably efficient, with an
estimated energy efficiency of 25 percent. The energy consumption per
metric ton of zinc produced is estimated at 4.5 million kg-cal of natural
gas for melting and 8.5 million kg-cal for electricity including electric
generation, giving a total of 13.0 million kg-cal per metric ton of
zinc.5
5. Waste Streams - Atmospheric emission of pollutants from electro-
lytic zinc reduction is limited to minor amounts of sulfuric acid mist
and participate, estimated to be 3.3 kg/metric ton of zinc produced.
The composition of the particulates is unknown but is thought to include
zinc, lead, cadmium, arsenic, and antimony. No S02 is released during
electrolysis; for the entire electrolytic cycle, S02 emission control
depends primarily upon generation of a strong SO? roaster effluent
stream through the use of modern roasting technology.
The precipitate contains copper, germanium, cadmium, nickel,
arsenic, antimony, cobalt, and iron. The spent electrolyte solution
contains zinc, aluminum, magnesium, calcium, sodium, chlorine, fluorine,
and sulfur. Because of the purification steps applied, no liquid or
solid wastes are discharged from electrolytic process streams.
Analyses of waste samples of anode sludge from one plant are as
follows:6
315
-------
Fresh anode sludge
Old anode sludge (from dump)
Concentration, ppm
Cd
12
1,400
Cr
10
8
Cu
85
1,900
Pb
170,000
89,000
Zn
12,800
39,200
6. Control Technology - In addition to ventilation, various electro-
lyte covers and additives are used with varying degrees of success to
curtail sulfuric acid misting.
The process circuit incorporates purification steps of chemical
precipitation and solids separation, with filter cakes providing the
outlet for impurities and spent electrolyte recirculating to the leach
step. The input of chlorides and fluorides must be carefully controlled
to maintain both the electrolysis reaction and product quality. Quality
of make-up water is usually maintained by use of special wells. The
electrolyte composition may be controlled by extraction of by-product
compounds, dependent on concentrate source. Zinc electrolysis is exo-
thermic; the process includes appropriate cooling loops in heat ex-
changers of flash cooling capability. All spills and wash water are
considered valuable metal-bearing material and are recycled to the
leach-purification circuit.
All precipitates from the solution are shipped to either copper or
lead smelters or to a cadmium recovery process. As with leaching re-
sidues, these precipitates are stored in unlined lagoons and are dredged
for shipment to the smelters. If leaching is a problem, the best avail-
able control technology is use of lined lagoons and prompt shipment of
dredged sludge to lead smelters. Cost of this technology is estimated
at $0.15 per ton of zinc produced.2
7. EPA Source Classification Code - Electrolytic Process 3-03-030-06.
8. References -
1. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. EPA-405/2-74-002a.
Office of Air and Waste Management, U.S. Environmental Pro-
tection Agency. Research Triangle Park, N.C. October 1974.
pp. 3-125 to 3-169, 5-18 to 5-27, 6-66 to 6-105.
2. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. Effluent Guidelines
Division Office of Water ahd Hazardous Materials. U.S. EPA,
Washington. November 1974.
316
-------
3. American Bureau of Metal Statistics. Non-ferrous Metal Data
1975. New York, New York. 1975. pp. 65-87.
4. Schlechten, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York
1967. Volume 22. pp. 555-603.
5. Study of Industrial Uses of Energy Relative to Environment
Effects. EPA-450/3-74-044. U.S. Environmental Protection
Agency. Research Triangle Park, N.C. July 1974. pp. XII-1
to XII-16.
6. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II, Primary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft-
April, 1975). pp. 81-127, 239-250.
317
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PRIMARY ZINC PRODUCTION PROCESS NO. 14
Melting and Casting
1. Function - The process involves melting and casting the zinc from
an electrolytic or pyrolytic plant into a marketable form. Pyrolytic
zinc is usually in a molten state whereas electrolytic stripped zinc
sheets must always be melted. Recent practice utilizes induction
heating, possibly combined with a gas-fired furnace, and the molten zinc
is cast into 27-kg slabs on an in-line casting machine. Some zinc is
cast into 640- to 1100-kg blocks in stationary moldsJ
Cathode zinc sheets from electrolytic plants are dried, melted, and
cast into various forms of slab zinc. Alloys of zinc are also prepared
and cast. Depending on market conditions, lead and other constituents
may be added to a relatively high grade of zinc to make a select grade
for galvanizing. Zinc dust is made at the plants for use in purifica-
tion of solutions.
Because molten zinc exhibits a strong tendency to form dross,
ammonium chloride flux is usually added to the melting furnace to retard
oxidation at the surface and to collect any oxides formed. Ordinary
slab zinc an be melted in a reverberatory furnace with the formation of
1 percent or less of dross, whereas electrolytic cathodes lose 6 to 17
percent under these circumstances, depending on the presence of glue,
cobalt, and the like in the electrolyte.l
2. Input Materials - Inputs are zinc, ammonium chloride flux, and
various alloying materials to meet special requirments. Quantities are
not specified.
3. Operating Conditions - Melting and casting are at atmospheric
pressure.Zinc must be heated above its melting point (420°C) to form
liquid zinc.
4. Utilities - Gas- or oil-fired melting pots are used for melting
zinc.The heat requirments are not specified. Water is used in some
plants to cool the molds rapidly, but usually does not come into contact
with the metal.
5. Waste Streams - Some zinc oxides and chloride compounds are emitted
by the melting pots into the atmosphere; quantities are not given. To
achieve high overall melting efficiency, dross is skimmed from the
melting pot so that the globules of metal may be separated from thin
oxide shells.
Casting cooling water generally contains suspended solids and oil
and grease in terms of metal oxides, mold washes, and lubricants from
casting equipment.
318
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6. Control Technology - Control of atmospheric emissions is with
scrubbers or fabric filters. Any processing to recover the zinc values
from the control residue must deal with its chloride content. In
electrolytic plants the reaction is very sensitive to the deleterious
effects of chloride.
The dross is retreated by liquation, centrifugal separation of
metal, fluxing, etc. Recovered metal may be used to make zinc dust for
electrolytic purification or it may be returned to the melting pots.
7. EPA Source Classification Code - None
8. References -
1. Schlechter, A.W., and A. Paul Thompson. Zinc and Zinc Alloys.
In: Kirk-Othmer. Encyclopedia of Chemical Technology.
Interscience Division of John Wiley and Sons, Inc. New York,
1967. Volume 22. pp. 555-603.
319
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PRIMARY ZINC PRODUCTION PROCESS NO. 15
Cadmium Leaching
1. Function - Leaching selectively dissolves as much cadmium as possi-
ble from various cadmium-bearing dusts, fumes, and sludges from primary
zinc plants, precipitating lead and other impurities without precipi-
tating any of the dissolved cadmium. There is no separate primary
cadmium industry in the United States. Cadmium-recovery processes use
by-products of other operations, all involving zinc, in four major
categories:"!
(1) Fumes and dusts from roasting and sintering of zinc concen-
trates.
(2) Recycled zinc metal containing cadmium.
(3) Dusts from smelting of lead-zinc or copper-lead-zinc ores.
(4) Purification sludge from electrolytic zinc plants.
The first and fourth of these categories are of concern here, since they
provide the inputs to the leaching process. Since the second category
involves recycled zinc metal, it is not relevant here; cadmium-bearing
dusts from lead and copper smelting, the third category above, are sent
directly to cadmium purification, Process No. 17.
During sintering, cadmium, lead, and thallium chlorides form and
are drawn off as a fume to be recovered in ESP's. After collection, the
fumes and dusts are leached with dilute sulfuric acid and sodium chlo-
rate to ensure complete dissolution of cadmium sulfide. Cadmium goes
into solution by the following reaction:
CdO + HoSO, + CdSO, + H00
'2
yi ' vWOWyi • I IrtV
Sodium chlorate is a strong oxidizing agent, added to prevent reduction
of any sulfur to sulfide and to prevent reprecipitation of cadmium or
thallium as sulfides. Cadmium and lead are converted to sulfates and
chlorides. The cadmium compounds remain in solution, but lead is almost
completely converted to insoluble lead sulfate. This is filtered out
and sent to lead recovery together with other insoluble materials such
as quartz and silicates. The lead residue may contain small but sig-
nificant quantities of gold, silver, and indium.2 More sulfuric acid is
then added to the solution to bring the concentration up to 10 percent
and to raise the temperature.
320
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Instead of a direct leach with sulfuric acid, in at least one plant
the dust is first roasted and then water leached. The sinter fume is
heat-treated in a four-hearth roaster, which selectively sulfates the
cadmium and makes about 90 percent of it water-soluble. The water leach
that follows produces relatively pure cadmium solutions containing about
40 grams per liter cadmium and 10 grams per liter zinc. Addition of
sodium bichromate to this solution removes about 90 percent of the
soluble lead. The residual solids are batch treated with scrubber
liquor and concentrated acid.3
Another element in the flue dusts from roasting and sintering that
is economically recoverable is thallium; this processing is not done at
any primary zinc smelters. One refiner in Colorado is the sole domestic
producer of thallium.4
Another source of high-cadmium residues for leaching is electro-
lytic zinc processing. In the purification stage, both copper and
cadmium are eliminated from solution by treating the electrolyte with
powdered zinc. Generally this is done stagewise, making possible rough
separations such as high-copper and high-cadmium precipitates. Copper
sulfate and arsenic trioxide may be added during these stages. The
cadmium precipitate is high in zinc and contains virtually all the
cadmium originally present in the electrolyte. This purification sludge
constitutes the raw material for the cadmium plant. It is oxidized,
either by allowing it to stand exposed to air or on a steaming platform.
The acid-soluble cadmium is leached out by sulfuric acid and spent
electrolyte. Filtration to remove insoluble copper may follow.
2. Input Materials - The fumes and dusts from roasting and sintering
processes and the purification sludge from electrolytic plants are the
primary inputs to leaching. Copper sulfate and arsenic trioxide may be
added during pre-leaching purification stages. Dilute sulfuric acid
and/or spent electrolytic are the leaching agents. The electrolyte
contains 200 g/1 sulfuric acid and 65 g/1 zinc. Sodium chlorate is
added to serve as an oxidizing agent, and in some plants water is used
to leach the fumes and dusts. The residual solids from water leaching
are batch-treated with scrubber liquor bolstered with concentrated acid.
Sodium bichromate may be added to remove soluble lead; one plant report-
ed using 7.5 grams per kilogram of cadmium product. Sixty-five grams of
caustic per kilogram of cadmium produced is added.3 Quantities for
other inputs are not specified.
3. Operating Conditions - Cadmium leaching takes place at atmospheric
pressure. A temperature of 80°C is reached after the final addition of
sulfuric acidJ
321
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4. Utilities - Electricity in unspecified quantities is used to pump
liquids. The water leaching process requires 1.3 cubic meters of
natural gas per kilogram of cadmium product to roast the dust, followed
by the addition of an unspecified quantity of water.3
5. Waste Streams - No data were found on the quantities of residuals
from acid leaching, but they are probably very small. Typical analysis
of this residue at one plant is 32 percent lead, 8 percent zinc, 0.7
percent cadmium, 0.13 percent indium, 0.45 percent arsenic, 0.30 percent
copper, 0.23 percent silver, and 4 g/metric ton gold. The residue may
also contain other insolubles such as quartz and silicates.3 Water
vapor is released from purification and leaching. An exhuast gas
temperature of 60°C from leaching was reported at one electrolytic
plant, which also reported exhaust gases from preleaching purification
stages ranging from 60 to 90°C.3
6. Control Technology - Lead sulfate is filtered out from the solution
along with other insolubles. All residues are recycled. They are
typically high in lead content and are sent to a lead recovery plant.
7. EPA Classification Code - None
8. References -
1. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II, Primary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft -
April 1975). pp. 81-127, 239-250.
2. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer.
Encyclopedia of Chemical Technology. Interscience Division of
John Wiley and Sons, Inc. New York 1967. Volume 3. pp. 884-
899.
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. Battelle Columbus Laboratories. EPA 68-01-
1518. Draft data.
4. Commodity Data Summaries 1976. U.S. Department of Interior,
Bureau of Mines. Washington, D.C. 1976.
322
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PRIMARY ZINC PRODUCTION PROCESS NO. 16
Cadmium Precipitation
1. Function - The purpose of precipitation is to treat the cadmium-
zinc sulfate solution from the leaching process with zinc dust to pre-
cipitate cadmium as a metallic sponge and then to separate it from most
of the zinc dust while the solution is agitated. To avoid excess zinc
contamination, usually only 90 to 95 percent of the cadmium in solution
is precipitated. The initial cementation with zinc dust may result in a
liquor containing a residual 0.2 gram per liter of cadmium and 30 to 40
grams per liter of zinc. To further decrease overall cadmium discharge,
the stripped liquor is heated to 40°C and recemented with 1.6 times the
stoichiometric amount of zinc to reduce the liquor to 0.04 gram per
liter of cadmium and 30 to 40 grams per liter of zinc.1
The cadmium sponge is then filter-pressed. It contains about 69
percent cadmium, 30 percent moisture, and small amounts of lead and
zinc.2 It is steam dried or dewatered in a centrifuge. The solution
from filtration, containing practically all of the zinc added and about
10 percent of the cadmium as chlorides and sulfates, is returned to the
sintering operation.
For cadmium production at electrolytic plants, the leach liquor is
first filtered to remove insoluble copper introduced from the electro-
lyte. The filtrate is then precipitated with zinc dust in two or three
stages to minimize zinc concentration in the cadmium sponge. Strontium
carbonate may be added in one of these stages. The sponge will contain
about 80 percent cadmium and less than 5 percent zinc. Because the
electrolysis step that follows is not highly sensitive to the presence
of impurities, the purification step with zinc dust is usually adequate.
If further purification is desirable, cobalt may be precipitated with
nitroso-2-naphthol or potassium xanthate, and thallium precipitated with
potassium chromate or dichromate. The sponge is then oxidized again by
steam drying to enhance the solubility of cadmium and is leached in
spent electrolyte and filtered. The filtrate contains about 200 grams
per liter of cadmium as sulfate and is ready for introduction into the
electrolytic cells.'
2- Input Materials - The cadmium-zinc sulfate solution from leaching
and zinc dust are the principal inputs to precipitation. Strontium
carbonate may be added during purification in electrolytic processing.
Impurities may be removed by the addition of nitroso-2^naphthol, potas-
sium xanthate, and potassium chromate or dichromate. Quantities for
these inputs are not specified.
3- Operating Conditions - Precipitation takes place at atmospheric
pressure. Temperatures range from ambient to 40°C.l
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4- Utilities - Electricity is used to pump liquids, and natural gas or
oil is used to heat the stripped liquor. Quantities were not found.
5. Waste Streams - The filtration solution from this process contains
almost all of the zinc added as well as about 10 percent of the cadmium
as chlorides and sulfates.^ The purification precipitates may contain
arsenic and mercury, both of which are relatively low in concentrates
and ores. One plant reported that a materials balance established 8 ppm
mercury and 23,000 ppm arsenic in an iron precipitate from the cadmium
process, the only place where significant concentrations were found.4
Water vapor is released from both purification and precipitation opera-
tions at temperatures ranging from 45 to 60°C.
6. Control Technology - Much of the residual zinc and cadmium in the
filtration solution can be precipitated by lime treatment. It has been
shown that freshly precipitated cadmium hydroxide leaves approximately 1
milligram per liter of cadmium in solution at pH 8; this is reduced to
0.1 milligram per liter at pH 10. Even lower values of 0.002 milligram
per liter have been shown at pH 11.2 Evidence has been presented that
high levels of iron are beneficial for removal of cadmium during lime
precipitation. The resultant light slurry goes to sludge settlement
ponds, where the solids can settle and be retained for recycling. At
one plant the slurry is mixed with the neutralized roaster scrubber
liquor before being sent to a series of settling ponds.
7. EPA Source Classification Code - None
8. References -
1. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer.
Encyclopedia of Chemical Technology. Interscience Division of
John Wiley and Sons, Inc. New York. 1967. Volume 3. pp.
884-899.
2. Assessment of Industrial Waste Practices in the Metal Smelting
and Refining Industry - Volume II. Primary and Secondary
Nonferrous Smelting and Refining. Calspan Corporation (Draft -
April 1975). pp. 81-127, 239-250.
3. Development Document for Interim Final Effluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. EPA 440/1-75/032. Effluent Guidelines
Division Office of Water and Hazardous Materials. U.S.
Environmental Protection Agency, Washington. November 1974.
4. Development Document for Interim Final Eflluent Limitations
Guidelines and Proposed New Source Performance Standards for
the Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category. Battelle Columbus Laboratories. EPA Con-
tract No. 68-01-1518. Draft data.
324
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PRIMARY ZINC PRODUCTION PROCESS NO. 17
Cadmium Purification and Casting
1. Function - The purification and casting step purifies the cadmium
sponge by melting it with a caustic flux, with distillation and perhaps
redistillation, or by redissolving the sponge with sulfuric acid and
collecting the cadmium by electrolysis.
In pyrometallurgical processing, the dried cadmium sponge is first
mixed with coal or coke and lime. It is then transferred to a conven-
tional horizontal-type retort, where the cadmium is reduced and col-
lected as molten metal in a condenser. Occasionally, for ultra-purity,
the metal is distilled in graphite retorts. Thallium is removed by
treatment with zinc ammonium chloride or sodium dichromate. The metal
may then be cast into a marketable form or further purified by redistil-
lation. Typical impurities in the product cadmium are 0.01 percent
zinc, 0.003 percent copper, 0.015 percent lead, less than 0.001 percent
thallium, less than 0.0005 percent tin, and less than 0.001 percent
antimonyJ Cadmium recovery has been reported as 94 percent from the
feed to the leach plant and 67 percent from the zinc concentrates.
Electrolytic processing of cadmium is carried out in banks of cells
similar to zinc cells. The sponge is first dissolved in dilute sulfuric
acid (return electrolyte). The anodes are lead. The cathodes of
cadmium are 97 percent pure and represent 90 to 95 percent total re-
covery of cadmium from ore to metal. Recovery in the electrolytic step
is 96 percent from the cadmium sponge. The stripped cathode metal is
washed, dried, and melted under a flux, such as caustic or rosin, and
cast into various shapes. Total depletion of cadmium from the solution
is not carried out. When the ratio of the thallium to cadmium sulfate
in the electrolyte reaches 1:10, the cadmium cathodes must be removed
and replaced with new insoluble cathodes. Continuing electrolysis
deposits an alloy containing 5 to 20 percent thallium. These cathodes
are then leached with steam and water to separate thallium into the
filtrate, leaving cadmium as a residue. Small amounts of cadmium in
solution are precipitated with sodium bicarbonate. Thallium is pre-
cipitated with hydrogen sulfide, then dissolved in sulfuric acid. This
sulfate solution can be electrolyzed for recovery of pure sponge thal-
lium. The sponge is washed, pressed into blocks, melted, and cast.
Processing of concentrated lead-smelter baghouse dust in an elec-
trolytic cadmium plant is similar to processing of cadmium sponge.
Since the dust is generally higher in impurities than the sponge, some
additional purification is necessary. The dust is mixed with sulfuric
acid and water to form a paste, then calcined to a sulfated cake which
is crushed and agitated with spent electrolyte. Milk of lime may be
added to neutralize the solution, and sodium sulfide or impure cadmium
325
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sulfide is added to precipitate copper and other metal impurities.
After filtration to remove a lead cake, an agent such as sodium chlorate
oxidizes the iron and lime precipitates iron and arsenic. Heating
ensures complete precipitation. Precipitated thallium chromate or
dichromate is filtered off after the addition of a soluble chromate or
dichromate. Excess chromate remaining may be reduced by sodium sulfide
and precipitated by neutralizing with caustic. The filtered solution is
fed to the electrolytic plant. Where chlorine is present, irons high in
silicon are used as anodes instead of lead. The finished cathodes are
melted and cast.2
2. Input Materials - Cadmium sponge from precipitation or lead smelter
baghouse dust are the principal inputs to purification and casting. In
pyrometallurgical processing, coal or coke and lime or sodium hydroxide
are added, and zinc ammonium chloride or sodium dichromate is used to
remove thallium. One plant reported using 65 grams of caustic per
kilogram of cadmium. In electrolytic processing, dilute sulfuric acid
or return electrolyte, a flux such as caustic or rosin, water, sodium
bicarbonate, and hydrogen sulfide are the inputs. The cell feed may run
100 to 200 grams cadmium, 30 to 80 grams zinc, and 70 to 80 grams
sulfuric acid per liter.I Processing of lead smelter baghouse dust may
entail addition of sulfuric acid, water, milk of lime, sodium sulfide or
cadmium sulfide, sodium chlorate, a soluble chromate or dichromate, and
a caustic, added at different stages. Quantities are not specified for
any of these inputs.
3. Operating Conditions - Cadmium retorting furnaces are not pres-
surizecTOperating temperatures of 455°C have been reported for the gas
melting and ammonium chloride stages, and 790° to 910°C in the furnace
itself. Electrolysis also takes place at atmospheric pressure, with the
temperature held at about 30°CJ
4. Utilities - Electricity, oil, or natural gas is used for melting
the cadmium product. Air is brought into the furnace. Specific quanti-
ties were not found. In electrolysis, current density may ranc,e from
140 to 360 amps per m2.2
5. Waste Streams - Residues from retorting furnaces contain 1.5 to 6.0
percent cadmium, together with varying amounts of zinc and lead.
Filter cakes may also include iron, arsenic, indium, mercury, and cop-
per.2 Thallium is not regarded as a serious impurity since it is re-
moved during processing. A sample of a cadmium filter cake residue was
obtained from one electrolytic plant. After cadmium had been recovered
from the flue dust, this waste amounted to 514 kilograms per day.
Analysis was as follows: cadmium - 280 ppm, chromium - 24 ppm, copper -
1150 ppm, lead - 215,000 ppm, zinc - 39,000 ppm, and thallium - < 40
ppm. The filter cake amounts to 1.8 kg/metric ton of zinc product.3
Particulates are released at several stages in the process. One plant
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reported the release of 20 kilograms per hour of ammonium chloride from
fluxing and 108 kilograms per hour of cadmium from the retorting
furnace.'
6. Control Technology - The filter cake derived from purification is
returned to the sintering operation. In some cases it is disposed of by
open dumping or transferred to unlined ponds for liming and settling.
Sludge dredged from the lagoons is stored on the ground for variable
periods of time before shipment to lead smelters. No plants are known
to use lined lagoons or to treat soil areas where dredged sludges are
stored. In electrolytic processing, the spent electrolyte is returned
to various leaching stages in the circuit. In some cases the electro-
lytic cadmium process cycle is completely closed with no discharge.
Particulates can be controlled with fabric filter systems.
7. EPA Source Classification'Code - None
8. References -
1. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Zinc
Segment of the Nonferrous Metals Manufacturing Point Source
Category. Battelle Columbus Laboritories. EPA Contract No. 68-01-
1518. Draft Data.
2. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer. Encyclo-
pedia of Chemical Technology. Interscience Division of John Wiley
and Sons, Inc. New York 1967. Volume 3, pp. 884-899.
3. Assessment of Industrial Waste Practices in the Metal Smelting and
Refining Industry - Volume II, Primary and Secondary Nonferrous
Smelting and Refining. Calspan Corporation (Draft - April 1975).
pp. 81-127, 239-250.
327
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5.0 AIR MANAGEMENT
The most noticeable pollution from the primary copper, lead, and
zinc industries is the discharge of materials to the atmosphere. Al-
though mining and concentrating operations produce considerable fugitive
dusts, there are no direct atmospheric process discharges. Electrolytic
copper refineries also cause very little atmospheric pollution. It is
the smelters that produce hazardous particulates, toxic fumes, and S02
gas, all of which may be discharged into the air or lost to the sur-
roundings as fugitive emissions.
EMISSION CHARACTERISTICS
Almost every smelting process produces atmospheric emissions.
Despite the variety of processes, many of the emission streams have
common characteristics. In production of these metals, an ore concen-
trate is heated to a high temperature. Solid particles are entrained in
the exhaust gas and are carried from the process. Sulfur in the ore
concentrate burns to form S02. Pyrometallurgical operations vaporize
several metallic elements including antimony, arsenic, boron, cadmium,
lead, mercury, selenium, tellurium, vanadium, and zinc. The quality of
these emissions varies widely and depends mainly on the ore composition
and the type of smelting operation. The metallic constituents in the
emission stream are characterized by their small particle size; largely
less than one micron. Little quantitative data on particulate composi-
tion as a function of particle size are available. However, since most
of the trace metallic compounds condense from the vapor state, the more
volatile metals tend to form a major portion of the finer particulate.
An illustration of this trend is shown in Table 5-1 for samples col-
lected in a flue serving a copper ore roaster and arsenic plant after
large particulate has settled out. These data are not representative of
other processes since this ore had a high arsenic content, but the trend
to higher metallic content in the fine fraction is evident.
All of the major pyrometallurgical processes release hot waste
gases that contain inorganic particulates, metallic fumes, S02, and
combustion products. The gas streams contain no significant quantities
of organic material, oxides of nitrogen, or carbon monoxide. Smelter
gas streams are characterized as rich or lean in regard to their S02
content with lean streams containing less than about 4 percent S02-
Gases from several processes within a smelter are usually combined prior
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Table 5-1. ELEMENTAL ANALYSIS OF PARTICULATE SIZE FRACTIONS
Average
par ticu late
size, y
14.3
7.5
4.6
2.3
1.3
0.58
0.11
-------
to control. Frequently, the principal exhaust gas stream produced at a
smelter is composed of gases from all the major processes. Other waste
streams, containing fewer particulates and less S02, are usually vented
separately with or without treatment. This practice of combining sev-
eral waste streams allows the use of a few atmospheric emission control
devices; each, however, may be a large piece of equipment.
The combining of waste streams creates a problem in air management,
since it can cause overloading of the control equipment. Most conven-
tional pyrometallurgical processes are batch operations, whose gaseous
emissions are not uniform. Unless the operations are carefully se-
quenced, both the flow rate and pollutant loadings of the waste gases
can be highly variable. Control devices are not usually sized to
handle simultaneously the peak loads from all the processes, and metal-
lurgical operations therefore can result in excessive emission rates.
Smelter emission streams generally have a high opacity due to the forma-
tion of sulfuric acid mist and the fine particulate present.
EMISSION CONTROL SYSTEMS
Particulate Control
Control of waste gas in a smelter is a multi-step process. The
first step is always to capture and cool the gases so that they can be
passed through ductwork and control equipment of standard construction.
Flue gases frequently leave a pyrometallurgical process at 1000°C or
higher. Some of this heat may be recovered in a waste heat boiler,
reducing the gas temperature slightly. Only a limited amount of heat
can be recovered in this way, however, since a smelter does not require
large quantities of steam.
For further reduction of gas temperature, dilution with ambient
air, radiation coolers, and water sprays is used. Smelters are usually
built with very tall stacks, whose principal purpose is to create a
strong natural draft and thus allow a large quantity of cool air to be
drawn into the gas stream. Design of smelter equipment has evolved with
the assumption that tightly sealed ductwork is frequently not desirable.
For example, copper converters have been designed for use with a loosely
fitted hood with the intent that large quantities of air would mix with
the converter gases. Reverberatory furnaces have been designed so that
natural drafts would keep the furnace ,under a slight vacuum to prevent
emissions from the charging and tapping ports.
The second step in control of waste gases is the removal of the
larger particulates. Most of the larger solid particles entrained in
the gas streams are composed of raw and product materials, usually in
amounts too great to discard economically. Larger particles usually
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contain more of the product and fewer impurities than fine particles.
Gravity settling is used to collect these large particulates. Special
balloon flues or settling chambers are installed to reduce gas velocity
enough for their collection. The collected particles are directly
recycled to the process operations. The settling chambers also act to
further cool the gas, since they are not insulated and lose considerable
heat by radiation to the atmosphere.
These first two steps in control of waste gases can be considered
as processing operations since their principal function is not to mini-
mize air pollution but to prevent product loss. The third step involves
removal of fine particulates, with fabric filters and electrostatic
precipitators, with or without additional cooling.
An electrostatic precipitator (ESP) is usually described either as
a "hot" ESP or a cold ESP. The hot version handles the gas at tem-
peratures up to 370°C and collects the particulates as a dry dust. The
cold ESP system utilizes additional gas cooling via radiation or water
sprays to cool the gas stream to about 150°C prior to collection.
Selection of one type rather than the other depends largely on the final
disposition of the gas and plant layout. If the stream is to be used to
feed a sulfuric acid plant, the gas must be scrubbed for more complete
removal of particulates and fumes. If the gas is not processed for
sulfur recovery, cooling is not required and the hot ESP is more common.
Baghouses are also utilized for fine particulate control; especially
at zinc and lead smelters. Baghouse installations usually require
additional cooling of the gas to meet the temperature limit of the
filter fabric (approximately 260°C). To prevent condensation, the
stream is not cooled to its dew point. Particulate collection at higher
temperatures such as in the hot ESP prevents efficient collection of the
more volatile metals since they remain in a gaseous state.
ESP's and fabric filters can achieve collection efficiencies in
excess of 95 percent and with careful design and operation, in excess of
99 percent. ESP's, and to a lesser extent fabric filters, lose effi-
ciency rapidly for particles less than 0.2 microns in diameter.
Particulate collected by an ESP or baghouse is recycled or dis-
carded depending on its value. Many of the trace elements such as
arsenic and cadmium tend to concentrate in the fine particulate frac-
tion. All zinc smelters process the particulate to recover cadmium.
One copper smelter processes the particulate to reclaim part of its
arsenic content.
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S0_2 Control
The final step in treatment of some smelter waste gases is removal
of S02- In this respect, there are only two classes of smelter gas
streams: those that are suitable for sulfur recovery, primarily as
sulfuric acid, and those that are not. Sulfuric acid plants require a
clean gas stream which averages at least 4 percent S02 for efficient
operation. If the gas stream meets this requirement, sulfuric acid
manufacture via the double contact process represents the best currently
available control technology.
Prior to acid manufacture, the gas stream must be conditioned to
remove any residual particulate matter. This is accomplished in a
series of scrubbers and electrostatic precipitators. Particulate re-
moved in these steps forms a sludge which must be disposed of or re-
cycled to the smelting process. The remaining gas stream and the final
acid plant tail gas contain essentially no particulate matter. Acid
plant S02 emissions are less than 500 ppm and about 2000 ppm for a
double and single contact plant respectively.
Marketing of the acid may be a problem in some parts of the country
and neutralization may be required at times.
Sulfuric acid manufacture requires not only a high concentration of
S02, but also a stream that is reasonably constant in both composition
and flow rate. Increased attention to the correct sequencing of batch
operations and instrumentation to prevent sudden changes in flow or
composition has allowed some smelters to more efficiently produce sul-
furic acid.
If the stream averages less than approximately 4 percent SOg (or 2
1/2 percent as the absolute minimum according to some advertised claims),
control via wet scrubbing with a chemical solution could be implemented.
Two or three copper smelters have incorporated a DMA absorptic.. plant.
This process can be used to produce a concentrated stream of S02 for
export or as a buffer between the smelter and the acid plant, to absorb
surges in S02 production and to release S02 to the acid plant during
periods of low production. These scrubbing systems also utilize gas
conditioning systems which remove essentially all particulate matter.
Best air management in a smelter, therefore, consists of modifying
pyrometallurgical processes and waste gas control equipment to produce a
gas stream rich in S02 suitable for acid manufacture. Improvements in
ore concentration have eliminated the need for most multipie-hearth
roasters, in both the copper and zinc industries. Newer processes have
been adopted that use less fuel and yield more concentrated exhaust
streams. Oxygen enrichment has been utilized in a number of smelter
332
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operations and may soon be used more commonly. This practice also
produces more concentrated exhaust streams.
Less fundamental changes have been made at some smelters and could
be used at others to minimize diluting waste gas streams. The changes
are directed toward minimizing infiltration of outside air and substi-
tuting other methods for cooling the gas stream. Copper converters
formerly operated with as much as a 12-inch gap between the converter
and the hood. Substitution of a water-cooled hoods allows reduction of
this gap to a few inches. Replacement of some ducts with new ones made
of high-temperature alloys has minimized the need for infiltration air.
Water spray chambers have been installed to cool the gas stream. At
some plants, holes and leaks in the ductwork have simply been repaired.
Weak S02 bearing streams could be controlled via caustic scrubbers.
In this country, no smelter gases are scrubbed for S02 removal except
for those utilizing the DMA process. In other countries, smelters
employ caustic scrubbing techniques similar to those used in this coun-
try with flue gases from coal-fired boilers. In one Japanese copper
smelter, waste gases are segregated into two streams, one with high $62
content that is sent to an acid plant and one with low S02 content that
is sent to a wet scrubber. The scrubber also treats all process gases
before discharging them to the atmosphere, including acid plant efflu-
ent, fire refining furnace flue gas, and ventilation fan discharges that
collect fugitive S02 emissions. More than 99 percent of the sulfur that
enters with the concentrate is contained.
FUGITIVE EMISSIONS
Fugitive emissions consisting of gaseous and particulate matter
include all atmospheric emissions that are not contained in a vent.
These emissions occur from raw materials handling and sizing operations
storage piles, roadways, and mining areas, slag piles, leaks in process
equipment, inadequate hooding systems, and during furnace charging,
slagging and tapping operations. Even semi-quantitative data on smelter
fugitive emissions are sparse since these emissions are very variable
and depend on many factors including:
0 In-plant maintenance
0 Plant and yard housekeeping practice
0 Type of furnaces
0 Smelter throughput
0 Fume hood systems and draft
0 Care in process operation
0 Railfall and wind patterns
0 Feed composition
333
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The effect of these emissions on ambient air quality may be appreciable
since these emissions are emitted near ground level.
As described earlier, the S02 content of exhaust gas streams is
maximimized by preventing the infiltration of air into process exhaust
ventilation equipment. The elimination of dilution air may, however,
cause increased fugitive emissions of fumes, particulates, and S02
around furnace openings. Thus, efforts to strengthen a gas stream to
enhance pollution control may result in deterioration in the workplace
environment. In Japan, one building enclosing a well-controlled smelter
was exhaust ventilated through a fabric filter for particulate control
to eliminate fugitive particulate emissions, but indoor S02 concentra-
tions were noticeably high.
Fugitive emissions from pyrometallurgical processing equipment can
be greatly reduced through the use of hoods, enclosures and exhaust
systems over charging and tapping operations. These devices can be
vented through particulate control devices and into the furnaces final
discharge stack. Furnace configuration and movement of the ladles to
tap and charge various types of furnaces present design difficulties and
maintenance problems.
Particulate emissions from materials handling, roadways, slag
piles, etc. may be reduced through use of covered conveyors, enclosures,
and storage buildings, spray wetting systems, frequent sweeping and good
plant housekeeping practices.
Technology is available for all phases of fugitive emission con-
trol, but successful application of this technology requires a complete
engineering study rather than a piece-by-piece approach. The most
critical design point is the application of hoods and suction vents to
collect emissions from point sources of fugitive emissions. Experience
has shown that cross-currents of air past a hood can seriously affect
its efficiency; no hood design is complete unless provision is ,nade for
control of all air movements in the vicinity.
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6.0 WATER MANAGEMENT
Pollution of water by an industrial operation arises from three
fundamental causes. Liquid materials may be purposely discharged into
natural waters. These are the direct process wastes, easily quantified
and recognized. Liquids may also accidentally enter a watercourse.
These are the fugitive losses, difficult to quantify or predict.
Finally, solids or gases may be discharged that enter otherwise con-
taminated waters. These are secondary sources, frequently the most
difficult to locate and most troublesome to control.
The primary copper, lead, and zinc industries contribute water
pollution in all three classifications. However, few direct process
wastes are liquids; the major industry wastes are solids and gases.
Each of the few processes that have predictable discharges of clearly-
recognized liquid wastes is discussed in this report in connection with
the process that originates it.
Opportunities for fugitive liquid losses are limited in these
industries. Most of the processes for copper, lead, and zinc production
do not handle water solutions. Only the copper and zinc electrolytic
refineries, and a few auxiliary operations such as sulfuric acid absorp-
tion and storage are likely to produce substantial discharges of toxic
liquids. As with any other liquid-handling industry, control must be
through the provision of facilities to capture and recycle or treat any
accidental losses.
The major threat to water quality from these industries is second-
ary pollution from the vast quantities of minerals that are extracted
from the earth and artifically exposed to the environment. The remain-
der of this report section is a detailed examination of the character
and control of this threat; centering on sulfide ores, recognized as a
major contributor to water pollution.
SOURCES OF SECONDARY WATER POLLUTION
Sulfide ores will contain one or more of the following elements:
iron, nickel, copper, zinc, lead, arsenic, antimony, bismuth, cadmium,
selenium, silver, gold, manganese, and tellurium. In addition, molyb-
denum and rhenium are frequently in some copper ores and germanium is
335
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found in the Missouri lead and zinc ores. The platinum group of metals
is recovered from electrolytic slimes. Indium, gallium, thallium,
cobalt and tin are reported from copper and zinc concentrates. Mercury
and chromium appear in some water analyses from smelting operations.
The common rock-forming elements of calcium, magnesium, aluminum, the
alkali metals and silicon undoubtedly occur closely associated with all
metal ores. Stack particulates from a copper reverberatory furnace were
found to contain boron, barium and vanadium. Thus many potentially
hazardous elements may appear in the wastewaters from these industries.
In the mining, concentrating and smelting of sulfide ores, five
sources of impurities in water can be identified:
1. In the operation of either a surface or an underground mine,
wastewaters will be produced that must be pumped or drained from the
workings. Some of this water will be brought into the mine for equip-
ment cooling, dust control, and as life support for the miners. Some
will be the result of natural seepage, or in the case of a surface mine,
from rain and snow.
2. Spoil is waste rock moved during the mining process as over-
burden, in the construction of shafts and tunnels, or as low grade
inclusions found within the ore body. It is disposed of near the mine
workings. The spoil generally contains appreciable amounts of sulfide
minerals, and may occasionally be rich in unwanted "outrider" deposits
of metals other than the one being produced. With greater or lesser
attention to the environmental effects, these spoils are placed where
they are also exposed to rainfall, seepage, and surface run-off.
3. In mining the main body of the ore, stockpiles of ore are
frequently placed either near the mine or near the mill, or both, in
order to avoid having to synchronize operations of these two processes.
The stockpiles are also exposed to the weather, and in some cases are
purposely hosed down to minimize blowing dust.
4. In one of the major processing operations, the ores from the
mine are very finely ground and circulated through equipment as a slurry
in water. This concentrates minerals containing the desired metals into
a usable fraction. A continuous bleed of this milling and flotation
water is necessary, and the dumps or ponds containing the tailings are
subject to rainfall and seepage.
5. In the smelters, fugitive losses of concentrates and their
decomposition products are common. Fugitive losses of dusts with simi-
lar characteristics are frequent. These materials are subject to rain-
fall, and may also be hosed down in attempts at housekeeping.
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Unless controlled, water from all these sources will follow the
natural drainage of the area, and will enter natural watercourses. All
these waters share the characteristic of having been in contact with
sulfide minerals, and to some degree, all of them will be:
1. More highly mineralized than normal surface waters of the
area, containing especially the sulfate ion.
2. Likely to contain measureable amounts of one or more of the
heavy elements found in the ore body.
3. More acidic than the normal surface waters of the area.
The waste stream that is the largest in volume offers an exception
to the third characteristic. Concentrator effluent has been treated to
be neutral or alkaline. Instead of acidity, it substitutes a very high
suspended solids loading, and also contains various flotation and
decomposition chemicals.
CONCENTRATOR EFFLUENT
The most easily observed source of potential water pollution is the
process water used for ore concentration. The flotation process re-
quires water in large quantity. After being used to separate a fraction
of high grade ore, the water is sluiced into a pond, carrying with it
the waste rock. The volumes of both water and waste rock are very
large. As much as 80 percent of all the ore mined is thrown away as
tailings, and two-thirds or more of the water used by the industry is in
the concentrator plant. Some of the overflow of the tailings pond is
recycled back to the mill or concentrator. Most economical operation of
the flotation process requires a recycle, and in arid regions, economy
frequently dictates the maximum possible percentage. There is, however,
a limit to the amount that can be recycled, since soluble minerals in
the ore, the introduction of organic matter, and chemicals added for
processing all accumulate. If their concentration gets too high, flota-
tion is inefficient.
With the added incentive to minimize water discharges, most flota-
tion mills have approached this maximum limit of recycle. It is un-
likely that further reduction in volume of waste water from the concen-
trator will be possible unless the industry adopts better waste segrega-
tion practices or more complete waste treatment before recycle.
Many of the materials added to the water in the concentrator are of
increasing concern in the management of the wastes. At some milling
operations, high concentrations of partially oxidized sulfur compounds
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such as thiosalts and thionates are produced. Some concentrators also
discharge quantities of cyanide salts and organic compounds. These
materials must be biochemically oxidized if they are not to appear in
the plant effluent. Some of these materials oxidize only with diffi-
culty and will not be decomposed in a simple settling pond.
Various proposals have been advanced on how best to handle and
treat the effluent from the concentrator, but at best these are but
exercises. The fact is that in those industries it is the almost
unanimous practice to combine all waste streams into one. Water from
the mines, acid seepage, storm drainage, plant floor drains, and even
sewage from the plant facilities all are run together, and in most
installations are discharged into the tailings pond. This is in many
cases regrettable, but any proposals to modify the waste management
practices of these industries must start with the consideration that
waste waters now exist as a combined stream.
SULFIDE WEATHERING
The mechanism by which sulfide minerals enter water solution has
been extensively studied, in connection with both sulfide ore mining,
and the sulfide content of the Appalachian coal mines. It is thought
that contamination of drainages with acid and heavy metals starts with
iron pyrite, (Fe$2). This is the most common mineral in all sulfide
ores. When this mineral is in contact with both water and air, it
oxidizes to produce soluble ferrous sulfate and sulfuric acid.
2 FeS2 + 2H20 + 702 -»• 2 FeS04 + 2H2S04
The ferrous iron may oxidize further to form insoluble ferric hydroxide
and still more acid.
4 FeS04 + 02 •»• 4 Fe(OH)3 + -4 H2S04
Other reactions may form a complex sulfate, or jarosite, thus adding
another quantity of acid.
These reactions take place naturally in any outcrop of a sulfide
mineral. They are the natural weathering reactions of pyrite. If the
outcrop is undisturbed, they take place slowly, and the acid is quickly
neutralized by reaction with other minerals or with the natural alka-
linity in surface water. The iron hydroxide remains virtually in place,
changing slowly into the mineral limonite.
Mining operations greatly speed up these reactions, partly by
exposing more sulfides to weathering action, and partly by forming many
small particles from the brittle pyrite crystals. A particle smaller
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than 25 microns oxidizes very rapidly, and soon the natural alkalinity
of the water cannot neutralize the acid as fast as it is produced.2
Free acid appears in the water and the pH drops.
Sulfides of other metals, normally more stable than pyrite, become
more susceptible to oxidation at low pH and they enter into solution
along with more pyrite, contributing still more acid to the water.
When pH drops below 6.5, conditions are now favorable for the
growth of iron bacteria, such as ferrobacillus ferrooxidans,2 also
called "thiobacillus",3 these bacteria obtain their life energy by
oxidizing ferrous iron to ferric, and they act as catalysts to greatly
speed up one of the slower steps of the oxidation process. As water pH
continues to drop, action of the iron bacteria speeds up still more,
until they reach their preferred level of pH 4.3.2 At this degree of
acidity, the sulfides of copper and zinc can be readily oxidized, and
many heavy elements can enter into solution.
All the mining, milling, and run-off sources of wastewater are
affected by this process. The rate of the reaction is influenced by
several factors. The longer the water stands in contact with the min-
erals, the lower the pH becomes and the faster the reactions proceed.
Temperature also affects the rate, and in warm, acid, aerated water, the
speed of dissolution of the minerals is fast enough to see.
Spoil dumps and ore stockpiles are generally considered to be less
important as contributors to water pollution than are mines and tailings
ponds. Stockpiles are usually placed where they will not stay saturated
with water. It is assumed that they generally drain quickly when wetted
by rainfall, and that they will be turned over often enough so condi-
tions favorable to the establishment of iron bacteria will not develop.
These assumptions have been tested primarily with coal stockpiles, and
in reference to the acidity and iron content of the drainage. Tests
specific to the heavy metal content of sulfide ore stockpiles have not
been reported.
Run-off from smelter property has not been well quantified from the
installations in this country. Since many of the minerals in the dusts
on smelter property are partially oxidized by smelter processing, their
rate of solution may be more rapid than from unaltered sulfide minerals.
Spoil dumps are apparently quite variable in their effect on water
quality. It seems to be assumed that old dumps are not important
contributors, since they generally contain quantities of calcium and
aluminum-based minerals that themselves decompose on weathering. They
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act to neutralize the acids produced and to re-precipitate the heavy
metals in new and often complex crystalline structures. There is, over
a period of time, a rearrangement of the elements into a more stable
combination of compounds that tends to resist further weathering. The
extent of this internal purification can be judged by the observation
that whereas fresh spoils contain no sulfates, a well-weathered spoil
may contain as much as 5% calcium sulfate.2
The problem with this analysis is that not all reactions take place
at the same rate, and at some period, even years after deposition, there
may be present quantities of soluble heavy metal salts which have not
recombined into stable molecules. Even the best stability in a body of
oxidized heavy metals is in itself relative. Deposits of oxidized ores
are not found at the surface of the earth except in regions of low
rainfall.
That soluble heavy metals do occur in many spoil dumps has been
clearly demonstrated. Botanical research has shown that plants can be
affected only by soluble salts, and studies have been made on a bent
grass, Agrostis tenuis.4 A strain of this grass, growing on the dumps
of U.S. lead and zinc mines, has developed a high tolerance to zinc and
nickel in solution. European plants have similarly adapted to the
mining spoils of Wales,5 and Australian plants to the spoils of the
Cloncurry district of that country. Few plants other than metal-
resistant strains can live in these soils.^
Tailings are expected to be even more efficient in neutralizing any
acid that is formed, since the other minerals are finely divided and
intimately mixed with the residual sulfides. One test, however, shows
that sulfide oxidation continues. Leached tailings that contained
sulfides were used to grow grasses in a greenhouse environment. At the
end of 2 1/2 years, most samples had dropped in pH from 2 to 3 units,
and were still dropping when the test was discontinued."
WASTE CHARACTERISTICS
Waste waters from the mining and processing of copper, lead and
zinc sulfide ores have been shown in previous sections, to have one or
more of the following characteristics:
1. Highly turbid, containing both settleable and colloidal
insoluble inorganic material. Effluent from the milling and
ore concentration processes is most representative of this
characteristic, as is run-off from smelter property.
2. Strongly acidic, containing free sulfuric acid. Mine drainage
and seepage from sulfide containing tailings or spoil dumps is
representative.
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3. Containing heavy metal ions, frequently toxic metals, in
concentrations higher than is allowable for discharge into
public waters. Most of the wastewater sources of these
industries have this characteristic.
4. Containing high concentrations of metallic and nonmettallic
ions that are not toxic in moderate concentrations. Most of
the waste waters from these industries are high in sulfates.
Concentrator effluent water is frequently high in sodium and
calcium.
5. Containing materials that have a chemical or a biochemical
oxygen demand, some of which may be toxic to animals or
plants. Mill effluent contributes such substances.
To correct the second listed characteristic, excess acidity,
requires that an alkali be added to neutralize the acid. This is the
most fundamental treatment process that is required with these wastes,
and has been characterized here as "pH adjustment".
To correct the fourth listed characteristic, the presence of so-
called "permanent" ions such as sulfate or sodium, requires either a
change of state of the body of water or demoralization. These are
considered as advanced treatment processes, and are discussed in later
paragraphs.
Present technology indicates that there are only two basic methods
that can be directly applied to the other three waste characteristics.
Merely allowing the water to stand for several days or weeks, if the pH
has been properly adjusted, will reduce the oxygen demand, decrease the
heavy metals concentrations, and allow some of the suspended solids to
settle. In these industries, this is accomplished with the tailings
pond, a basic waste treatment system.
The mechanized man-made approach to the same result is the other
basic process. Precipitation of some of the heavy metal ions and their
separation from the water by clarification is a tested and workable
process. It can also be expanded, if needed, to biologically oxidize
organic or reduced waste components.
WASTE HANDLING AND TREATMENT
Oxidation Prevention
Prevention of the formation of harmful waste waters in a sulfide
mining and ore processing operation is in principle very simple. Either
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the mineral can be kept dry, or cxxygen can be excluded to prevent the
ore from weathering.3 In practice, either is difficult.
Techniques of mining cannot keep rain from falling into an open
pit, or keep all water out of an underground mine. It is possible,
however, to design either of these operations so the water that does
enter is pumped out as rapidly as it accumulates. It may be possible to
modify local drainage so the ores do not stand in ponds of water, or to
design an underground mine that will best avoid seepage.
It is possible to provide cover for ore stockpiles, waterproof
storage for concentrates, and to seal spoil and tailings dumps with a
waterproof material. Stabilization of spoil heaps to restore natural
appearance, if done properly, will minimize run-off of polluted waters.
In extreme cases, covering a dump with latex, clay, tar, or plastic,
and then with topsoil and shallow-rooted vegetation, will keep water out
of contact with the minerals.
Excluding oxygen was considered in a feasability study of mining
coal in an oxygen-free atmosphere.9 In the coal industry, such a
procedure would both prevent explosions and prevent pyrite oxidation.
Its feasability in a mine where there is less explosion hazard is
questionable.
It may be possible to exclude oxygen from some worked-out under-
ground mine sections. The use of air-tight seals has been suggested for
tunnels that are to^be abandoned but that must continue to drain into
operating sections. If the air behind the seal were replaced with
nitrogen, sulfide oxidation would cease. Complete flooding of an under-
ground mine section would also stop oxidation. Neither method, however,
would instantly stop pollution of the water due to continued solution of
the sulfite minerals already oxidized.
Sealing a spoi^dump to exclude water, as described above, will
also exclude oxygen. Heavy compaction of some types of fine-grained
waste rocks may also prevent air from pentrating far below the surface.
This compaction may occur naturally in some deposits of tailings.
The addition of oxygen-scavenging chemicals, such as calcium sul-
fite, may minimize the oxidation of the sulfides of heavy metals for a
short period of time. Sulfites may therefore assist in temporary sta-
bilization until more permanent treatment can be accomplished.
The use of sewage sludge or other decaying organic material has
also been suggested as a means of scavenging the oxygen from spoil or
tailings, thus minimizing the rate of sulfite oxidation.3
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The role of the iron bacteria in increasing the rate of formation
of acids and soluble metals has been studied in recent years.10 These
bacteria are controlled in other industries by chlorination or ozona-
tion,11 but this procedure may be self-defeating when used with sulfide
ores. Antibiotics have been examined for control in mine waters. Three
of 15 tested were effective.12 An attempt to develop a virus that would
infect the bacteria was apparently unsuccessful, but a reduction in the
rate of sulfide oxidation was obtained when a slower-acting species of
iron bacteria (caulobactous) was purposely introducted into a mine.12
One report points out that there must be some undiscovered mecha-
nism, other than iron bacteria, to account for the speed of oxidation
observed in some sulfite localities.2
One technique that does not work is simply sealing a mine, unless
the mine can be guarded in perpetuity. Literature cites serious inci-
dents from this source. For example, one lead mine in Great Britan,
sealed for about 50 years, was accidentally opened by a bulldozer. A
large quantity of water ran out into a river, and the analysis indicated
140 mg/1 lead and 230 mg/1 zinc.!3 if exposed to air and water, there
seems to be no limit to the degree of weathering that can take place.
Wastewater Collection
If the assumption is made that any sulfide mine, spoil dump, tail-
ings bed, smelter property, or concentrator plant is a potential source
for water pollution, collection and treatment of all those waters is a
major effort. In some areas, all operating units may be concentrated
into a single reasonably small drainage area, but in other places,
possible sources may be widely distributed and discharge into several
different watersheds. In an old, heavily worked area, the identifica-
tion of all the possible sources and assignment of responsibility for
them may be very difficult. No general rules can be listed that will
assist in this task, other than the well-known techniques of following
the ecological clues that nature provides.
If effluent quality from existing operating installations is to be
improved, however, the copper, lead and zinc industries must at some
stage attend to waste classification. There is, for example, no benefit
in the practice of including sewage from operating facilities in process
wastes. It is recognized that the organic molecules thus introduced can
seriously limit re-use of the water for ore concentration. Conditions
for best biochemical oxidation are not compatible with the conditions
for best removal of heavy metals. The volume of sewage discharged by
even a large plant can normally be handled by a treatment as simple as a
septic tank or a similar unattended device, so the major cost in segre-
gation of this waste is installation of plumbing rather than the treat-
ment plant itself.
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If cooling towers or steam boilers are part of the mine, smelter,
or concentrator installation, blowdown from these sources also intro-
duces chemicals not compatible with either metals removal or biochemical
oxidation. Sequestering agents used in boiler treatment are designed
specifically to keep metals from precipitating. Algicides in cooling
towers act specifically to limit biological activity. Treatment of
undiluted blowdown waters is difficult, but if dilution were made down-
stream of the main treatment facility, better overall effluent quality
might be obtained.
Highly acidic wastes, or wastes known to contain high metals con-
centrations, can be treated more effectively if they remain concen-
trated. If mine waters were passed through a bed of coarse limestone
and/or scrap iron prior to mixing with other wastes, heavy metals con-
centration would probably be reduced and less lime would be required for
pH adjustment. Tests to confirm the effectiveness of such simple treat-
ments on specific streams are easy and inexpensive to perform, and could
very well decrease loadings in the main waste stream.
Separate treatment of cyanide wastes is virtually the only method
for preventing discharge of these poisonous ions, since cyanides are
stable in alkaline solutions. Mixing of cyanide wastes includes a
second hazard, in that if acid streams are intermixed prior to lime
treatment, poisonous HCN gas can evolve very rapidly from the mixture.
The Tailings Pond - Primary Treatment
It has been the practice in the copper, lead and zinc industries to
combine most water wastes into one stream to be sent to a tailings pond.
Usually chemicals have been added to remove acidity from the water. If
effluent quality has been inadequate, in many cases the next step has
been to add another pond in series with the first. If this failed,
another pond might be added. Using this theory of waste treatment,
ponds have been strung together for up to eight km before the ,;ater
finally was discharged from company property.
There are many advocates of pond treatment of wastes. Ponds seem
to be the only practical means for disposing of the large tonnages of
waste rock.'4 The large surface area speeds up oxidation processes and
breaks down many sulfides, sulfites, and organic materials. The reten-
tion time allows most of the suspended solids to settle. In many cases,
they can be quite efficient and can provide a sizable reduction in
mineral constituents.
Combining highly metalliferous wastes with the finely divided
tailings allows the same sort of internal purification process to pro-
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ceed in a tailings pond that has been previously described for spoil
dumps. These chemical changes are extremely complex, involving not only
chemical reactions but also intercrystalline substitutions. However,
the rearrangements of the elements to approch the most stable oxidized
configurations occur naturally, requiring no sophisticated analyses, no
precise control, and no complex equipment.
It is the apparent simplicity of the pond system that has caused it
to be adopted so widely. In an isolated mining area, it is frequently
the least expensive treatment system, requiring no utilities or energy
and needing no highly technical staff of people for its operation. In
addition, the results obtained by some tailings pond systems are impres-
sive. One published report states that a properly designed and con-
structed pond, carefully operated to maintain a pH between 9.5 and 10.5,
can consistently produce an effluent with no more than the following
metals concentrations:!4
Copper 30 ug/1 (extractable)
Zinc 150 ug/1 (extractable)
Lead 100 ug/1 (extractable)
Total iron 1 yg/1
These levels are lower than reported pilot plant studies using coagula-
tion, clarification and filtration. Other reports state that best
removal is at a higher pH - up to pH 11.0.15
Despite these results, a tailings pond has its disadvantages. It
is not possible to use sophisticated analyses or precise control because
there is no way to collect a representative sample of in-process waste-
water or to optimize the performance of the system. Any treatment will
be trial and error, and an error cannot be recognized until, days later,
the effluent analysis indicates inadequate quality. Any change in
treatment will then require several days before an improvement in efflu-
ent quality. Meanwhile, the pond is subject to continuous variation due
to changes in temperature, rainfall, and wind. The phenomenon of "ther-
mal skimming" may allow warmer waste water to slide rapidly across
colder pond water. It is therefore difficult to assess the results of
different treatment strategies.
A single pond, or a string of ponds, cannot easily perform all the
functions required in the treatment of these mixed wastes. Biochemical
oxidation of components such as thiosalts, ammonia, and organic mole-
cules is aided by turbulence and works better at neutral or slightly
acid pH. Removal of heavy metals requires precise control of pH at an
alkaline level and requires quiescent conditions for sedimentation,^
whereas in a pond, degree of turbulence depends on the wind.
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There is very limited information on variations in pH in a pond
treatment system. The usual tailings pond pH control consists of lime
additions at a manually set rate, periodically readjusted by the opera-
tor on the basis of grab samples. This degree of precision would be
inadequate for some chemical processing operations, but it has seemed
appropriate for ore processing since none of the chemical control in the
ore flotation process itself requires anymore precision than this.
The tailings pond system is firmly entrenched, however, and there
is no basis on which to recommend its abandonment. The most likely
combination of effluent treatments will continue to use the tailings
pond, with lime added to maintain an approximate pH. A secondary treat-
ment plant, described in the following paragraphs, would accept whatever
water flows out of the pond. If pond effluent is acceptable, the
secondary plant would serve only as a wide spot in the effluent line.
If quality becomes unacceptable, the plant would be placed in operation,
using a more scientific approach to remove the remaining pollutants.
Precipitation and Clarification - Secondary Treatment
If plant effluent is not acceptably free of heavy metal ions, the
best known process for reducing their concentration is chemical preci-
pitation. The soluble ions are converted to insoluble compounds by
chemical mechanisms that are complex and variable. The solids are then
separated from the water. This process is routinely used to remove iron
and manganese from public water supplies, and the techniques are well
known and proven in hundreds of installations.
It is known that the following metals can be successfully precipi-
tated: copper, Jead, zinc, cadmium, nickel, arsenic, chromium, manga-
nese, and iron.lb In theory, each metal precipitates to its lowest
concentration at a definite pH. Cadmium and manganese require a pH near
10, while zinc is least soluble near a pH of 9. Above or below this
critical pH, more of the metal will remain in solution. It is also
theoretically impossible to precipitate all of the ions from solution.
In practice, there has been confirmation that optimum pH levels do
in fact exist, and that a change of as-little as 0.2 pH can affect
metals concentration of the effluent.17 Practice has also shown that
interactions between the ions and with other substances in the water can
greatly influence the degree of removal. In practice, the optimum pH is
also often found to be higher than that predicted by simplified theory.
At the same time that an alkaline chemical is added for pH adjust-
ment, flocculants are added to coagulate or agglomerate the precipitated
metal compounds and the other suspended solids in the water. Floccu-
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lants form sticky or gelatinous precipitates when mixed with water; they
collect and bind together small particles and form agglomerates that are
heavy enough to settle upon standing.
The flocculants traditionally used are inorganic. They are the
salts of aluminum or ferric iron, or are ferrous iron salts added along
with an oxidant (chlorinated copperas). Being metals, they also have a
rather small range over which they form the gelatinous hydroxide "floe".
Aluminum does best at pH less than 8.5, iron at pH 9 to 10.
In recent years there has been an increased use of organic floccu-
lants. These are usually proprietary mixtures, classifiable as poly-
electrolytes or polymeric materials, and are used either alone or in
combination with the metal salts. They are usually more effective then
inorganics. Although they are more expensive, they reduce the load on
filters and frequently result in lower overall operating costs. In
waste treatment applications, they may make filtration unnecessary.
The addition of adsorbents or fillers during coagulation is oc-
casionally practiced, usually either to add weight to the floe to in-
crease settling or to adsorb organic materials. It has been observed
that occasionally such additives increase the degree of removal of metal
ions. The mechanism of this action is unclear but it may explain the
unexpectedly low concentrations of metals in some tailings pond overflow
waters. One test of ferrous acid mine drainage with activated carbon
has been reported J8
The equipment used for this clarification process can be an engi-
neered and locally-constructed assembly, but is most often a commer-
cially packaged clarifier. With either type, there must be three dis-
tinct treatment sections:
1. A flash mix section, in which all the chemical additives are
quickly and thoroughly mixed with the water.
2. A flocculating section, in which the water is subjected to a
slow, rolling agitation. This causes the floe to capture the
fine solids and grow to large size.
3. A sludge concentration section, in which the solids are
separated from the water.
A number of proven designs are available that incorporate all these
requirements. Energy must be provided for flash mixing and floccula-
tion, and in the treatment of a large flow of water the process may
require a sizable amount of energy.
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The sludge concentration section operates on either of two prin-
ciples. The simplest is gravity settling, in which the water is allowed
to move slowly through a large basin or pond while the solids settle to
the bottom. This method is not entirely satisfactory, since the sludge
cannot be continuously removed. It remains in contact with clarified
water, and re-solution of the precipitate can occur. Simple settling
has been largely replaced by up-flow devices in which the solids settle
by gravity against an upward flow of water. The solids accumulate in a
blanket at some level in the clarifier, from which they are continuously
purged. Part of the sludge may be returned to the flash mix section,
increasing flocculant loading and thus collecting more small particles
of precipitated heavy metals and other suspended solids.
A recent pilot plant study used a polymer flocculant in the clari-
fication of sulfide mine drainage. The concentrations of metals in the
effluent were:19
mg/1, extractable
Average Minimum Maximum
Copper 0.05 0.03 0.11
Zinc 0.37 0.13 0.90
Lead 0.25 0.01 0.62
Iron 0.28 0.09 0.60
These apparently high concentrations do not reflect the degree to
which the metals were precipitated by the clarification process. Sand
filtration removed more than 60 percent of the lead and zinc, down to an
average of 0.09 and 0.15 mg/1, respectively. On the other hand, most of
the copper passed through the filter as a soluble ion. Its average
concentration was 0.04 mg/1 after filtration.
Various proposals have been advanced for disposition of sludge from
a heavy metals clarifier. Many consider sludge disposal to be one of
the most problematic and potentially expensive problems to be over-
come. '9 Returning it to the tailings pond may upset the pond's complex
chemistry, especially if organic flocculants have been used. U could
be discharged to its own pond for storage in perpetuity, but any over-
flow would undoubtedly be highly mineralized and would itself require
treatment. It could be de-watered and dried with mechanical filtration
and drying equipment and disposed of as landfill, but secondary seepage
could be a serious problem because of its high mineralization. Experi-
ments with stabilizing agents (Dravo, Calcilox, etc.) such as are now
being tested with power plant fly ash have not been reported. Drying
and incineration of the sludge may also be feasable, with the resulting
ash recycled to a pyrometallurgical process for recovery of valuable
metals.
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Filtration
It is known that clarification alone cannot remove all suspended
solids from water. Some of the small particles will not become enmeshed
in a particle of floe, and some of the smaller floe particles will be
carried out with the water. As mentioned previously, a pilot plant test
showed that more than 60 percent of the lead and zinc found in a clar-
ifier effluent was present as insoluble particles, capable of being
removed by filtration.^
If filtration is necessary, the equipment that is most suitable for
handling large volumes of water is the "rapid sand filter". The water
flows downward through a bed of smooth quartz sand or anthracite coal.
Larger particles of carry-over floe are caught in the spaces between the
sand grains, filtering out the remaining solids from the water. The
effluent is clear and sparkling and usually completely free of suspended
solids.
Periodically the filter must be backwashed by pumping a portion of
the clear water back through the sand in reverse direction. The dirty
stream of filter backwash water will contain all the suspended solids
not previously removed by the clarifier. The filter backwash is more
dilute than sludge from the clarifier. It is also produced in large
intermittent slugs. The most usual method of disposal is an agitated
tank which catches and feeds it back into the clarifier at a constant
slow rate. There it is re-concentrated; the solids leave as part of the
clarifier sludge.
Sand filters represent a large percentage of the capital cost of a
water treatment facility, and they account for much of the operating
labor.
\
It is not possible to use sand filtration as a substitute for
flocculation and clarification. In itself, sand is not a good filter
media. The actual filtration is accomplished by the thin layer of floe
that forms on the surface of the sand. Without flocculation, sand
filter effluent would be dirty. The filter would also blind rapidly,
and be difficult to backwash, with the result that it might not filter
much more water than is needed to backwash it.
Diatomaceous earth filters could be used for filtration without
prior clarification. These filters use a layer of a natural cellular
silica as the filter media. More of this diatomaceous earth may also be
fed into the water as it passes into the filter. When the filter is
backwashed, no attempt is made to recover the diatomaceous earth and it
is discarded along with the solids from the water. This type of filtra-
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tion is widely used in the clarification of food and chemical solutions
and in small water systems such as swimming pools. The operating costs
and sludge disposal problems in an application to a major waste stream,
however, do not recommend it.
. Most other types of mechanical filters used throughout the process
industries are much too small for application to fine filtering of a
large water stream. Direct filtration of a special type has been
evaluated by at least one installation in another mining industry (tar
sands), and may some day be directly used to replace the tailings pond.
The preceding description of a sand filter describes the "rapid"
type as it has developed in the last 50 years. Recently, as problems of
sludge disposal have increased, there has been a revival of interest in
the older "slow" sand filter, once used for clarification of public
water supplies. These are large basins, fitted with a network of
perforated drainage pipes and covered with layers of graded gravel and
a top layer of filter sand. Water passes through this filter at a much
slower rate than through the rapid filter, and solid particles are
caught by the same mechanism.
These filters are not backwashed. As they accumulate more and more
solids, they finally cannot handle a reasonable quantity of water. They
are then allowed to dry, and the accumulated materials are removed
mechanically. It might also be necessary to discard the sand, gravel,
and underdrain piping, or in extreme cases, to abandon the entire filter.
The slow sand filter operates efficiently for only a short time
after it is rebuilt. As solids accumulate, vertical channels develop
and much of the water flows, unfiltered through these channels. Algae
causes the filters to plug and bacterial infestations of many kinds may
cause problems.
Slow sand filters are no longer used for public water supplies in
this country, but their principle may find application in waste treat-
ment or in sludge disposal.
Post-Treatment
Most effective removal of heavy metal ions will probably require
precipitation at a pH that is more alkaline than the allowable level for
discharge to a lake or stream. Re-acidification may be required in
order to meet these environmental standards.
The conventional method to acidify water, and until recent years
the least expensive, is submerged combustion of natural gas or propane.
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The combustion gases, containing carbon dioxide, bubble up through the
water, forming carbonic acid. This is a weak acid, easily controlled.
There is no danager of over-acidification since the water can absorb
only a limited quantity of the gas. The carbon dioxide introduces no
permanent negative ion, and the bicarbonate ion is environmentally
compatible with most ecosystems. The rising bubbles from the submerged
combustion effectively agitate the water and provide good mixing so a
water of uniform pH is discharged. However, the availability and cost
of natural gas and propane may force the use of mineral acids which have
none of these advantages for this re-acidification. Sulfuric acid will
undoubtedly be employed in the copper, lead and zinc industries. Al-
though the amount used will be very small, a mechanically agitated
vessel for the final pH adjustment may be required.
Canadian researchers believe that water adequately treated for
discharge suffers a drop in pH after release due to the presence of many
different incompletely oxidized species. They therefore maintain that a
slight excess of pH is of no great consequence, and that purposeful
reacidification may be less desirable than releasing the water at higher
pHj9 if this drop in pH also applies to U.S. waste waters, extended
storage in a special pond prior to release may be the best overall
ecological solution.
Basic Treatment Effectiveness
The basic treatment methods outlined in the previous sections can
never be totally effective. Given enough time, many of the heavy metal
compounds exposed to the moist, oxygenated environment of the earth's
surface will become soluble. Eventually, the toxic metals artifically
exposed by the mining of copper, lead, and zinc ores will flow to the
ocean. The best control that can be expected is a slowing of down their
rate of solution, concentration and containment, and the prevention of
gross discharges.
Pond treatment, coagulation, clarification, and filtration are the
only proven and practical methods for the concentration of large volumes
of wastes containing dilute quantities of metallic ions. Reports show
that these treatments are capable of removing copper to less than 0.5
mg/1, lead to less than 1.0 mg/1, and zinc to less than 10 mg/1.
Theoretically, copper could be reduced to 1 to 8 micrograms/1, lead to 1
microgram/1, and zinc to 10-60 micrograms/1. As mentioned previously,
pilot plant tests20 have been performed that are consistent with these
levels and are comparable to tertiary treatment of domestic sewage.
Other references show slightly lower concentrations, down to 0.10 mg/1
for lead and 0.15 mg/1 for zinc.16
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There is much less definitive information regarding the effective-
ness of this treatment method in removing other elements. It is known
that such processing should theoretically be effective for any heavy
metal that can form in a low energy compound as a positive ion. This
would include nickel, cadmium, bismuth, manganese and chromium. In
Canada, it has been reported that 0.04 mg/1 is believed to be an attain-
able level for cadmiumJ5 This same report indicates that arsenic can
be consistently reduced to 0.05 mg/1. There is, however, little oper-
ating data or theoretical evaluation of the effectiveness of basic
treatment in reducing the concentrations of arsenic, antimony, selenium,
and tellurium, elements which can form low energy compounds as negative
ions (arsenates, selenites, etc.).
ADVANCED TREATMENT
Present technology indicates that there are no simpler or less
costly techniques that could replace tailings ponds, coagulation, clar-
ification and filtration. Many of the methods proposed for this indus-
try, or being used in other industries, are not applicable to large
volumes of water containing large amounts of suspended solids, high
concentrations of non-hazardous dissolved solids, and much smaller
concentrations of hazardous ions.
The techniques that are described in the following paragraphs are
therefore applicable either in conjunction with or following basic
treatment, or else applicable only to small, selected streams. In many
cases, such as reverse osmosis and ion exchange, a clear water, free
from suspended solids, is necessary. In other cases, cost may be pro-
hibitive for large volumes.
The treatment techniques can be roughly classified as either
physical, chemical, or biochemical in nature. Each will be discussed
separately.
Physical Processes
Change of state is one physical process that has been considered
for treatment of wastes containing large quantities of dissolved impuri-
ties. If water is evaporated, the minerals remain as a sludge or scale
or as a more concentrated solution. If the water is frozen, the ice
that forms is relatively pure and the minerals are concentrated in the
smaller quantity of liquid that is left un-frozen.
Freezing is more applicable to these wastes than evaporation, since
calcium salts form a tough insulating scale if they are concentrated by
evaporation. A study has been made of the purification of mine waters
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20
by freezing. The cost of this method is likely to be quite high, and
the residual salts present a substantial waste disposal problem.
Reverse osmosis is another system that can physically concentrate
soluble impurities into a smaller, more concentrated stream. This was
also tested on waters from mines in 1971.21 The capacity of commercial
reverse osmosis units are very small, and disposal of the concentrated
salts is also a problem with this method.
Chemical Processes
A variety of chemical processes have been proposed for wastes such
as those found in copper, lead, and zinc processing. One method sug-
gested is electrochemical. It is proposed that the wastes be sent
through electrowinning cells to deposit the heavy metals at the cathode.
No data exists on this method of waste treatment, but it is likely that
the metals concentrations are well below the practical technology for
electrolytic devices. The metals would also probably be formed as a
slime, and would likely require secondary coagulation and filtration for
removal.
Chemical absorption methods have been proposed. A few chelating
chemicals are known that react quantitatively with selected metallic
ions. The chelates, plus unreacted chelating chemicals, are then
extracted from the solution. The chelates are broken, the metals are
salvaged, and the chelating materials regenerated to be recycled. This
technique is being used to extract copper from solution in some proc-
essing operations. Several developing hydrometallurgical processes in
the copper industry may use this -extraction technique. For treatment of
wastes, however, the chelating chemicals have the disadvantage of being
specific to only one or two ions, usually copper and iron. Published
literature indicates none that react with most other metals. This
procedure may become important, if the demand exists for the development
of the required chemical materials.
The addition of sodium sulfide to waste waters has been proposed as
a method for the removal of heavy metal ions from a waste stream.'" The
solubilities of the sulfides of the metals is much less than their
carbonates or hydroxides. The soluble sulfides would be added along
with the flocculating chemicals in a clarification installation, and the
metals would be removed with clarifier sludge.
The efficiency of this system in removing heavy metals is appar-
ently not known from a full-scale test. Its efficiency must be bal-
anced, however, against the other problems that sulfide addition would
raise. In general, any waste treatment process must be based on the
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assumption that any harmful components in the waste will be oxidized, in
order that they may exist as stable substances in the oxygenated en-
vironment of the earth's surface. Sulfides are strongly reduced mate-
rials and are not stable in contact with oxygen. As a result, before
soluble sulfides can effect removal of metals, they would have to be
added in sufficient quantity to completely de-oxygenate the water. Any
soluble residual sulfides would have a large chemical oxygen demand and
would be toxic to aquatic life, thus the water would require treatment
before it could be discharged. Sulfide treatment also includes the
possibility of releasing toxic and malodorous hydrogen sulfide gas into
the atmosphere. In addition, the strongly reduced sludge that would be
produced by this operation would be very unstable since it would contain
many finely divided particles of sulfide minerals, not organized into
crystalline lattices, and the sludge would weather rapidly to form
acidic, soluble heavy metals.
The use of soluble sulfides should therefore not be considered
unless adequate facilities are available for re-oxidation of both the
effluent water and the sludge. With these facilities, very low effluent
concentrations of metals might be realized.
Two classes of chemical treatments are in use that remove metal
ions from water by substituting other more desirable ions. One, usually
called "cementation", uses an active elemental metal to precipitate the
less active metals. This process has been studied largely in connection
with the more concentrated solutions encountered in the processing of
lead or copper ores. Iron is the metal most often used in process
operations, but zinc metal has also been tested. Under optimum condi-
tions, it has been shown that iron shot could reduce a copper solution
down to less than 1 mg/1 in 10 minutes or less,22 but these optimum
conditions included very low pH, absence of oxygen, and strong agita-
tion. Even if conditions are not optimum, however, reduction of metals
concentration is significant, especially in acidic streams. The process
is in use with some mine drainage streams, and its application could be
increased if guidelines for its use were available. A special applica-
tion of cementation has been studied, in which zinc dust was used to
remove selenium from a smelter scrubber effluent.23 Similar applica-
tions may be applicable to wastes from mining and concentrating.
The second metal substitution process is ion exchange. This is
considered by some authors as the standard method for concentrating
heavy metal ions,24 and was extensively tested in 1972 for use on acidic
mine drainage. However, when applied to the bulk of the waters from
mining and concentrating operations, the usual commercial ion exchange
resins have a disadvantage. Although they can remove heavy metals, they
also remove calcium and magnesium, and sometimes even sodium and potas-
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slum. These elements exist in waste waters from mining and milling at
concentrations often thousands of times greater than the hazardous
metals. It is therefore necessary to use very large ion exchangers,
removing many more materials from the water than is necessary, in order
to insure that all the hazardous ions are removed.
Ion exchange has another disadvantage. When the resin is regen-
erated, the heavy metals reappear as a solution mixed with large amounts
of regenerating chemical, usually a sodium chloride brine, this pre-
sents a large secondary disposal problem, for which at this time there
is no satisfactory solution.
There has been speculation, however, that some of the manufacturers
of ion exchange resins were attempting to develop a calcium or magnesium
based resin. Whether such a resin would work in the presence of sul-
fates and carbonates is not known, but a magnesium-based resin would
probably not be affected by sodium, potassium, or calcium, and therefore
would be reactive only to those elements less active than magnesium.
Such a resin would be ideal for treatment of clarified and filtered
wastes containing heavy metals.
Biochemical Processes
Biological precipitation of heavy metals has been proposed, but no
references to recent research have been found. It is known that some
strains of microorganisms have the ability to actively absorb heavy
metals through their cell wall. This seems to occur quite rapidly, and
if the strain of the organism is resistant to the metals, they can then
be filtered from the water and concentrated, along with the metals, into
a smaller volume. A problem noted in some of the initial research was
that the microorganisms showed a tendency to be easily replaced by a
strain that merely rejected the metals.
Selenium occurs naturally in the soils of several of the western
states, and a group of plants has evolved on these soils that accumulate
selenium in their tissues. They are well known by the ranchers of the
west, who call them collectively "poison vetch" because they cause
selenium poisoning in the livestock that consume them. They consist of
about twenty species of the genus Astralagus, and some can concentrate
this element a thousandfold or more, reaching selenium concentrations up
to 15,000 mg/1.25 When they die, these plants may return so much sele-
nium to the soil that other plants growing on that spot can become
toxic.
There has apparently been no research to establish whether or not
this ability can be of benefit in waste treatment. All of the plants
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are adapted to highly mineralized and poor soils. It is not beyond
possibility that they could thrive on seleniferous tailings, or be
irrigated with seleniferous waste waters. It is also possible that the
plants could be harvested to produce enough selenium to make this
operation profitable, especially if "high yield" strains of these plants
were developed. It is also a possibility that if the mechanism by which
the element is accumulated by the plants were studied, it -might open a
new route for the artificial concentration of this element.
Selenium is the only element known to be actively accumulated by
plants in the United States. Other elements, however, are passively
absorbed. A few food crops become poisonous if they are grown on soils
high in molybdenum. A grass, Agrostis tenuis, mentioned earlier, can
accumulate high concentrations of zinc in its tissues. Most research in
the relationship of plants to metallic elements has been directed toward
the plants that do not absorb metals, and thus can be grown for food in
the presence of hazardous ions. Much less attention has been directed
toward those that do absorb metals, thereby accumulating them for safe
disposal.
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REFERENCES
1. E.E. Smith, et.al., "Sulfide to Sulfate Reaction Mechanism. U.S.
Dept. of Interior, Federal Water Quality Administration. FPS 02-
70.
2. H.G. Glover, "Acidic and Ferruginous Mine Drainages". The Ecology
of Resource Degradation and Renewal, 1975. Blackwell Scientific
Publications, London.
3. Mine and Mill Wastewater Treatment, Canadian Environmental Protec-
tion Service, Report EPS 3-WP-75-5.
4. Ian Tribe The Plant Kingdom, Grossett & Dunlap, New York.
5. S.J. Wainwright, et. al. "Physiological Mechanisms of Heavy Metal
Tolerance in Plants". The Ecology of Resource Degredation and
Renewal, Blackwell Scientific Publications, London.
6. W.A. Berg, E.M. Barrau, and L.A. Rhodes. Plant Growth on Acid Mill
Tailings Colorado State University, Fort Collins, Colo.
7. Use of Latex as a Soil Sealant to Control Acid Mine Drainage. EPA,
Wat. Poll. Control Res. Ser. EFK 06-72.
8. D.A. Hall. "The Sealing of Coal Dumps with Road Tar". J. Inst. of
Fuel 40, 474-6.
9. Feasibility Study of Mining Coal in an Oxygen-Free Atmosphere.
U.S. Dept. Inter., Water Pollution Control Res. Serv. DZM 08-70.
10. C.M. Lau, et. al. "The Role of Bacteria in Pyrite Oxidation Kinet-
ics". Proc. 3rd Symp. on Coal Mine Drainage Research". Pitts-
burgh, Pa.
11. Beller M., Waide C. & Steinberg M. (1970) Treatment of Acid Mine
Drainage by Ozone Oxidation. U.S. EPA, Water Pollution Control
Res. Ser. FMH 12-70.
12. R.E. Shearer, W.A. Everson & J.W. Mausteller, 1968 & 1970, 2nd and
3rd Symposium on Coal Mine Drainage Research, Pittsburgh.
357
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13. A.N. Jones & W.R. Haura!! "The partial Recovery of the River
Rheidol". The Ecology of "Resource Degredation and Renewal". 1975.
Blackwell Scientific Publications, London.
14. A.V. Bell "The Tailings Pond as a Waste Treatment System". Cana-
dian Min. & Metal!. Bull., April 1974.
15. Base Metal Mine Waste Management in Northeastern New Brunswick.
Canadian Environmental Protection Service, Report No. EP S8-WP-73-
1, 1973.
16. Development Document for Interim final and Proposed Effluent Limi-
tations Guidelines and New Source Performance Standards for the Ore
Mining and Dressing Industry. Point Source Category Vol. 1. EPA
440/1-75/061. Effluent Guidelines Division Office of Water and
Hazardous Materials, U.S. Environmental Protection Agency. Wash-
ington, D.C. October 1975.
17. Unpublished, Operating1 data, Water Department, Batesville, Ar-
kansas.
18. Ford T. & Boyer J.F. (1973) Treatment of ferrous acid mine drainage
with activated carbon. U.S. Environ. Prot. Agency, Eviron. Pro-
tect. Technol. Ser. EPA-R2-73-150.
19. A.V. Bell, et. al. "Some Recent Experiences in the Treatment of
Acidic Metal-Bearing Mine Drainages", Canadian Min. & Metal!.
Bui!., December, 1975.
20. Purification of mine water by freezing. EPA, Water Poll. Control
Res. Sen/., 1971. DRZ 02-71.
21. Acid mine waste treatment using reverse osmosis, EPA, Water Poll.
Control Res. Serv. , 1971, DYG 08-71.
22. O.P. Case. Metallic Recovery from Waste Waters Using Cementation.
EPA, Office of Res. & Development. 1974.
23. W.N. Marchant, R.O. Dannenberg & P.T. Brooks. Selenium Removal
from Acidic Waste Water Using Zinc Reduction and Lime Neutraliza-
tion. U.S. Dept. Interior. Salt Lake City Metal!. Res. Center,
1975.
24. H.F. Lund. Industrial Pollution Control Handbook. 1971. McGraw-
Hill, New York.
25. J.M. Kingsbury. Deadly Harvest. Holt Rinehart 7 Winston, New
York.
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7.0 EMERGING TECHNOLOGY
Considerable research is being conducted, both in the U.S. and in
many other metals-producing countries, involving the primary copper,
lead, and zinc industries. The goals of this research are more efficient
and economical operations that will result in fewer environmental
problems. The greatest amount of emerging technology is for the pro-
duction of copper. To some extent, the lead and zinc industries have
preceded copper in adopting new procedures or building new plants. In
addition, recent economic events have discouraged further developments
in these industries while at the same time emphasizing the need for
changes in copper processing. Process development has taken several
directions. Although already in widespread use in the copper and zinc
industries, the technology of electrolysis continues to be improved.
There are a number of new continuous pyrometallurgical technologies that
offer great promise in terms of improved economics and reduced environ-
mental impacts. There are several new hydrometallurgical processes at
the laboratory scale that have permitted recovery of metals without the
use of high temperature reduction. Unique technologies are being de-
veloped that for the first time permit the separation of metals from
certain raw materials or that eliminate steps usually taken in conven-
tional processing. There are several possible methods of treating slags
from continuous pyrometallurgical processes. This section of the report
will outline the most promising of these developments. Coverage of the
various processes varies considerably; many of the new technologies are
proprietary, and little information has been released. In each case, a
brief description of the process and the status of the technology is
given. Where available, data on energy requirements, environmental
residuals, and costs are also presented.
Electrolytic Processing
Copper and zinc refining in the U.S. is most often accomplished by
electrolytic methods. The technology continues, to develop, however,
particularly in other countries. Foreign smelters make wider use of
electrolytic principles and employ different auxiliary techniques than
do U.S. smelters.
Electrolytic Refining of Lead - In a process sometimes known as
the Betts process, many foreign lead smelters eliminate the pyrometal-
lurgical purification steps used in the U.S. These steps are thought to
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be major sources of pollution by trace elements. Electrolytic refining
permits direct recovery of bismuth and other elements from electrolytic
slimes.
Process Description - After dressing to remove copper and tin, lead
bullion is cast into anodes. The anodes are placed in concrete electro-
lytic cells, usually lined with asphalt or polyethylene. The electro-
lyte is a solution of lead fluosilicate and fluosilicic acid (^SiFe),
and typically analyzes lead, 67 g/1; free H2SJF6, 95 g/1; and total
acid, 142 g/1. Electrolyte temperature is maintained at about 40°C.
Energy Requirements - Current density is 160 to 190 amp/m? and
cell voltage ranges from 0.3 to 0.7 volt. Current efficiency is 90 to
95 percent.
Environmental Residuals - Electrolytic refining of lead accumulates
the trace elements normally released by pyrometallurgical purification
into a slime, from which they may be isolated by chemical procedures.
Status of Technology - Many foreign smelters practice electrolytic
refining of lead.
Direct Electrolysis or Lead from Galena? - Australian research has
indicated that it is possible to extract lead directly from galena
concentrates by electrolysis. This development would replace the con-
ventional lead smelter.
Process Description - Galena concentrate is first compacted with
5.5 percent graphite to provide good conductivity and mechanical strength.
The electrolyte is a perchlorate. A pure lead cathode is used to start
deposition, and it is reported that 75 percent of the lead in the con-
centrate is extracted. Current efficiency is 85 percent, a considerably
higher figure than is found in conventional electrolysis.
Energy Requirements - Total energy requirements are about c600 kg-
cal per kg of lead produced.
Environmental Residuals - Although details of the residual slurries
or recycling operations are not available, it is likely that the per-
chlorate solutions used create highly reactive residuals. The process
has potential for secondary water pollution.
Status of Technology - Electrolysis of lead directly from galena
has been demonstrated on a laboratory scale. No pilot or commercial
plants have been built using this process.
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Purification of Zinc Electrolyte - During the electrolytic manu-
facture of zinc, the roasted concentrate is first leached with sulfuric
acid to dissolve zinc and to precipitate iron and other impurities. The
solution is then purified to remove remaining impurities before transfer
to the electrolytic cells. Treatment reagents such as zinc dust, or-
ganic chemicals, and occasionally arsenic and copper sulfate are used at
U.S. zinc smelters. Purification is multi-stage, with the heavier
elements precipitated out first, followed by the most soluble metals.
There are three additional processes for purification of the leach
liquor in use at foreign smelters.
Process Description - The formation of jarosite for iron precipita-
tion can be used to purify the zinc electrolyte.3,4 jarosite is an
insoluble hydroxylated sodium iron sulfate (NaFe3(S04)2(OH)s). It is
formed in stages, by adding roasted ore, sodium sulfate, and manganese
oxide to the electrolyte under carefully-controlled conditions.
Goethite formation is another process for iron removal.^»4 Goe-
thite, a hydroxylated iron oxide, is formed by the addition of ore
concentrate, lime, and compressed air to the liquor.
More soluble impurities can be removed by a "reverse antimony"
process3»4 which is similar to the multi-stage U.S. technology. Instead
of the highly toxic arsenic, antimony is added to the leach liquor in
two stages with zinc dust. Cobalt and germanium are removed separately
from other elements, allowing for their simplified recovery.
Environmental Residuals - Although insoluble, the stability of
jarosite and goethite as waste materials is not known. The sludges
receive normal by-product recovery, so no additional environmental
problems should be created by jarosite or goethite formation. Substitu-
tion of antimony for arsenic as a treatment reagent may be of environ-
mental benefit, but data are insufficient for evaluation.
Status of Technology - Jarosite formation for iron precipitation is
used by ten smelters in eight countries. Two smelters in Belgium employ
goethite formation for iron removal. Four smelters in Belgium and the
Netherlands use the "reverse antimony" process.
Periodic Reverse Current (PRC)3»5 - Although offering no environ-
mental advantages or disadvantages, use of the PRC process increases the
production rate for an electrolytic copper refinery. Using this tech-
nique, each cell can process about 15 percent more copper than a conven-
tional cell.
Process Description - PRC operates with cathodes of much larger
size and much higher current densities than are found in conventional
electrolytic cells. Every few minutes the current density is reversed
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for a few seconds to dislodge deposited impurities and to prevent the
growth of "needles" on the cathode. The period of current reversal is
reported to be about 10 seconds during each 190 seconds of operation.
Energy Requirements - Current densities up to 375 amp/m^ can be
used.
Environmental Residuals - PRC waste streams are no different from
conventional refineries in proportion to the quantity of copper proc-
essed.
Status of Technology - At least two copper refineries in Sweden and
Japan are operating with the PRC process. There may also be other
installations using this technology.
Pyrometallurgical Developments
There are several pyrometallurgical designs which, like the Noranda,
continuously smelt concentrates to matte and produce continuous streams
of off-gas high in S02 content, suitable for direct application to
sulfuric acid production. The differences between them are in mechan-
ical detail and furnace configuration. Although all the designs are
based on sound theoretical and practical calculations, continuous smelt-
ing must be considered to be in an early stage of development because
none has yet been built in a variety of sizes. The attractions of
continuous processes are readily apparent. With steady, uninterrupted
feed of raw materials and withdrawal of product and wastes, equipment
remains in full-time operation. Continuous plants are easily automated
and can be built in very large sizes. As a result, they are more
economical than other designs, since more product can be made per input
of capital investment, fuel, and operating labor.
WOJRCRA6>7>8 - The WORCRA process, in which smelting and converting
occur in a single furnace, is a development of Conzinc Riotinto of
Australia. Countercurrent flows within the furnace are claimeo to be
one of the principal advantages of this process, since they are said to
provide better separation of slag than in-line flow designs such as
Noranda. The system can be built with equal efficiencies in small-
capacity units.
Process Description - The WORCRA furnace has three distinct zones.
In the center is the smelting zone, into which concentrate and flux are
injected by high-velocity pneumatic or mechanical devices. The re-
sulting agitation helps to separate the slag from the matte. They flow
in opposite directions into converting and slag cleaning zones located
at opposite ends of the furnace. It is this countercurrent flow that
provides good slag separation. Because of the interaction of iron in
the underlying matte, the copper content of the slag is continuously
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reduced as it moves through the slag-cleaning zone. The matte entering
the converting zone is reduced to blister copper which flows contin-
uously from the furnace. The blister copper contains 97.0 to 98.5
percent copper and 0.6 to 0.9 percent sulfur, a low-grade product re-
quiring considerable fire-refining before it can be electrolytically
refined.
Energy Requirements - Fuel consumption is expected to be 50 to 60
percent of that needed for a reverberatory-converter process of compa-
rable throughput.
Environmental Residuals - The WORCRA furnace produces two gas
streams. One consists primarily of combustion gases with very low S02
content, and the other contains most of the S02, in the range of 9 to 12
percent. The slag flows continuously from the end of the slag-cleaning
zone. It contains 0.32 to 0.81 percent copper, a lower concentration
than in other continuous smelters.
Costs - Capital costs are said to be 20 to 30 percent below those
of reverberatory smelting.
Status of Technology - A WORCRA pilot plant processing 73 metric
tons of concentrate per day operated for several months. For a short
period it processed about 100 metric tons per day. There are no com-
mercial installations.
Mitsubishi6*^'^'IC)'I"! _ Mitsubishi Metal Corporation of Japan has
developed a continuous smelting process which is basically similar to
the flash smelting process but which uses three interconnected furnaces.
This arrangement eliminates the need for equipment to transfer molten
matte and also reduces manpower requirements. As with the flash and
Noranda processes, the oxidation of sulfur and iron supplies part of the
process energy requirements and a steady concentrated stream of S02
suitable for acid manufacture is produced. The Mitsubishi is therefore
more energy efficient and less polluting than conventional reverberatory
smelters.
Process Description - Mitsubishi provides for smelting, converting,
and slag cleaning in three separate furnaces, with gravity flow of
product and slag from one furnace to the next. The smelting furnace is
charged with copper concentrates, fluxes, gas or oil fuel, and oxidizing
air through lances installed vertically through its roof. Oxygen en-
richment of air is possible. The matte and slag that are produced flow
continuously to the slag-cleaning furnace. Pyrites mixed with coke are
added to release excess copper trapped in the slag layer. After clean-
ing, the slag is skimmed out continuously and granulated, with the matte
being sent on to the converting furnace. The matte contains 60 to 65
percent copper. The converting furnace is also equipped with lances
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that blow air or oxygen-enriched air and limestone flux through the
mixture, converting matte to blister copper. The slag is recycled to
the smelting furnace by a system of moving buckets, while the blister
copper flows continuously from the furnace. The product analyzes 98 to
99 percent copper and 0.4 to 0.8 percent sulfur, a quality higher than
from conventional furnaces.
Very high smelting rates, about six times greater than that of a
reverberatory furnace and at least twice as much as a flash smelter, are
possible in the Mitsubishi process for two reasons. First, each furnace
has only a single reaction zone, allowing the smelting rate to be in-
creased far more than in conventional furnaces, which have separate
reaction zones for oxidation and settling. Conventional furnaces oper-
ate at a slower rate because oxidation is promoted by good mixing while
settling requires quiescent conditions. Second, the Mitsubishi uses the
reaction heat generated by the oxidation of sulfur and iron efficiently
because the concentrates and fluxes injected through the lances into the
bath are reduced very rapidly.
Energy Requirements - Mitsubishi furnaces are energy efficient
since the reaction heat of the oxidation of iron and sulfur supply part
of the energy requirements. The reduction in energy consumption is at
least 30 percent, and can be a high as 50 percent when coupled with
oxygen enrichment. Energy consumption per metric ton of product has
been reported as follows:
Oxygen 0.3 metric tons
Fuel oil 9.9 x 109 joules
g
Natural gas 1.5 x 10 joules
Electricity 350,000 kg-cal
Total energy consumption is 14.0 x 109 joules per metric ton ot copper,
as opposed to 26.8 x 10$ to 30.2 x 10^ joules per metric ton for conven-
tional reverberatory smelting.
Environmental Residuals - The mixed exhaust gas leaving the fur-
naces contains over 10 percent S02 when the smelting furnace is operated
with air enrichment to 25 percent oxygen. It is a steady stream which
can be readily recovered as sulfuric acid, with total sulfur recovery
over 90 percent. The gas is prepared for the acid plant by cooling and
dry cleaning in ESP's or fabric filters followed by wet scrubbing to
remove fine particulates and excess moisture. Very small particulate
loadings are claimed since the molten liquids flow continuously over
very short distances and the solids are trapped by the furnace bath.
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Since the converter slag is entirely returned to the smelting
furnace, the only slag to be disposed of comes from the slag-cleaning
furnace. An estimated three metric tons of granulated slag per metric
ton of copper product are generated by the Mitsubishi furnaces. Its
major constituents are iron silicates. Flue dusts, consisting of copper
oxides and minor elements, are generated in amounts slightly smaller
than in flash smelting. Dust carryover is about 3 to 4 percent of the
charged concentrate. The flue dusts have to be bled to the extent
necessary to avoid build-up of volatile impurities. The dust bleed
amounts to 0.3 metric ton per metric ton of copper product. Dusts rich
in lead and zinc can be sent to a lead-zinc smelter.
Water effluent streams are created by slag granulation, acid plant
blowdown, and contact cooling of the anodes. They are similar in nature
to those from a conventional smelter, but somewhat larger in volume.
Pollution is minimal. Amounts of water produced per metric ton of
copper are as follows:
Slag granulation
Acid plant blowdown
Contact cooling
50,000 liters
14,000 liters
7,800 liters
Costs - The Mitsubishi process results in slightly higher capital
costs than a conventional smelter, but lower operating costs because of
energy savings. Air pollution control costs are slightly lower because
more concentrated S02 streams are handled. A cost comparison with
conventional reverberatory smelting is as follows:
Pollution control costs
($/metric ton of copper)
Investment
(I/annual metric ton)
Pollution control and
operating cost*
($/metric ton)
Mitsubishi
$ 65
$827
$367
Reverberatory
smelting
$ 59 - 71
$717 -827
$375 -408
* Includes pre-tax return on investment; excludes concentrate
cost.
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Status of Technology - Mitsubishi began development work on this
process in 1961. They have operated a prototype pilot plant with a
monthly capacity of 450 metric tons of copper at their Onahama smelter.
A semi-commercial plant with a capacity of 1500 metric tons of copper
has been operating since 1971. A commercial operation producing 45,000
metric tons per year is being started at Naoshima and a 118,000 metric
ton per year plant is being designed for the Kidd Creek smelter of
Texasgulf.
Kivcet?»T2,13 - The Kivcet process is being developed in the Soviet
Union. It is said that any combination of copper, coal, and zinc con-
centrates may be used as feed, simplifying concentrating procedures and
permitting greater by-product recovery.
Process Description - The Kivcet smelting process contains three
steps. In the first, dried concentrates are reduced in a cyclone fur-
nace with pure oxygen. The melt is transferred to an electric furnace,
where a matte forms which is then sent to a second electric furnace, in
which blister copper and slag are separated and collected.
Energy Requirements - The Kivcet process requires no additional
fuel except when the sulfur content is below 20 percent.
Environmental Residuals - Off gases from the Kivcet process contain
70 to 85 percent S02, and are said to contain all the zinc in the con-
centrate feed because of their high temperature. Zinc metal is re-
covered directly through condensation of off gases from the cyclone
furnace.
Status of Technology - A 45 metric-ton-per-day pilot plant is
believed to be operating in the Soviet Union. No commercial plants
using the Kivcet process have been reported.
Q-S Oxygen Process^, 15 - Drs. Paul E. Quenaw and Reinhardt Schulmann,
Jr., of the United States have developed a continuous smeltint, process
that uses pure oxygen for the autonegous conversion of concentrates to
metal in a single vessel. The process can be used for sulfide ores of
copper, nickel, cobalt, or lead. As in other continuous operations,
over-all capital and operating costs are minimized.
Process Description - The Q-S reactor is an elongated, kiln-like
vessel with crude metal and slag discharge ports at opposite ends. The
cylinder is gently sloping and stepped downward to the lower metal
discharge port. The furnace is rotatable along its axis, simplifying
maintenance and improving heat and mass transfer. The furnace design
and sequential staging of oxygen admission creates a countercurrent flow
of the crude metal and the slag. Bottom-blowing achieves regulated,
localized turbulent baths for optimal gas-liquid-solid contact with
minimal eddy and splash.
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Energy Requirements - Specific energy inputs are not available, but
the use of oxygen minimizes energy consumption and conserves oil and
natural gas. Energy use should be reduced even after the energy used to
manufacture oxygen is taken into account.
Environmental Residuals - The Q-S process creates a single off-gas
stream, minimizing the costs and problems of sulfur fixation and par-
ticulate removal. Slags are said to be low enough in product-metal
content to allow direct disposal.
Status of Technology - A Q-S pilot plant has been built for lead.
No commercial or pilot plants have been announced for copper.
Momoda Furnace7 - One replacement for the conventional copper blast
furnace is the Momoda furnace, a new small-sized design suitable for
lower-grade concentrates which has been developed by the Sumitomo Metal
Mining Company of Japan. It is designed for decentralized installations
where it is economical to minimize transportation of ore. The rated
capacity of this furnace is about 450 metric tons of charge per day.
The Momoda would not be economical in large installations because many
units would be required. However, for smaller operations, smelting
rates are about twice as fast as conventional blast furnaces.
Process Description - The Momoda is a vertical cylindrical furnace.
The concentrate feed is ground to 50 percent minus 325 mesh and is
plasticized to a stiff mass by kneading with flue dust and 10 to 15
percent water. It is extruded into the center of the furnace. Fluxes
such as silica and limestone are ground coarsely and added at the
periphery of the cylinder. This manner of charging creates a bed in
which the center contains fine-grained ore materials and the perimeter
contains coarse-grained flux. During reduction, the gas generated flows
upward through the coarse material rather than through the finer con-
centrate. With the radiant heat being directed into the concentrate,
less dust is contained in the exhaust gas. Matte and slag flow from
separate taps in the bottom of the furnace. The matte can be processed
to blister copper by any standard copper converter.
Energy Requirements - The Momoda uses coal for fuel. It requires
only 28 percent of the heat that a reverberatory furnace uses to produce
the same quantity of matte.
Environmental Residuals - As a result of the charging procedure,
the off-gas from the Momoda blast furnace is very low in particulates.
The exhaust gas is produced continuously and has an S02 content of 7.3
percent and is therefore suitable as feed to a sulfuric acid plant.
Status of Technology - The Momoda furnace is currently in use at
two of Sumitomo's smelters.
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Hydrometallurgical Processing
Hydrometallurgical processes use chemicals in water solution to
break down an ore mineral, permitting recovery of a metal without the
use of high temperature reduction. There are three basic classifica-
tions of hydrometallurgy:
1. Simple heat or vat leaching of ores that are not suitable for
conventional pyrometallurgical processing.
2. Hybrid processes, usually roasting followed by leaching and
electrolysis.
3. Complete hydrometallurgical processes, with no pyrometal-
lurgical steps.
Simple Leaching - Simple heap or vat leaching as practiced in the
western states is an accelerated form of natural weathering (Primary
Copper Production, Process No. 30). In one test, iron bacteria, which
speed up the rate of natural weathering, were also found to speed up the
process in a vat leaching operation.16 They are most effective for ores
containing pyrite, and the weathering rate can increase 5.5 times as
fast with bacteria as without. Since iron bacteria are not known to be
harmful to plant or animal life, no adverse environmental effects should
result from their use.
In-situ leaching, using ammonia or sulfuric acid and ferric sul-
fate, has been tested in Michigan and NevadaJ7 However, if uncon-
trolled, these processes could cause much wider pollution of ground
water than surface operations.
Hybrid Systems - In hybrid hydrometallurgical processes, roasting
of the ore concentrate is generally followed by leaching and electro-
lysis. The "sulfation roasting" used at one U.S. copper plant (Primary
Copper Production, Process No. 34) is an example of a hybrid pt-.cess, as
is the electrolytic process in use at three of the six primary slab zinc
plants. Many types of hybrid systems are being investigated, but with
the exception of the Treadwell process, none of those described has been
operated beyond laboratory scale.
I Q
Treadwell'° - The most extensively tested hybrid hydrometallurgical
process has been developed by Treadwell Engineering Company of New York.
Although there have been problems, the Treadwell process may yet find
future applications since it provides a means of concentrating copper
values prior to pyrometallurgical processing.
Process Description - Copper concentrate is leached in a kiln with
sulfuric acid, forming soluble copper sulfate which is separated from
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insolubles and precipitated with hydrogen sulfide gas. Metallic copper
is produced by reacting cuprous cyanide crystals with hydrogen gas.
Environmental Residuals - Although the potential exists for the
loss of poisonous cyanides, this material is used in many chemical
processes without apparent environmental damage.
Status of Technology - A Treadwell pilot plant with a capacity of
0.9 metric ton of concentrate per day was operated by the Anaconda
Company. The tests were discontinued and Anaconda chose the Arbiter
process because of corrosion problems with the Treadwell.
Mjjiemetl9,20 _ /\ process developed by Minemet Recherche of France
is said to be non-polluting, economically competitive with pyrometal-
lurgical processes, and economically viable even for small installations
of 10,000 to 30,000 metric tons per year.
Process Description - Sulfation-roasted copper concentrate is first
selectively leached with a cupric chloride solution to form cuprous
and ferric chlorides and elemental sulfur. The use of cupric chloride
is favorable since it is not necessary to separate the metal from the
agent after leaching. Pyrites remain unchanged in the residue. The
leach solution is then divided into two streams. Air is injected into.
the first stream and the dissolved iron is precipitated by oxidation as
goethite. Copper is recovered as a cupric ion by solvent extraction
from the second stream, again with air injection to maintain good ex-
traction conditions. The cupric chloride leaching agent is regenerated
in both of these processes. The solvent is then stripped with spent
electrolyte, producing copper sulfate. Copper is recovered by con-
ventional electrolysis.
Energy Requirements - Energy use is said to be low. Only elec-
tricity and air are used, and the extraction and electrolysis stages are
conducted at about 50°C.
Environmental Residuals - If it does not receive further proc-
essing, the leach residue could prove to be an unstable solid waste.
However, the elemental sulfur it contains could be easily recovered by
flotation or other means for internal use or sale. After copper re-
moval, the electrolyte solution can be processed for the recovery of
silver and other metals.
Costs - The Minemet process is considered to be very economical.
Capital costs are low due to the simplicity of process equipment since
all steps are carried out at atmospheric pressure. In addition, the
capacity of the solvent is utilized to the maximum, thereby reducing the
size and number of mixer-settlers required. Operating costs are low
because of the low energy requirements and the complete regeneration of
the leaching agent. There is also the possibility of reclaiming other
metals in the concentrate, including precious metals.
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Status of Technology - Miiiemet has been studying this process on a
laboratory scale. Construction of a pilot plant is planned for 1977.
Fused-Salt Electrolysis^ »22,23,24 _ /\ hybrid hydrometallurgical
process that uses fused-salt electrolysis for the production of lead and
zinc has been developed by the U.S. Bureau of Mines.
Process Description - A mixed lead-zinc sulfide concentrate is
leached with an acid solution at 100°C and a pressure of 3.5 kg/cm? in
the presence of chlorine and oxygen. Zinc, copper, and cadmium are
dissolved as metal chlorides, and lead and silver remain in the leach
residue and are dissolved in a subsequent brine leach at 90°C. Upon
cooling of the solution, lead chloride crystallizes and is separated by
filtration. The remaining filtrate is treated by solvent extraction to
recover copper, then with zinc powder to recover remaining impurities.
Anhydrous zinc chloride is recovered from the resulting purified solu-
tion by evaporation. To form a stable low-melting electrolyte, potas-
sium and lithium chlorides are mixed with the lead and zinc chlorides,
which are then electrolyzed individually in sealed cells with graphite
electrodes. Electrolysis is conducted at 500°C, and the electrodes
operate at a current density of 0.8 amp/cm^. Chlorine gas is liberated
at the anode and captured for recycle. The lithium and potassium are
not affected by electrolysis and can be used for months without replace-
ment. In an alternative process applicable only to lead concentrates,
ferric chloride solution is the leach solvent used for preparing the
lead chloride crystals.
Energy Requirements - Energy required for electrolysis is 95 kg-
cal/kg for lead production and 455 kg-cal/kg for zinc. Current effi-
ciency is 95 percent.
Environmental Residuals - The disposition of impurity metals in
these fused-salt electrolytic processes has not been described. How-
ever, they must be discarded as wastes since the purity of the lead or
zinc product is expected to be more than 99.9 percent.
Costs - Costs should be lower than in conventional electrolysis
because of the energy conservation and the elimination of much of the
labor involved in charging, stripping, and remelting cathodes.
Status of Technology - Experiments with fused-salt electrolysis are
being conducted on the laboratory scale.
_ A hybrid hydrometallurgical process using a different
approach has been developed by Battelle-Pacific Northwest Laboratories.
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Process Description - Following a neutral roast of copper concen-
trate, the roasted calcine is leached in hydrochloric acid, dissolving
the iron and generating hydrogen sulfide gas. Because much of the iron
and sulfur have been removed, the leach residue contains high concentra-
tions of copper. The residue is processed to copper by conventional
pyrornetallurgical methods.
Environmental Residuals - S02 from the pyrometallurgical operations
can be reacted with the hydrogen sulfide from leaching to form elemental
sulfur, although proper control could prevent the formation of either
gas. The leach solution can be processed to recover the hydrochloric
acid and to produce a high-grade iron oxide suitable for sale.
Costs - Economics of the Battelle process are said to be competi-
tive with conventional technology.
Status of Technology - Laboratory-scale tests are being conducted.
Lime-Concentrate-Pellet Roast^6 - Another modification of sulfation
roasting which has not been as thoroughly studied as the Minemet is
1ime-concentrate-pellet roasting.
Process Description - Lime is mixed with the copper concentrate and
the mixture is pelletized before roasting. Most of the sulfur in the
ore is converted to calcium sulfate. A sulfuric acid leach dissolves
the copper, which is then extracted by electrolysis.
Environmental Residuals - This process would add no significant
environmental impacts to those of conventional sulfation roasting.
Status of Technology - Research has been limited to laboratory
tests.
Nitrogen Roast^7 - Another hybrid hydrometallurgical process being
investigated by the U.S. Bureau of Mines involves roasting in an atmo-
sphere of nitrogen.
Process Description - Copper concentrate is roasted in a nitrogen
atmosphere, removing about 20 percent of the sulfur present. A subse-
quent hydrochloric acid leach removes most of the iron. A second leach
dissolves the copper, which can then be extracted electrolytically in a
special cell in the presence of S02.
Environmental Residuals - Environmental hazards cannot be deter-
mined based on available data.
Status of Technology - Laboratory testing only has been conducted.
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Distillation with AcetonitriTe^ - A process for removing copper
from sulfation-roasted concentrate is being developed in Australia.
Process Description - A sulfation-roasted concentrate is leached
with sulfuric acid and the organic chemical acetonitrile. Metallic
copper is formed upon distillation of the solution, thereby eliminating
electrolysis.
Energy Requirements - Energy use is said to be less than 60 percent
of that for conventional processing.
Environmental Residuals - Acetonitrile is poisonous and its wide-
spread use might cause serious pollution.
Status of Technology - Australian research into distillation with
acetonitrile has proceeded at the laboratory level only.
Complete Hydrometallurgy - Complete hydrometallurgical processes
are being tested primarily for copper production. They treat the ore
concentrates directly with chemicals, using no pyrometallurgical steps.
The CLEAR, Cymet, and Arbiter processes (Primary Copper Production,
Process Nos. 36 to 45) are such processes. Current research appears to
be in progress primarily in the United States and Canada.
Sherritt-Gordon/Cominco^0.18,29,30 _ /\ complete hydrometallurgical
process has been developed as a joint venture of Sherrit-Gordon and
Cominco.
Process Description - This process is expected to be a ferric
chloride leach, but few details have been released.
Environmental Residuals - As with other chloride-based hydrometal-
lurgical systems, problems in the disposal of soluble waste chlorides
are likely.
Status of Technology - The Sherritt-Gordon/Cominco process is
probably now operating in a pilot plant at Fort Saskatchewan in Alberta,
Canada.
DuPont-Kennecott30,31 _ DuPont and Kennecott Copper have patented a
complete hydrometallurgical process based on a mixture of nitric and
sulfuric acids.
Process Description - Ore concentrate is first leached with a
mixture of nitric and sulfuric acids, after which the nitric acid is
decomposed and recovered. Iron is removed by precipitation and filtra-
tion, and the copper is extracted by electrolysis. The process is
pressurized, operating at temperatures up to 200°C.
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Status of Technology - A pilot plant is now under construction, and
this has the appearance of a serious and well-organized development
effort.
29
Lurgi-Mittenburg - The Lurgi-Mittenburg process is a complete
hydrometallurgical system developed in Germany and Austria.
Process Description - Sulfuric acid is used in a high-pressure
leach. Pressures up to 20 kg/cm? and temperatures of 115°C promote
direct oxidation of the minerals with oxygen gas. Copper is recovered
by electrolysis.
Status of Technology - A demonstration plant is now operating.
Other Developments in Hydrometallurgy - There are two other de-
velopments in hydrometallurgical technology which are finding applica-
tion in existing processes.
O A
Treatment of Zinc Silicate Ores - An auxiliary application of
hydrometallurgy is being tested in connection with zinc production from
silicate ores by the Electrolytic Zinc Company of Australia. The prob-
lem in leaching these ores is the formation of colloidal silica which
must be removed before the solution can be treated by electrolysis. If
successful, this procedure may find application in many industries,
since silicate ores have proved to be one of the more difficult indus-
trial separations.
Process Description - A complex coagulation procedure is being
developed to solve this problem. The leach solvent is sulfuric acid,
spent electrolyte from the zinc cells.
Environmental Residuals - Application of the process should create
no new environmental problems.
Status of Technology - A pilot plant with a capacity of 5-metric-
tons-per-day is operating.
Miscellaneous Technologies
There are several metallurgical technologies that have been pro-
posed or that are in use outside the U.S. that fit no generalized clas-
sification.
7 35
Torco * - Copper silicates cannot be extracted by flotation
methods from gangue rock. Torco overcomes this problem. It is a pyro-
metallurgical process for recovering copper from silicate ores such as
chrysocolla and dioptase. It is based on the fact that copper silicates
form growths of elemental copper when they are crushed, mixed with
sodium chloride, and smelted in an inert atmosphere.
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Process Description - Coarse silicate ore is first crushed, dried,
and ground. It is then mixed with coal and heated in a fluidized-bed
reactor. The heated mixture overflows into a second fluidized-bed
reactor, into which sodium chloride and additional coal are added.
Copper forms as granules, and the mixture is water quenched, ground, and
then separated by flotation. Cyclones are integral parts of the flu-
idized-bed reactors. The low-grade copper product is mixed with the
feed to a pyrometallurgical smelter.
AMAX Base Metals Research and Development Corporation has experi-
mented with a modification of the Torco process for use with sulfite
ores of copper, such as chalcopyrite. After dead roasting to remove all
sulfur, the ore concentrate is mixed with sodium chloride and heated to
700° to 800°C. Granules of copper form, which are separated by flota-
tion and melted to form blister copper of 98 percent purity.
Environmental Residuals - Sulfur dioxide emissions depend on the
sulfur content of the silicate ore. Additional particulate control
equipment may be used in addition to the cyclones.
Status of Technology - The Torco process is in use in Zambia and
Mauritania. The AMAX modification has only been demonstrated in labora-
tory tests.
TBRC6.7.10.36.37 _ The JBRC op blQwn rotary converter) was orig-
inally designed for nickel production. However, tests by its developer,
the International Nickel Company of Canada (INCO). have shown that it is
applicable to other metals. They claim that the TBRC, using U.S. con-
centrates, can produce copper that can be directly cast as anodes, thus
eliminating fire-refining. The TBRC is said to provide increased opera-
tional flexibility through close control of both turbulence and tempera-
ture while maintaining high thermal efficiency.
Process Description - The TBRC is a rotatable refractory-lined
furnace that can be tilted for filling, blowing, or emptying. The
vessel atmosphere is controlled by injecting natural gas and a'r, oxy-
gen, or oxygen-enriched air onto the molten surface through a water-
cooled lance in the hood. The vessel rotates constantly, providing
excellent gas-solid-liquid contact, rapid mixing of the charge, and even
heat distribution which increases the rate of conversion reaction.
Oxygen efficiency greater than 95 percent can be obtained by adjusting
rotational speed and angle of the injection lance. When pure oxygen is
used, no heat is carried away by atmospheric nitrogen. The high tem-
peratures resulting eliminate more impurities and cause better separa-
tion of slag from the copper than with other converters.
Energy Requirements - Because of its high thermal efficiency and
elimination of fire-refining, the TBRC uses less fuel than other con-
verters.
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Environmental Residuals - When oxygen is used with the TBRC, the
exhaust gases contain about 50 percent S02« However, it is a batch-type
mechanism that does not produce a continuous stream of gas. Because of
the turbulence in the converter, most solid particles are captured by
the liquid or by the wet surfaces above the liquid. This results in
lower particulate loadings and fugitive emissions. As with conventional
converters, slags contain sufficient copper to be economically returned
to the smelting process.
Costs - By eliminating fire-refining, the TBRC can reduce a smel-
ter's capital investment. However, it does require a source of oxygen
for highest operating efficiencies.
Status of Technology - The first commercial copper TBRC plant, now
under construction in British Columbia, is scheduled for completion in
late 1977.
Oxygen-Enriched Smelting ' - In addition to the continuous Q-S
Oxygen process already discussed, other applications of pure oxygen or
oxygen-enriched air to smelting have been developed. They provide the
benefits of decreasing fuel requirements while increasing the process
capacity. Oxygen enrichment can increase smelting rates by 25 to 50
percent; an additional 3 to 8 tons of charge can be smelted for every
ton of oxygen.
Process Description - Oxygen enrichment can be applied to conven-
tional reverberatory smelting, increasing fuel efficiency. The in-
creased operating temperature decreases refractory life, so enrichment
of converter air is normally limited to about 27 percent oxygen for
conventional bottom-blown converters. The increased temperature im-
proves separation of the matte and reduces metal losses. In addition to
increasing S02 concentrations, enrichment also increases the converter's
scrap melting capability.
The specific capacity of furnaces such as flash smelters can be
increased by the use of oxygen. Since waste heat recovery is already
practiced in these furnaces, the increase in energy efficiency is not as
great as in the case of converters.
Energy Requirements - Oxygen enrichment results in lower energy use
than conventional processes, even after the energy used in oxygen manu-
facture is taken into account. Fuel efficiency is increased particu-
larly where waste heat is not recovered.
Environmental Residuals - The use of oxygen at a smelter does not
significantly change the nature and amount of effluents. Although there
are no published data in this area, the higher operating temperatures
may result in an increased circulating dust load. The higher tempera-
tures may also result in more NOX formation, although these gases are
probably absorbed in the acid after scrubbing and treatment in an acid
375
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plant. The oxygen plant itself uses significant quantities of indirect
cooling water, and it may contain oil and grease, corrosion inhibitors,
and chlorides.
Costs - Oxygen enrichment should result in slightly lower capital
costs because of the increase in capacity. Maintenance requirements
will be higher, but operating costs should decrease because of lower
fixed costs.
Status of Technology - Oxygen enrichment of combustion air has been
tested with Noranda, Mitsubishi, and WORCRA furnaces. It has been
practiced since about 1972 at the flash smelter at Harjavalta, where it
has enabled almost completely autogenous furnace operations and resulted
in a large increase in furnace capacity. The only major reverberatory
smelter using oxygen is at Onahama in Japan.
38
Sea Nodule Processing - A highly specialized hybrid process in
Canada is recovering copper, nickel, and cobalt from sea nodules.
Process Description - This process features reduction in a pyro-
metallurgical kiln followed by smelting in an electric furnace and
hydrometallurgical leaching in an oxygen atmosphere. Although intended
primarily for nickel recovery, copper equivalent in weight to 60 percent
of the nickel is also extracted.
State of Technology - Sea nodule processing is being practiced, but
the scale of operations is not known.
Slag Treatment
Slags from most continuous copper pyrometallurgical processes
contain sufficient quantities of the product metal to justify recovery
before disposal. The metal is in the slag either as entrapped globules
of a second phase such as matte or copper or as dissolved metals. Slag
treatment in an electric furnace is now used at the single U.S. flash
smelter and will be used at the continuous smelter soon to be built.
This process has been previously discussed (Primary Copper Production,
Process No. 12).
There are two other basic recovery techniques available for slag
treatment. At some Japanese and Canadian smelters, flotation techniques
are used with slags that have a high sulfur content.7>39>40 Another
possibility in use outside the U.S. is hydrometallurgical processing of
slags, consisting of grinding and leaching with acids to form copper
solutions, from which metal is recovered by conventional concentrating
techniques or electrolysis.41 With both these methods, however, the
slags are discarded as finely ground tailings. Without adequate stabil-
ization or protection, they are less stable in weathering and would
generate more fugitive dusts than electric furnace slags, which are
masses of good crystalline composition. In addition, crystals in hydro-
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metallurgical slags would be corroded and saturated with acids, and
therefore would be more likely to cause secondary water pollution.
Although capital costs are similar to electric furnace treatment,
operating costs are significantly higher in either case. Hydrometallur-
gical treatment is probably the most expensive, since it requires both
grinding and electrolysis for complete recovery of the metal.
377
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REFERENCES
1. Encyclopedia of Chemical Technology. New York, Interscience Pub-
lishers, a division of John Wiley and Sons, Inc., 1967.
2. Shewes, H.R. Electrowinning of Lead Directly from Galena. Proc.
Australasian Inst. Win. Met., December 1972. (244). pp. 35-41.
3. Kawakita, T., Kitamure, T. Sakoh, Y. and Sasaki, K. Design, Con-
struction, and Operation of Periodic Reverse Current Process at
Tamano. Presented at the 105th annual meeting of the AIME. Las
Vegas, Nevada. February 22-26, 1976.
4. Lindstrom, R. Production Unit with Current Reversal at the Ronnskar
Works of Boliden AKTIE bolag, Skelleftehamn, Sweden. Presented at
the 104th annual meeting of AIME. New York, New York. February
16-20, 1975.
5. Claessen, P.L. Electrowinning of Zinc under Periodic Reverse and
Pulsating Current Conditions. Presented at the 104th annual meet-
ing of the AIME. New York, New York. February 16-20, 1975.
6. Background Information for New Source Performance Standards: Pri-
mary Copper, Zinc, and Lead Smelters. Volume I, Proposed Stan-
dards. Environmental Protection Agency, Research Triangle Park,
North Caroline, EPA-450/2-74-002a. October 1974.
7. Price, F.C. Copper Technology on the Move. Chemical Engineering.
New York, New York. April 16, 1973.
8. Two Direct Smelting Processes Designed for Pollution Control in
Copper Plants. Engineering and Mining Journal. New York, New
York. April 1971.
9. Nagano, T. and Suzuki, T. Commercial Operation of Mitsubishi
Continuous Copper Smelting and Converting Process. Presented at
the 105th annual meeting of the AIME. Las Vegas, Nevada. February
22-26, 1976.
10. Dolezal, Henry et al. Environmental Considerations for Emerging
Nonferrous Metal-Winning Processes (Field Review Copy). U.S.
Bureau of Mines. Salt Lake City Metallurgy Research Center.
378
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11. Arthur D, Little, Inc. Environmental Considerations of Selected
Energy Conservation Manufacturing Process Options. Vol. 14,
Primary Copper Industry (Draft), EPA Contract No. 68-03-2198.
Cambridge, Mass.
12. Kivcet. Pamphlet by V/0 Licensintorg Mostva-SSSR. Distributed by
Southwire Company. Carroll ton, Georgia.
13. Kivcet Process for Complex Ores. World Mining. June 1974. pp.
26-27.
14. Queneaw, P.E. and Schuhmann, Jr., R. The Q-S Oxygen Process.
Journal of Metals. New York, New York. August 1974. pp. 14-16.
15. Matyas, A.G. and Mackey, P.J. Metallurgy of the Direct Smelting of
Lead. Metallurgical Society of AIME. New York, New York. IMS
Paper No. A75-80. February 1975.
16. Michalek, Z. and Malik, A. The Influence of the Mineral Composi-
tion of Copper Ores on the Effect of Biological Leaching. Archives
Hutn., 1975, 20, (3) pp. 421-432.
17. Hockings, W.A. and Freyberger, W.L. Laboratory Studies of In-Situ
Ammonia leaching of Michigan Copper Ores. Presented at the 105th
annual meeting of the AIME. Las Vegas, Nevada. February 22-26,
1976.
18. Chemical Route to Copper? Chemical Engineering. New York, New
York, April 20, 1970. pp. 64-66.
19. Demarthe, J.M., Gandon, L. and Georgeawx, A. A New Hydrometal-
lurgical Process for Copper. Presented at the 105th annual meeting
of the AIME. Las Vegas, Nevada. February 22-26, 1976.
20. Demarthe, J.M. et al. A New Hydrometallurgical Process for Copper.
Presented at the 105th Annual Meeting of AIME. Las Vegas, Nevada.
February 22-26, 1976.
21. Haver, F.P., Elges, C.H., Bixby, D.L. and Wong, M.M. Recovery of
Lead from Lead Chloride by Fused-Salt Electrolysis. U.S. Depart-
ment of the Interior. Washington, D.C. Bureau of Mines Report of
Investigations 8166. 1976.
22. Haver, F.P. and Wong, M.M. Ferrix Chloride - Brine Leaching of
Galena Concentrate. U.S. Department of the Interior. Washington,
D.C. Bureau of Mines Report of Investigations 8105. 1976.
379
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23. Haver, P.P., Shanks, D.C., Boxby, D.L. .and Wong, M.M. Recovery of
Zinc from Zinc Chloride by fused-Salt Electrolysis. U.S. Depart-
ment of the Interior. Washington, D.C. Bureau of Mines Reprot of
Investigations 8133. 1976.
24. Bureau of Mines Research 1975. U.S. Bureau of Mines. Washington,
D.C.
25. Rohrmann, C.A. and Fullam, H.T. Control of Sulfur Dioxide Emis-
sions from Copper Smelters. Volume II, Hydrogen Sulfide Production
from Copper Concentrates. EPA, Office of Research and Development.
Research Triangle Park, North Carolina. PB>237 928. September
1974.
26. Bartlett, R.W. and Haung, H.H. The Lime-Concentrate-Pellet Roast
Process for Treating Copper Sulfide Concentrates. Journal of
Metals. December 1973. pp. 28-34.
27. Gabler, R.C., Dunning, B.W., Brown, R.E. and Campbell, W.J. Proc-
essing Chalcopyrite Concentrates by a Nitrogen Roast-
Hydrometallurgical Technique. U.S. Department of the Interior.
Washington, D.C. Bureau of Mines Report of Investigations No. 867.
1975.
28. Propose Copper Leaching Process. Industrial Research. August
1976. pp. 38.
29. Habashi, F. Pressure Hydrometallurty: Key to Better and Nonpol-
luting Processes. Engineering and Mining Journal. New York, New
York. May 1971. pp. 88-94.
30. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometallurgy.
Chemical Engineering. Volume 83, No. 1. pp. 79-81. January 5,
1976.
31. Brenneck, H.M. Recovery of Metal Values from Ore Concentrates.
U.S. Patent No. 3,888,748. June 10, 1975.
32. Arbiter, N. and Milligan, D. Reduction of Copper Ammine Solutions
to Metal with Sulfur Dioxide. Presented at the International
Symposium on Copper Metallurgy. 105th annual meeting of the AIME.
Las Vegas, Nevada. February 22-26, 1976.
33. Arbiter, N., Milligan, D. and McClincy, R. Metal Production from
Copper Ammine Solution with Sulfur Dioxide. Presented at IChemE
International Symposium on Hydrometallurgy. Manchester, England.
1975.
380
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34. Matthew, I.G. and Eisner, D. The Hydr(metallurgical Treatment of
Zinc Silicate Ores. Metallurgical Society of AIME. New York, New
York. TMS Paper No. A75-74. February 1975.
35. Opie, W.R., Coffin, L.D. and Casanelli, D.C. A Minimum Pollution,
Low-Energy Pyrometallurgical Process for Treating Chalcopyrite
Concentrates. Presented at the 105th annual meeting of the AIME.
Las Vegas, Nevada. February 22-26, 1976.
36. Daniele, R.A. and Jaquary, L.H. TBRC A New Smelting Technique.
Dravo Corporation, Pittsburgh, Pennsylvania. Also presented at
1972 annual meeting of the AIME. February 20-24, 1972.
37. Mason, J.L. and Siscoe, S.F. Afton Mines Smelter to Use Top Blown
Rotary Converter Process. The Northern Miner. April 22, 1976.
pp. B12-B13.
38. Sridhar, R. and Jones, W.E. Extraction of Copper, Nickel and
Cobalt from Sea Modules. Presented at the 104th annual meeting of
the AIME. New York, New York. February 16-20, 1975.
39. Hallett, G.D. Continuous Copper-Smelting Process Uses Single
Vessel. Chemical Engineering. New York, New York. April 26,
1976. pp. 62-64.
40. Mills, L.A., Hallett, G.D. and Newman, C.J. Design and Operation
of the Noranda Process Continuous Smelter. Presented at the Inter-
national Symposium on Copper Extraction and Refining. 105th annual
meeting of IAME. Las Vegas, Nevada. February 22-26, 1976.
41. Kmetova, D. Hydrometallurgical Method of Copper Recovery from
Slags. Zbornik V.S.T. Kosiciach. 1971. pp. 191-197.
381
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8.0 RECOMMENDED RESEARCH AND DEVELOPMENT PROGRAMS
It is evident that many areas within the primary copper, lead, and
zinc industries warrant investigation. Presented below are several
suggested programs - some applicable to only one industry, some to all.
The purposes of these programs would be to eliminate some uncertainty
about the potential environmental effects of the industry operations and
to develop control techniques and policies that minimize environmental
impacts.
Determination of the most effective method of disposition of "black acid"
from a copper electrolytic refinery.
Need:
Almost one percent of the input to an electrolytic refinery is impurity
elements that refining removes from the product. The process yields a
strongly acidic waste stream that contains many of these elements,
including much of the nickel, cadmium, arsenic, antimony, and bismuth
that enters with the anode copper. There is apparently no effective
method to dispose of this stream.
Project Scope:
(a) Establish the quantity and composition of waste electrolyte
solution from operating electrolytic refineries.
(b) Determine what methods are currently used to dispose of this
material.
(c) Define best control technology for this waste.
Determination of heavy metals concentrations in effluents from tailings
piles, mine waters, and milling Wastes.
Need:
Mining and milling operations generate large quantities of wastewater
containing various heavy metals. Before environmental effects of these
heavy metals can be properly evaluated, the specific compounds must be
identified and their chemical states and quantities defined.
382
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Project Scope:
(a) Perform field survey and sampling of water from various
operating mines and mills.
(b) Perform elemental analysis and identification of specific
compounds.
(c) Establish the quantity of release of each of these compounds
from industrial sources.
Investigation of high-boiling liquids for the removal of particulates
from a smelter gas stream.
Need:
Smelter gases are high-temperature streams containing hazardous solid
particles. Methods for removal of these particulates require cooling of
the gases, usually by evaporation of water into the stream, resulting in
an excess of water that must be removed prior to the treatment of the
gas for S02 removal. Use of stable high-boiling liquids to remove these
particulates from a hot gas stream could eliminate the excess of water,
enable recovery of heat in an auxiliary boiler, and provide a method for
handling the hazardous dusts without fugitive losses.
Project Scope:
(a) Identify liquids of high boiling points that might serve for
scrubbing liquors at temperatures above 100°C.
(b) Design and evaluate a system for high-temperature treatment of
smelter gases that minimizes introduction of water vapor into
the gas stream.
Disposal of DMA process wastes.
Need:
DMA absorption of SO? produces a small, concentrated liquid waste stream
for which there is no simple method of disposal.
Project Scope:
(a) Determine stream condition.
(b) Determine methods for fixation or removal of hazardous materials
determined in a).
(c) Demonstrate disposal technique selected.
383
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Determination of a complete mass and energy balance, including all
operations from mine to product.
Need:
A detailed process flow sheet with a chemical mass balance would be
especially useful in identifying feasible emission controls and the
effects of process changes. An energy balance on the complete mine-to-
product mass flow would be used to identify areas where the process
could be modified to improve energy utilization. A smelter operator is
better able to afford pollution control when his process is optimized
for energy utilization.
Project Scope:
(a) Generate a complete mass balance for smelter operation, using
concentrate analyses, stream analyses, thermodynamics and
chemical kinetics data, and process flows.
(b) Prepare an energy balance for smelter operation based on
available and field-generated data.
Assessment of the quantity and concentration of metallic elements in
the waste gas streams of lead smelters.
Need:
It is known that fumes of cadmium, arsenic, antimony, and other metals
are evolved in lead smelters from sintering, slag fuming, reverberatory
softening, and recovery of cadmium and antimony. These fumes are cooled
and condensed in chambers from which the exhaust gases are sent to the
atmosphere, generally without further treatment. Published literature
gives no data on the effectiveness of the control devices being used.
Project Scope:
(a) Define the quantity and composition of fume emissions from the
above-listed waste gases.
(b) Establish the effectiveness of simple condensation in removing
sublimed fumes from such streams.
(c) Define acceptable methods of control of these emissions.
384
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Determination of slag cleaning furnace exhaust stream characteristics
arid emission control methods.
Need:
Electric slag cleaning furnaces are being installed in conjunction with
new pyrometallurgicai processes for copper. Gaseous exhausts from these
furnaces are low in volume but have not been fully characterized.
Emissions may include trace element fume, carbon monoxide, and hydrogen.
Requirements for the development of emission control technology should
be determined.
Project Scope:
(a) Determine exhaust gas flows and pollutant exhaust concentra-
tions, including suspected heavy metals, carbon monoxide, and
metal carbonyls.
(b) Determine types of pollution control systems that are effec-
tive to control the emissions that were characterized in a).
(c) Establish emissions performance standards for slag cleaning
furnaces.
Evaluation of the use of sulfides to precipitate heavy metals from
waste waters.
Need:
Sulfides have been proposed for use to remove dissolved metals from
wastewaters. The treatment may effectively lower metals concentrations,
but several attendant disadvantages should be investigated and addressed.
Project Scope:
(a) Determine residual sulfide concentrations in effluent neces-
sary to remove metallic species from wastewaters.
(b) Determine heavy metal removal effectiveness and secondary
resolution of precipitated metals.
(c) Evaluate potential of ^S formation from treated waters.
(d) Determine disposal techniques for metal sulfide sludges that
result from the treatment process.
385
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Study the problems of arsine evolution from tank houses.
Need:
Arsine, a very toxic form of arsenic and a suspected carcinogen, is
formed in electrolytic tank houses. Arsine is a gas evolved when the
outlet concentration of the electrolyte falls below 3 grams of copper
per liter. There is evidence of additional arsine formation where
overall copper concentrations in the electrolyte are higher. Arsine may
pose a significant hazard to the environment.
Project Scope:
(a) Measure arsine concentrations at various suspect points in the
refinery operation.
(b) Develop means of extracting arsenic before the electrolytic
refinement stage.
(c) Investigate transport of arsine in the ecosystem.
(d) Determine toxicity level of arsine for various organisms.
Evaluation of the environmental aspects of expanded hydrometallurgical
practices.
Need:
Gradual reductions in the available quality of copper ores and increased
supplies of sulfuric acid resulting from S02 control requirements tend
to support an expansion of hydrometallurgical methods for the recovery
of copper. The disposal of increased quantities of resulting spent
leach solutions should be assessed.
Project Scope;
(a) Determine the chemical characteristics of spent leach solutions.
(b) Assess the environmental effects resulting from the disposal
of untreated spent leach solutions.
(c) Determine treatment requirements to insure that effluent
liquids from hydrometallurgical processes meet Federal and
local standards.
(d) Fund R&D efforts to develop additional control processes that
may be required.
386
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Investigation of the use of cementation as a wastewater treatment
procedure.
Need:
Metallic iron has been used to remove copper and other heavy metals from
solution in acidic streams, and laboratory tests have indicated that the
technique may be suitable as a wastewater treatment method under certain
conditions. Practical applications have not been demonstrated, and
potential environmental effects have not been assessed.
Project Scope:
(a) On the basis of current laboratory results prepare a descrip-
tion of field conditions where cementation may be a viable
wastewater treatment method.
(b) Predict treated wastewater chemical characteristics and deter-
mine whether effluent guidelines can be met.
(c) Determine modifications or additional treatment methods that
may be required to meet effluent guidelines.
(d) Analyze economics of treatment of wastewater by cementation in
comparison with other treatment methods.
(e) Pilot plant demonstration under field conditions.
Investigation of pretreatment techniques for removal of impurities from
concentrate prior to smelting.
Need:
Most of the environmental problems associated with base-metal production
are related to the impurity content of the concentrates entering the
smelters. Removal of the impurities prior to smelting will eliminate
volatile metals and reduce the potential for adverse effects from flue
dusts and slags. Volumes of waste will be considerably decreased.
Project Scope:
(a) Investigate new leaching procedures for preleaching of the
concentrate.
(b) Develop more efficient flotation operations or additional
stages of flotation to remove impurities or by-products.
(c) Demonstrate technology in cooperation with the industry.
387
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Assessment of the environmental effects froni tailings containing
elemental sulfur, and determination of methods for their control.
Need:
Most of the developing hydrometallurgical processes for copper produc-
tion will produce large quantities of elemental sulfur, which will be
discarded along with waste rock as a tailings. Sulfur is a strongly
reduced material, which is used for many purposes, including application
as an insecticide and as an acid additive in soil treatment. Prior to
large-scale application of these new processes, a study to evaluate the
long-term stability of the tailings and to devise effective methods of
control is recommended.
Project Scope:
(a) Estimate the quantities of sulfur and other constituents to be
expected from proposed hydrometallurgical processes.
(b) Outline the most likely chemical changes that will take place
in this material during extended outdoor exposure.
(c) Confirm the above findings by controlled tests.
Investigate conversion of ferric oxide sludge from hydrometallurgical
processes to a form suitable for use in other industries.
Need:
Many copper concentrates contain such high levels of iron that they
could also be used as iron concentrates. The residues from leaching
operations contain significant amounts of iron and other marketable
metals. The use of these residues for raw materials in iron and other
metal industries warrants investigation. Such a procedure cou'd make
hydrometallurgy more economically attractive and reduce the environ-
mental problems of sludge disposal.
Project Scope:
(a) Phase-identify residual solids from conventional hydrometal-
lurgical processes.
(b) Investigate possible conversion of leach residue into concen-
trate for iron production.
(c) Consider alternative solvents for recovery of copper and iron
in a desirable form.
388
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Decontamination of wastewater from lead milling operations.
Need:
The water that is used in a lead concentrator is recycled to a certain
extent, but eventually it is discharged in a highly mineralized condi-
tion. Net evaporation conditions in the West and high natural pH
conditions in Missouri tend to minimize the necessity for further
treatment, but a full assessment of environmental effects has not been
made.
Project Scope:
(a) Determine concentrations of metallic species in concentrator
effluents.
(b) Evaluate potential environmental hazards associated with
discharges determined in a).
(c) Determine control system requirements and costs to reduce
effluent concentrations to acceptible levels.
Determination of leaching properties of lead mine spoil and other lead
smelter residues.
Need:
Solid wastes from lead mining and smelting operations are dewatered and
impounded or used for backfill for road building. The natural alkali-
nity of surface waters in the Missouri area is reported to inhibit
dissolution of heavy metals, but possible long term effect of leaching
and transfer of metals to other areas have not been studied.
Project Scope:
(a) Determine ultimate disposition of heavy metal values in
discarded materials and the stability of metallic species in
spoil materials and mill refuse ponds.
(b) Determine leach transfer rates for heavy metals in discarded
materials.
(c) Suggest methods to stabilize residues if stability is found to
be unsatisfactory.
389
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Investigation of economics for hydrometallurgical production of lead.
Need:
Hydrometallurgical production of lead has been suggested for environ-
mental reasons, but economics are still not sound. Considerable develop-
ment will be required to make the process competitive.
Project Scope:
(a) Conduct research to determine more favorable kinetics for lead
chlorination hydrometallurgy.
(b) Investigate the possibility of bacterial leaching; some geo-
chemists believe that some galena deposits were bacterially
formed.
(c) Develop Sherritt-Gordon high-pressure leach technology for
lead.
(d) Determine economic factors, such as the price of coke, that
would promote the use of hydrometallurgy.
Optimal production of lead sinter exhaust gas for the manufacture of
sulfuric acid.
Need;
The windbox of an updraft sintering machine is generally divided into a
strong gas and a weak gas section, where the strong gas is taken to a
sulfuric acid plant. The weak gas S02 concentration is too low for acid
manufacture, so it is discharged only through a particulate control
device. Optimization of gas flows would minimize the S02 emissions.
Project Scope:
(a) Determine parametric relationships between sinter feed quality
and variation, windbox gas flow and recirculation, windbox S02
concentrations, sinter quality, process reliability, and acid
plant operability.
(b) Investigate potential use of oxygen enrichment to upgrade
sinter gas quality for sulfuric acid production. Compare with
cost feasibility for blast furnace S02 control.
390
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Lead blast furnace SOp recovery.
Need:
S02 emissions from a lead blast furnace account for only about 5-10
percent of the sulfur originally present in the lead concentrate. If it
is determined that these emissions require control, effective control
systems will have to be demonstrated.
Project Scope:
(a) Determine whether sufficient S02 control can be obtained at
lead smelters without controlling blast furnace emissions.
(b) Describe systems for S02 control and estimate control costs
for blast furnace control.
Metal and trace material emission control efficiency - baghouses,
electrostatic precipitatbrs, and scrubbers.
Need:
Collection efficiencies for various materials vary widely with the type
of control device that is employed. Vapors that are not condensed by
the control devices cannot be collected and may escape from baghouses
and electrostatic devices. Small particles may escape from all three
devices.
Project Scope:
(a) Determine penetration parameters for the three collection
devices in pyrometallurgical service. Zinc, arsenic, and
cadmium emissions, as well as other potential harmful emis-
sions should be determined.
(b) Using an approved health effect or toxicological rating
scheme, rate the three given control devices for the chosen
pyrometallurgical service.
Zinc fuming emissions evaluation.
Need:
Zinc is recovered from some lead slags, either as reduced metal or as
zinc oxide. The fuming process that is used may result in emissions of
metallic fume, carbon monoxide, or other materials. These emissions may
require additional pollution controls.
391
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Project Scope:
(a) Determine pollutant species and quantities emanating from zinc
fuming operations.
(b) Determine optimum control devices for the pollutants that are
emitted during the fuming operation.
(c) Determine emissions performance criteria for zinc fuming
plants.
Establishment of effective disposal procedures for slags from the kettle
softening and bismuth refining processes in lead smelters.
Need:
Two smaller-volume slags from a lead smelter are discarded in open
dumps, even though they contain water-soluble components. One contains
sodium salts of arsenic, antimony and tin; the other contains chlorides
of lead, magnesium, and calcium. At this time, there is no method to
dispose of these materials that will assure no secondary water pollution
from these sources.
Project Scope:
(a) Determine quantity and composition of this waste from opera-
ting lead smelters.
(b) Establish degree of water solubility of these slags when
exposed to weathering.
(c) Define acceptable treatment or disposal methods.
Investigation of the possible use of copper, lead, and zinc slags as
raw materials in cement production.
Need:
Waste slag produced by base-metal smelting constitutes a cumbersome
disposal problem. Composition of slags from lead and zinc smelters
resembles that of cement except for iron content. Slags from copper
smelters are too high in iron content for cement production, but the use
of slag in asphalt aggregate may be feasible. Suitable uses for these
slags as raw materials in other industries would eliminate disposal
problems and would maximize resource utilization.
392
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Project Scope:
(a) Determine chemical and physical properties of slags relative
to standards for construction materials.
(b) Investigate fluxing methods and additives for producing slags
with desirable chemical and physical properties.
(c) Determine whether the slags exhibit adverse solubility or
reactivity characteristics.
(d) Investiage possible treatment of slags to improve chemical or
physical properties.
(e) Compare cost of slags as raw materials with costs of materials
now used in cement production.
Evaluation of the environmental effects of the various chemicals being
used for ore flotation.
Need:
Flotation reagents in the environment produce adverse effects, such as
algae blooms. In addition, the reagents have caused ductwork fires in a
least one lead smelter. Furthermore, they discolor sulfuric acid in
subsequent smelter operations, decreasing the product's marketability.
Little is known regarding the surface kinetics and chemistry of flota-
tion separation. Other less-hazardous chemicals may offer an alterna-
tive once this operation is better understood.
Project Scope:
(a) Conduct research on the surface kinetics and flotation chem-
istry required for better understanding of the phenomenon.
(b) Test alternative reagents that could be used for flotation,
and investigate their environmental effects.
(c) Investigate alternative methods of separation such as liquid-
density methods, grain-size distribution, and settling pro-
perties.
393
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Assessment of the effects of minor wastewater streams when mixed with
mine and concentrator wastes.
Need:
It is a common practice in the mining industry to intermix most waste-
water streams for treatment and recycle. It is likely that inclusion of
certain smaller waste streams, such as sewage, boiler blowdown, and
cooling tower overflow seriously limits the reuse of this combined
wastewater and/or hinders effective treatment for removal of metallic
constituents.
Project Scope:
(a) Identify minor streams that may interfere with reuse or treat-
ment.
(b) Quantify effects of removing the detrimental minor streams
from the mixed waste streams.
(c) Define acceptable alternative methods for disposal of the
minor waste streams.
(d) Compare overall production costs of treating mixed versus
separated streams.
Development of mechanical screening and filtering equipment to separate
tailings from concentrator water.
Need:
Tailings are commonly removed from flotation water by allowing them to
settle in a pond, from which they are then removed by dredging and
earth-moving. This solid waste could be controlled more effectively by
means of a mechanical device that continuously removes the tailings.
Such a method would simplify hauling of the solids, minimize the time
solids remain in contact with water, and reduce the land area require-
ment for deposition of these wastes.
Project Scope:
(a) Establish conceptually a method for mechanically separating
mill tailings from flotation water.
(2) Develop detailed conceptual drawings for the device.
(3) Define the manufacturing and operating costs.
394
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Investigation of methods for disposition of heavy-metal sludge from
the clarification of base metal. wastewaters^
Need:
Coagulation of water has been shown to be an efficient method for
removal of heavy metals from wastewater, but no techniques have been
developed to effectively dispose of the highly metalliferous sludge that
results from these operations. Several routes have been proposed,
ranging from chemical fixing and incineration to recycling the sludge to
pyrometallurgical processing; none of these possibilities has been
developed or tested.
Project Scope:
(a) Identify possible methods for disposition of clarifier sludge
from heavy metals water treatment.
(b) Test each method and establish its effectiveness.
(c) Define relative costs of each disposal method.
Evaluation of the methods for reacidifying water from a heavy-metals
waste treatment process.
Need:
Effective treatment of wastewater for removal of heavy metals requires,
in most cases, that the water be made more alkaline than is allowable
for discharge to a public watercourse. Reacidification may redissolve
some of the precipitated heavy metals, especially if mineral acids are
being used. A study of the methods and necessary degree of reacidifica-
tion would be valuable.
Project Scope:
(a) Perform laboratory tests to establish optimum pH for removal
of heavy metals from typical base metal wastewater streams, by
treatment with lime, coagulation, and settling.
(b) Determine the effect of reacidification on suspended heavy
metals solids, using both carbonic and sulfuric acids.
(c) Establish the effects of high pH discharge on water quality of
a typical receiving body.
395
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Investigation of chelatiilg agents for the separation of heavy metals
from water solutions.
Need:
Many heavy metals are capable of forming complex molecules (chelates)
with certain organic materials. These chelates can be extracted with an
inert solvent, thus concentrating the heavy metal. Chelating chemicals
are in use that are specific to copper and are used in solvent extrac-
tion techniques for copper recovery. It is possible that other less-
specific chemicals or chemical mixtures could be effective in removing
many hazardous metals from industrial waste streams.
Project Scope:
(a) Define chemical compounds that can effectively remove heavy
metals by solvent extraction techniques.
(b) Determine by laboratory examinations effectiveness of these
compounds in removing low concentrations of heavy-metal ions
from a solution of typical industrial waste composition.
Development of an ion exchange resin that Will selectively remove toxic
metal ions from water solution.
Need:
Ion exchange techniques can effectively remove heavy metal ions from
water, but they cannot normally be applied to waste streams with high
concentrations of calcium and magnesium. Because most ion exchange
resins, being sodium based, are not selective to toxic metals, the water
must be completely softened to ensure complete removal of toxic elements.
It has been postulated that a resin based on magnesium would he insen-
sitive to the common nontoxic elements, and would therefore be selective
to iron and to metals less active than iron, including copper, lead,
zinc, cadmium, and mercury. No such resin has been developed.
Project Scope:
(a) Identify possible compounds that could act as an ion exchang-
ing substance specific to heavy metals in solution.
(b) Synthesize these materials and define their operating suita-
bility.
396
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Investigation of selenium-accumulating plants as a method of controlling
selenium'in base-metal production wastes.
Need:
There is no proven process for extracting the poisonous element selenium
from mine spoil, concentrator tailings, or wastewaters. It is known
that certain plants selectively accumulate selenium in their tissues in
concentrations often a thousand times higher than average concentrations
in their environment. Knowledge of the mechanism of accumulation might
open a route for control or recovery of this element, or the plants
themselves might be used for its accumulation.
Project Scope:
(a) Determine which plants effectively accumulate selenium in
their tissues, and establish horticultural requirements for
their growth.
(b) Grow these plants in test areas to determine their ability to
remove selenium from soil or water.
(c) Determine the mechanism by which they accumulate this element,
with a view toward artifically duplicating this mechanism.
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