October 1978
ENVIRONMENTAL ASSESSMENT OF THE
DOMESTIC PRIMARY COPPER, LEAD
AND ZINC INDUSTRIES
°PREPUBLICATION COPY0
by
PEDCo Environmental, Inc.
11499 Chester Road
Cincinnati, Ohio 45246
Contract No. 68-03-2537, Work Directive No. 1
EPA Project Officers
John 0. Burckle and Margaret J. Stasikowski
Industrial Pollution Control Division
Industrial Environmental Research Laboratory
5555 Ridge Avenue
Cincinnati, Ohio 45268
U.S. ENVIRONMENTAL PROTECTION AGENCY
Industrial Environmental Research Laboratory
5555 Ridge Avenue
Cincinnati, Ohio 45268
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NOTICE
This document has been reviewed by the Office of Research and
Development, U.S. Environmental Protection Agency, and approved for
publication as an INTERIM report. Approval does not signify that the
contents necessarily reflect the views and policies of the Environmental
Protection Agency nor does mention of commercial products constitute
endorsement or recommendation for use. This report is being circulated
for comment on its technical accuracy and policy implications. Following
receipt of these comments, 1t is planned to make appropriate changes and
publish a final report. Publication of this interim report is, therefore,
on a limited basis.
This INTERIM report is Intended to provide state-of-the-art information
on methods to assess the load of pollutants on watercourses from nonpoint
sources. The "user" should be aware that some of the technical aspects
may be changed 1n the final report. Nevertheless, this interim report
does outline the type of methods that can be used to generate the
assessments and specifies the data which are required.
11
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PREFACE
The purpose of this study has been to assemble the data needed for a
comparative assessment of all discharges (air, water, and solid waste) from
the production of primary copper, lead, and zinc in the United States.
This "multimedia" approach has resulted in a comprehensive overview that
allows the environmental problems of these industries to be addressed
collectively, rather than on an individual basis. The report outlines the
processes used for the production of these metals, and the associated
environmental discharges and control strategies. It does not include
economic analyses or detailed performance assessments of pollution control
devices and treatment technologies, as a principal objective of this effort
has been to provide the background information required to decide where
such studies are most needed. The report also identifies industrial
processes and practices that could result in more effective utilization of
this country's raw materials and energy resources. The data were assembled
largely from the open literature, with theoretical means used to fill gaps
in the data, resolve inconsistencies, and confirm literature findings. The
results of previous single-media investigations were assigned to the
appropriate categories. Emphasis was placed on fullest identification of
all polluting components from readily available information.
This report will be used as a reference by working groups within the
U.S. Environmental Protection Agency, as well as state and local control
agencies and individual researchers. It summarizes all data existing
extant in a format allowing continuous updating and further quantification
as new facts become available. The Metals and Inorganic Chemicals Branch
of the Industrial Pollution Control Division should be contacted for
additional information on this program.
m
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ABSTRACT
This report presents the results of a multimedia (air, water, and solid
waste) study of the environmental impacts of the primary copper, lead, and
zinc industries in the United States. All production processes currently
employed by these industries are identified and described, and all pollutant
effluents and environmental effects from those processes are characterized.
The various pollution control systems in use or potentially applicable are
described and evaluated for domestic application. In addition, alternate
production processes that are in use at foreign smelters or are being devel-
oped for use in this country are reviewed.
There are sixteen primary copper smelters in this country, all but two
of which are located in ore-producing regions in the West. Including mining
operations, the copper industry is the largest single employer in Arizona,
Montana, Nevada, and Utah. The principal environmental impacts of the
industry are atmospheric emissions of S02 and trace metals from the smelters.
Control of these emissions is complicated by the insufficient markets for by-
product sulfur compounds, the poor collection of some metallic fume in the
high-temperature particulate control devices often employed, and the problems
inherent in capturing "fugitive" emissions. Water pollution is most serious
at mines and concentrating plants, and includes acid drainage and trace metal
contamination. However, water problems are mitigated somewhat by the fact
that many mines and smelters are located in remote arid regions.
Lead is produced by six primary smelters, with somewhat over half of the
industry's capacity located in Missouri and the remainder in other Western
states. The Missouri smelters process ore from the "New Lead Belt" dis-
covered in that state in 1965, and a large percent of future lead supplies
are expected to originate from this region. Lead mining and concentrating
operations produce metal-laden wastewaters that are effectively controlled in
Missouri by biotic degradation and natural precipitation at high pH, although
they may be less well controlled in other mining regions. Lead smelters
release particulate matter and metal fume, and S02 emissions are currently
only about 50 percent controlled.
There are six primary zinc smelters, as well as one additional plant
that produces zinc oxide from a primary feed material. Several of these are
located in or near populated areas. A number of zinc smelters have closed in
the past decade because of environmental considerations and poor economic
prospects. The technologies now in use in the industry have facilitated
effective particulate control and reduced the potential for sulfur oxide air
pollution. The principal environmental problems are associated with waste-
water discharges resulting from gas-cleaning operations prior to sulfuric
acid production.
iv
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The report summarizes the air, water, and solid waste management prac-
tices in the primary copper, lead, and zinc industries. Programs for further
research and development are presented which are based on the gaps identified
in the data base.
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CONTENTS
Preface
Abstract
Figures
Tables
Acknowledgment
1. Introduction
2. Copper Industry
Industry Description
Industry Analysis, Copper
Process No. 1
No.
No.
No.
Process No
Process No
Process No
Process
Process
Process
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
Copper
Process No
Process No
Process No
Process No
Process No
Process No
Process No
Process No
Process No
Mining
2, Concentrating
3, Multiple-Hearth Roasting
4, Fluidization Roasting
5, Drying
6, Reverberatory Smelting
7, Electric Smelting
8, Flash Smelting
9, Peirce-Smith Converting
10, Hoboken Converting
11, Noranda
12, Electric Furnace Slag Treatment
13, Flotation Slag Treatment
14, Contact Sulfuric Acid Plant
15, DMA S02 Absorption
16, Elemental Sulfur Production
17, Arsenic Recovery
18, Fire Refining and Anode Casting
19, Electrolytic Refining
20, Electrolyte Purification
21, Melting and Casting Cathode
22, Slime Acid Leach
23, CuS04 Precipitation
24, Slimes Roasting
25, Slime Water Leach
26, Dore" Furnace
27, Scrubber
28, Soda Slag Leach
29, Selenium and Tellurium Recovery
30, Dore" Metal Separation
Process No. 31, Vat Leaching
m
iv
ix
ix
xiv
1
14
14
30
31
35
42
49
52
54
63
65
68
75
77
79
81
83
90
93
95
98
105
109
113
116
118
120
122
123
126
128
130
132
133
VI
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Process No. 32,
Process No. 33,
Process No. 34,
Process No. 35,
Process No. 36,
Process No. 37,
Process No. 38,
Process No. 39,
Process No. 40,
Lead Industry
Industry Description
Industry Analysis
Sulfide Ore Leaching
Cementation
Solvent Extraction
Electrowinning
Sulfation Roasting
Sponge Iron Plant
CLEAR Reduction
CLEAR Regeneration - Purge
CLEAR Oxidation
Process No. 1, Mining
Process No. 2, Concentrating
Process No. 3, Sintering
Process No. 4, Acid Plant
Process No. 5, Blast Furnace
Process No. 6, Slag Fuming Furnace
7, Grossing
8, 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
Process
Process
Process
Process
Process
Process
Process No
Process No
Process No
Process
Process
Process No
Process No
Process No
Process No
Process No
Zinc Industry
Industry Description
Industry Segment Analysis
No.
No.
No.
No.
No.
No.
No.
No.
11
12
13
14
15
16
17
18
19
20
21
22
No.
No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process
Process
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
Process No.
1, Mining
2, Concentrating
3, Ferroalloy Production
4, Multiple-Hearth Roasting
5, Suspension Roasting
6, Fluidized-Bed Roasting
7, Sintering
8, Vertical Retorting
9, Electric Retorting
10, Oxidizing Furnace
11, Leaching
12, Purifying
13, Electrolysis
14, Melting and Casting
Process No. 15, Cadmium Leaching
136
137
141
143
145
147
149
151
153
155
155
161
164
168
175
184
186
193
198
201
203
204
207
209
211
213
215
216
218
220
221
223
225
226
228
228
242
243
246
254
256
260
263
266
274
281
285
288
291
293
296
298
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Process No. 16, Cadmium Precipitation 301
Process No. 17, Cadmium Purification and Casting 303
5. Air Management 306
Emission Characteristics 306
Emission Control Systems 312
Fugitive Emissions 318
6. Water Management 320
Sources of Secondary Water Pollution 321
Concentrator Effluent 325
Sulfide Weathering 326
Waste Characteristics 328
Waste Handling and Treatment 329
Advanced Treatment 340
7. Solid Waste Management 346
Sources and Characteristics 346
Waste Treatment 350
8. Emerging Technology 352
9. Recommended Research and Development Programs 374
vm
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FIGURES
Number Page
1-1 U.S. Sources and Uses of Nonferrous Metals 4
1-2 Yearly Copper, Lead, and Zinc Commodity Prices, 1962-1976 5
1-3 Primary U.S. Nonferrous Smelting and Refining Locations 10
2-1 Copper Industry Flow Sheet 28
3-1 Lead Industry Flow Sheet 162
4-1 Zinc Industry Flow Sheet 240
TABLES
Number . Page
1-1 By-Product Elements of the Industries - 1975 3
1-2 By-Products - Primary Nonferrous Smelter 8
1-3 Mine Production of Copper, Lead, and Zinc by States (1976) 9
1-4 Materials Handled at Copper, Lead, and Zinc Mines (1972) 12
2-1 Copper Minerals Important in U.S. Production 15
2-2 Typical Analysis of Copper Ore Used at White Pine Copper 17
Company, Michigan
2-3 Consumption of Refined Copper in 1976 19
2-4 Statistical Data for the Primary Copper Industry in the 19
United States in 1976
2-5 Principal By-Product Sulfuric Acid Producers - 1974 20
ix
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2-6 U.S. Primary Copper Producers (Conventional Smelting/ 21
Refining Operations)
2-7 Twenty-Five Leading Copper Mines in the United States 22
in 1974
2-8 1976 Survey of Mine and Plant Expansions in the United 24
States
2-9 Annual Generation of Hazardous Pollutants from U.S. 26
Primary Copper Industry - 1968 (metric tons)
2-10 Raw Waste Load in Water Pumped from Selected Copper Mines 33
2-11 Analysis of Copper Concentrate 37
2-12 Typical Flotation Collectors 38
2-13 Metallic Elements in Concentrator Wastewater 40
2-14 Typical Size Profile of Multiple-Hearth Copper Roaster 44
Effluents
2-15 Concentration and Weight Analysis of Particulate 45
Effluents from a Multiple-Hearth Copper Roaster
2-16 Typical Levels of Volatile Metals in Domestic Copper 46
Ore Concentrations
2-17 Composition of Charge to a Reverberatory Furnace 55
2-18 Analysis of Particulates Emitted from a Reverberatory 56
Furnace
2-19 Composition of Reverberatory Furnace Exhaust Gases 57
2-20 Effluents from Slag Granulation (mg/1) 58
2-21 General Range of Reverberatory Furnace Slag Composition 60
2-22 Material Balance on Converters - Smelters in Arizona 69
(percent)
2-23 Composition of Converter Dust 70
2-24 Particle Size Distribution in Converter Dust 71
2-25 Particulate Emissions Analysis at Stack Outlet for 71
Reverberatory Furnace and Converter
2-26 . Converter Off-Gas Composition 73
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2-27 Estimated Maximum Impurity Limits for Metallurgical Off- 84
Gases Used to Manufacture Sulfur Acid (Approximate
limit, mg/Nm3)
2-28 Raw Waste Characterization: Acid Plant Slowdown 86
2-29 Acid Plant Slowdown Control and Treatment Practices 88
2-30 Analysis of Arsenic Plant Washdown Water 96
2-31 General Range Analysis of Anode Copper 99
2-32 Water Requirements for Copper Refineries 102
2-33 Waste Effluents from Anode Cooling Water 102
2-34 Contact Cooling Water Control and Treatment Practices 103
2-35 General Range Analysis of Electrolyte, Refined Copper 108
and Anode Slime
2-36 Waste Effluents from NiS04 Barometric Condenser 111
2-37 Analysis of Water Used to Cool Refinery Shapes 114
(Concentrations in mg/1)
2-38 Dore" Metal Analysis 124
2-39 Analysis of Tailings Effluent from a Precipitation Plant 140
3-1 Lead Minerals, By Name, and Composition 156
3-2 Principal Statistics for the Primary Lead Industry in 158
the United States in 1976
3-3 Domestic Primary Lead Producers 159
3-4 Analysis of a Missouri Mine Water 167
3-5 Analysis of an Idaho Mine Water 167
3-6 Typical Southeastern Missouri Lead Concentrate Analyses 169
(Percent by weight)
3-7 Western Lead Concentrate Analyses 170
3-8 Flotation Chemicals 172
3-9 Lead Mill Wastewater Analysis 173
3-10 Sinter Analysis 176
XI
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3-11 Sinter Machine Feed 177
3-12 Grain Loading and Weight Analysis of Input Feed and 173
Emissions, Updraft Lead Sintering Machine
3-13 Typical Size Profile of Emissions, Updraft Lead Sinter- 178
ing Machine
i
3-14 Analysis of Sinter Machine Exhaust Gases (Missouri Lead 180
Operating Company)
3-15 Atmospheric Control Systems on Primary Lead Sintering 181
Machines
3-16 Wastewater Treatment at Primary Lead Acid Plants 185
3-17 Scrubber Wastewater Treatment at Primary Lead Plants 185
3-18 Lead Bullion Composition 187
3-19 Typical Blast Furnace Slag Analysis 188
3-20 Typical Blast Furnace Charge 189
3-21 Exhaust Gas Analysis After Air Dilution and CO Combustion 190
3-22 Atmospheric Control Systems on Primary Lead Blast 192
Furnaces
3-23 Waste Effluents from Slag Granulation 194
3-24 Effluent Concentrations with Neutralization and 195
Clarification
3-25 Primary Lead Slag Granulation Wastewater Treatment 196
3-26 Lead Bullion Analysis (Basis: As Drossed) 199
3-27 Typical Compositions of Softened Lead Bullion and Slag 205
(Amounts in weight percent)
3-28 Typical Retort Analysis 215
4-1 Mining, Production, and Consumption of Zinc and Cadmium 229
(metric tons)
4-2 Twenty-Five Leading Zinc Mines in the United States 230
4-3 Common Ores Mined for Their Zinc Content 233
4-4 U.S. Slab Zinc Consumption - (1976) 235
xn
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4-5 Grades of Commercial Zinc 235
4-6 Primary Zinc Processing Plants in the United States 236
4-7 Principal Statistics for the Primary Zinc Industry in 238
the United States - 1976
4-8 Range of Compositions of Zinc Concentrates 247
4-9 Typical Flotation Reagents Used for Zinc Concentrations 249
4-10 Ranges of Constituents of Wastewaters and Raw Waste Loads 251
for Five Selected Mills
4-11 Product Sinter Composition (percent) 267
4-12 Particulate Emissions from a Zinc Sintering Process 270
4-13 Particulate Emissions from a Zinc Coking Furnace 271
4-14 Uncontrolled Particulate Emissions from a Zinc Vertical 278
Retorting Furnace
4-15 Analysis of Vertical Retort Furnace Residues 279
4-16 Zinc Oxide Impurities and Brightness 286
4-17 Selected Constituents of Oxidizing Furnace Residue 287
4-18 Analyses of Anode Sludges from Electrolytic Zinc Refining 295
5-1 Sources of Atmospheric Emissions from Mining, Concen- 307
trating, and Smelting Operations
5-2 Elemental Analysis of Particulate Size Fractions 315
5-3 Selected Trace Element Emissions in Particulate Matter 318
From ESP on a Copper Smelter Reverberatory Furnace
6-1 Sources of Liquid Effluents from Mining, Concentrating, 322
and Smelting Operations
6-2 Comparison of Tailings Pond Reservoir Linings 333
6^3 Metal Concentrations in Effluent from a Mine Drainage 336
Clarifier
7-1 Sources of Solid Waste From Mining, Concentrating, and 347
Smelting Operations
xm
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ACKNOWLEDGMENT
This report was prepared by PEDCo Environmental, Inc., under the direc-
tion of Mr. Timothy W. Devitt. The PEDCo project managers were Mr. Thomas
K. Corwin and Dr. Gerald A. Isaacs. Principal authors were Mr. Thomas K.
Corwin, Mr. Hal M. Drake, Dr. Gerald A. Isaacs, Mr. Douglas J. Morell,
and Mr. A. Christian Worrell, III.
Project officers for the Industrial Environmental Research Laboratory of
the U.S. Environmental Protection Agency were Mr. John 0. Burckle and Ms.
Margaret J. Stasikowski.
The authors appreciate the efforts and cooperation of everyone who
participated in the preparation of this report.
xiv
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SECTION 1
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 multimedia 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 significant 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. All production processes, unit operations, and
pollutant discharges were examined in detail to characterize their environ-
mental impacts completely and accurately. Emphasis was placed on fullest
identification of all polluting components from readily available informa-
tion, with the results of previous single-media investigations assigned to
the appropriate categories. In some cases, engineering studies were under-
taken to describe sources of pollution and to predict the quantity and
character of discharges based upon the materials processed, production capac-
ities, and process conditions. As a result of certain limitations on the
amount of information available, some important questions have been left
unanswered. However, the report constitutes a comprehensive multimedia
environmental overview, which identifies these information gaps as well as
research and development efforts needed for more effective 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 concentrating 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 there are companies that produce two, or
all three, of the metals. The industries share similar production tech-
niques and produce similar waste materials.
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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 fabrica-
tion 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 metals, and it
is the principal material used for storage batteries and for radiation
shielding. Zinc is most used as a base for die casting alloys, as galva-
nizing 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 many by-products. 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 primary 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 1976, 1.8 million metric tons of copper were used, com-
pared to 1.0 million metric tons of zinc and 1.1 million metric tons of lead
(1). Figure 1-1 provides information on sources and uses of copper, lead,
and zinc from 1950 to 1976. 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 2,000 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
has been in additives for motor fuel; this use is being phased out. Plastic
coatings are replacing some of the corrosion-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. One 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.
Commodity price trends for the past 10 years are shown in Figure 1-2.
They show an increase in price typical of other materials, but they also show
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TABLE 1-1. BY-PRODUCT ELEMENTS OF THE COPPER, LEAD, AND ZINC INDUSTRIES - 1975
By-product
Principal source
Approximate quantity
Remarks
co
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
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t-r-J DOMESTIC MINES jilll NET IMPORTS
m
f$$\ OLD SCRAP
OTHER 5%
FOUNDRIES 8%
BRASS 29%
ELECTR'CAL 58%
0 C -T-TV-T— t-~-1—~-f - >- -t — i—I"- <—i— -t- —i-" t~— i— -t — i—I -'1- -i- 4- -i— + -~t- i
1950 1955 1960 1965 1970 1976 1950 CURRENT
OTHER 5%
PIGMENT 6%
GASOLINE 18%
FABRICATED 18%
BATTERIES 53%
1950
1955
1960
1965
1970
1976 1950 CURRENT
OTHER 4%
OXIDES 11%
BRASS 14%
ALLOYS 35°/.
GALVANIZING 36%
1950
1955
1960
1965
1970
1976 1950 CURRENT
Figure 1-1. U.S. sources and uses of nonferrous metals (2).
(Million of tons)
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80
70
LU
o 60
S 50
o.
o
o 30
o
20
10
I
I
I
I
'62 '63 '64 .'65 '66 '67 '68 '69 '70 '71 '72 '73 '74 '75 '76
YEAR
Figure 1-2. Yearly copper, lead, and zinc commodity prices, 1962-1976.
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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 pro-
cessed.
The United States is almost self-sufficient in its smelter production 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 sec-
ondary 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 the manufacturing facilities of these industries.
Table 1-2 lists these by-products. One of the most important by-products 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; trans-
portation costs make it less expensive for the user to buy sulfur for on-site
manufacture of the acid. 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 1-3 shows the location of the 28 copper, lead, and
zinc smelters in the United States, and of the seven copper refineries.
Copper is refined in a separate, but often adjacent plant, whereas lead and
zinc smelters more typically include 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 produce 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. Table 1-4 indicates
the scale of these operations in terms of materials handled. Arizona mines
produce more than half the copper, and Missouri mines 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.
Many of the smelters and refineries in the United States, especially in
the copper industry, are old plants, some having begun operations 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 concentrates for a fee and at a very low profit. A search for
new, more efficient, processes is under way.
-------
TABLE 1-2. BY-PRODUCTS - PRIMARY NONFERROUS SMELTERS
Company
Location
Primary
product
By-products
AMAX, Inc.
The Anaconda Co.
ASARCO, Inc.
The Bunker Hill Co.
Cities Service Co.,
Inc.
Copper Range Co.
Inspiration
Consolidated
Copper Co.
East St. Louis, Illinois
Boss, Missouri
Anaconda, Montana
Tacoma, Washington
Hayden, Arizona
El Paso, Texas
East Helena, Montana
%
Glover, Missouri
Corpus Christi, Texas
Columbus, Ohio
Kellogg, Idaho
Copperhill, Tennessee
White Pine, Michigan
Miami, Arizona
Zinc
Lead
Copper
Copper
Copper
Copper, lead
Lead
Lead
Zinc
Zinc oxide
Lead, zinc
Copper,
Iron oxide,
Sulfuric acid
Copper
Copper
Sulfuric acid, zinc sulfate, cadmium
Sulfuric acid
Sulfuric acid, beryllium
Arsenic, arsenic trioxide, sulfur dioxide,
sulfur sulfuric acid, gold, silver
Sulfuric acid
Sulfuric acid, gold, silver
Zinc fume, sulfuric acid, soda ash matte,
soda ash speiss
Copper dross
Sulfuric acid, zinc sulfate, cadmium
Sulfuric acid
Sulfuric acid, zinc oxide, cadmium, copper
matte, copper residue, refined gold dor£,
refined silver, silver concentrates
Copper carbonate, copper sulfate, iron
sulfate (ferric), sodium hydrosulfite,
sulfur trioxide, surface-active agents,
p-toluenesulfonic acid, emulsifiable
liquid copper fungicide, chelated plant
and animal nutrients
Silver
Sulfuric acid, molybdenum, gold, silver,
selenium
-------
TABLE 1-2 (continued).
Company
Location
Primary
product
By-products
Kennecott Copper
Corp.
Magma Copper Co.
National Zinc Co.
New Jersey Zinc Co.
Phelps Dodge Corp.
St. Joe Minerals
Corp.
Garfield, Utah
Hurley, New Mexico
Hayden, Arizona
McGill, Nevada
San Manuel, Arizona
Bartlesville, Oklahoma
Palmerton, Pennsylvania
Morenci, Arizona
Douglas, Arizona
Ajo, Arizona
Hidalgo, New Mexico
Herculaneum, Missouri
Monaca, Pennsylvania
Copper
Copper
Copper
Copper
Copper
Zinc
Zinc
Copper
Copper
Copper
Copper
Lead
Zinc
Ammonium perrhenate, molybdenum disulfide,
molybdenum trioxide, nickel sulfate,
sulfuric acid, gold, silver, selenium
Sulfuric acid, molybdenum
Sulfuric acid
Molybdenum
Sulfuric acid, molybdenum, gold, silver
Ammonia, carbon dioxide, sulfuric acid,
zinc oxide, cadmium, metal powders,
spiegeleisen, zinc alloys
Sulfuric acid, zinc oxides, cadmium, lead
sulfate, spiegeleisen
Sulfuric acid, gold, silver
Gold, silver
Sulfuric acid,
Sulfuric acid,
gold, silver
elemental sulfur
silver bullion, copper
Sulfuric acid,
matte
Sulfuric acid, zinc oxide, cadmium,
ferrosilicon, mercury
-------
• COPPER SMELTER
«COPPER REFINERY
* COPPER SMELTER/REFINERY
•0 LEAD SMELTER
0 LEAD REFINERY
•f LEAD SMELTER/REFINERY
« ZINC SMELTER
Figure 1-3. Primary U.S. nonferrous smelting and refining locations.
-------
TABLE 1-3. MINE PRODUCTION OF COPPER,
LEAD, AND ZINC BY STATES (1976)
Alaska
Arizona
California
Colorado
Idaho
Maine
Michigan
Missouri
Montana
Nevada
New Jersey
New Mexico
New York
Pennsylvania
Tennessee
Utah
Virginia
Other states
Totals
Metric tons as 100% of the metals
Copper
929,339
340
2,205
3,050
1,602
39,650
10,024
82,655
52,762
156,362
10,098
168,245
230
11,456,562
Lead
13
307
49
24,266
48,658
196
454,491
83
528
2,899
14,784
1,765
4.9313
552,970
Zinc
8,619
154
45,923
42,262
7,085
75,777
58
1,304
30,363
66,833
20,212
74,854
20,394
10,197
35,236b
439,271
Includes Illinois, New Mexico, Oklahoma, Washington,
and Wisconsin.
Includes Illinois, New Mexico, Washington, and
Wisconsin.
Source: U.S. Bureau of Mines.
11
-------
TABLE 1-4. MATERIALS HANDLED AT COPPER, LEAD, AND ZINC SURFACE AND UNDERGROUND
U.S. MINES IN 1975 (MILLIONS OF METRIC TONS) (3)
Ore
Copper
Lead
Zinc
Total
Surface
Crude ore
217.68
-
0.07
217.75
Waste
624.92
-
0.04
624.96
Total9
841.70
-
0.11
841.81
Underground
Crude ore
26.48
8.93
7.70
43.11
Waste
1.23
2.22
2.49
5.94
Total8
27.75
11.16
10.16
49.07
All mines
Crude ore
243.98
8.93
7.78
260.69
Waste
625.83
2.22
2.52
630.57
Total3
869.81
11.16
10.34
891.31
ro
Data may not add to totals shown because of independent rounding.
-------
References
1. Commodity Data Summaries 1977. U.S. Department of the Interior. Bureau
of Mines. Washington, D.C. 1977.
2. Status of the Mineral Industries 1977. U.S. Department of the Interior.
Bureau of Mines. Washington, D.C. 1977.
3. Minerals Yearbook 1975. U.S. Department of the Interior. Bureau of
Mines. Washington, D.C. 1975.
13
-------
SECTION 2
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
changes will or should take. Whether the trend is toward improved pyrometal-
lurgical processing or toward adoption of hydrometallurgy, most experts agree
that some basic changes are imminent.
At six installations, newer technology has been or is now operating.
One new smelter now uses a continuous flash smelting process. A continuous
smelting process is beginning to produce copper in Utah. One installation
has produced copper with a roast-leach-electrowinning technique. Three
advanced hydrometallurgical processes are approaching semi commercial produc-
tion. The following description of the industry 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 "conventional"
pyrometallurgical methods that center on the energy-inefficient reverberatory
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 about 25 locations, all but
five of which were operating before World War II. In twelve of these loca-
tions, 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 gangue rock. Throughout the
world, copper in minerals is most often chemically combined with sulfur,
frequently 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 oxidize to form
sulfuric acid and metal ions. The metal ion may, in turn, react with rock
minerals to form metal oxides, or they may move with ground or surface waters
14
-------
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
Cu5FeS.
CuS
Cu3As5S4
CuC03-Cu(OH)2
2 CuC03-Cu(OH)2
Cu20
CuS03-2H20
Cu
Copper content,
percent weight
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.
15
-------
and subsequently precipitate to form secondary metal deposits. Weathering
may therefore create deposits of oxidized copper minerals. The table shows
four of these, the highly colored azurite and malachite being most abundant
in domestic ores.
The ore deposits of northern Michigan are a unique occurrence of primary
origin, and native copper mixed with sulfide minerals is mined in this area.
Table 2-2 gives an analysis of the ore from this deposit. Except for
the presence of elemental copper and the low-sulfur 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 concen-
trate, 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, concentrates 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 inter-
mediate 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.
Primary copper smelting and refining, together with the primary aluminum
industry, accounts for 3 percent of all energy consumed by manufacturing
industries in the U.S. (2). Currently, natural gas is the fuel most heavily
relied upon by copper smelters, with oil also being widely used. Energy
consumption in 1974 for all copper,Defining and smelting operations,,™ the
U.S. was: Natural gas - 1.52 x 10 kilocalories, Oil - 4.13 x 10 kilo-
calories, and electricity - 6.82 x 10 kilowatt-hours (2). The rapidly in-
creasing cost of these forms of energy, and their increasing unavailability,
is causing consideration of conversion to coal; however, coal is currently
limited to only one or two locations.
Products
Commerce recognizes a number of different grades of copper, classified
into two main groups. Relatively impure grades are directly produced in a
copper smelter. These are sold for use in alloys or for other special pur-
poses, or they may be exported to be refined elsewhere. More than 90 percent
16
-------
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 (Weight)
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
17
-------
of the copper produced is refined in this country into one of the electro-
lytic grades. Table 2-3 shows the distribution of consumption of electro-
lytic copper in 1976. More than two-thirds of it is used directly to manu-
facture 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.
Several by-product elements are isolated from the copper ores. In 1976
the copper industry produced all the arsenic and selenium manufactured in
this country, almost all the platinum and palladium, and almost half the
gold, silver, and" polybdenum. 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 sulfuric acid as a major product, and most
others have facilities to manufacture 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 1976 statistics of this industry, and Table
2-5 lists major sulfuric acid producers.
Companies
The United States is the world's largest copper producer, accounting for
20 percent of the total world production in 1976. Domestic mine output that
year was estimated at 1.5 million metric tons and valued at $2.3 billion (4).
In 1976 nine companies operated 16 primary smelters and 20 companies
operated 27 refineries and electrowinning plants (4). Table 2-6 lists some
of these companies, with applicable data. The three largest domestic pro-
ducers 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 25 largest copper mines
operating in 1974. Twenty of these were directly owned by a producing
company. At least three other large companies own mines or leaching operations
intended primarily for production 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 a few
smaller mines, or as by-products of other mining industries.
Several of the larger U.S. copper companies have invested substantially
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.
Other projected changes through 1985 include addition of new capacity
and improvement of pollution control. Technology should not change radically
18
-------
Table 2-3. CONSUMPTION OF REFINED COPPER
IN 1976 (3)
Consumer
Quantity,
metric tons
Wire mills
Brass mills
Secondary smelters
Chemical plants,
foundries, and
miscellaneous plants
1,221,195
521,515
2,814
32,659a
Estimated.
TABLE 2-4. STATISTICAL DATA FOR THE PRIMARY COPPER INDUSTRY
IN THE UNITED STATES IN 1.976 (3)
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
1,461.81
1,325.65
1,288.77
105.76
154.20
102.40
487.34
348.37
19
-------
TABLE 2-5. PRINCIPAL BY-PRODUCT
SULFURIC ACID PRODUCERS - 1974 (5)
Producer
Capacity,
metric ton/year
The Anaconda Co.
Anaconda, Montana
ASARCO, Inc.
Corpus Christi, Texas
El Paso, Texas
Hayden, Arizona
Tacoma, Washington
Cities Service Co.
Copperhill, Tennessee
Inspiration Consolidated Copper Co.
Inspiration, Arizona
Kennecott Copper Corp.
Hurley, New Mexico
Hayden, Arizona
Garfield, Utah
Magma Copper Co.
San Manuel, Arizona
Phelps Dodge Corp.
Ajo, Arizona
Hidalgo, New Mexico
Morenci, Arizona
210,000
104,000a
145,000
163,000
49,000
l,143,000b
397,000
181,000
249,000
544,000
803,000
91,000
524,000
544,000
aASARCO - Corpus Christi smelter is no longer operating.
Approximate composition is 5% smelter gases; 95% gases from pyrites.
20
-------
TABLE 2-6. U.S. PRIMARY COPPER PRODUCERS (6,7)
(Conventional Smelting/Refining Operations)
Company
AMAX, Inc.
The Anaconda Company
ASARCO, Inc.
Cerro Corporation
Cities Service Company
Copper Range Company
Inspiration Consolidated
Copper Company
Kennecott Copper
Corporation
Magma Copper Company
Phelps Dodge Corporation
Southwire Company
Location
Carteret, New Jersey
Anaconda, Montana
Great Falls, Montana
Tacoma, Washington
El Paso, Texas
Hayden, Arizona
Amarillo, Texas
St. Louis, Missouri
Copperhill, Tennessee
White Pine, Michigan
Miami, Arizona
Garfield, Utah
Hurley, New Mexico
Hayden, Arizona
McGill, Nevada
Baltimore, Maryland
Magna, Utah
San Manuel, Arizona
Morenci, Arizona
Douglas, Arizona
Hidalgo, New Mexico
Ajo, Arizona
El Paso, Texas
Laurel Hill, New York
Carrollton, Georgia
Description
Refinery
Smelter
Refinery
Smelter/ refinery
Smelter
Smelter
Refinery
Refinery
Smelter
Smelter/refinery
Smelter/refinery
Smelter
Smelter/refinery
Smelter
Smelter
Refinery
Refinery
Smelter/refinery
Smelter
Smelter
Smelter
Smelter
Refinery
Refinery
Refinery
Capacity,
metric ton/year
(Cu content)
236,000
180,000
229,000
91,000/142,000
104,000
163,000
381,000
236,000
20,000
82,000/82,000
64,000/136,000
254,000
73,000/93,000
73,000
45,000
250,000
169,000
181,000/181,000
161,000
115,000
91,000
64,000
404,000
83,000
65,000
Note: Refineries typically produce copper from both blister and scrap in varying
proportions, and for this reason the U.S. Bureau of Mines does not categorize
refineries as either "primary" or "secondary." In general, refineries located
in Western states or adjacent to primary smelters process chiefly a blister
feed, while those refineries in the East produce a higher proportion of copper
from scrap.
21
-------
TABLE 2-7. TWENTY-FIVE LEADING COPPER MINES IN THE UNITED STATES IN 1974 (4)
Rat*
1
2
3
4
5
6
7
8
9
10
11
12
13
Mine
Utah Copper
San Manuel
Mo rend
Butte oper-
ations
Tyrone
Ray Pit
Plma
White Pine
Slerrlta
Chi no
Twin Buttes
New Cornelia
Inspiration
County
and state
Salt Lake, Utah
Final, Arizona
Greenlee, Ariz-.
Silver Bow, Mont.
Grant. N. Mexico
Plnal, Arizona
Plma, Arizona
Ontanagon, Mich.
Pima, Arizona
Grant, N. Mexico
Plma, Arizona
Plma, Arizona
Glla, Arizona
Company
Kennecott Copper Corp.
Magma Copper Co.
Phelps Dodge Corp.
The Anaconda Co.
Phelps Dodge Corp.
Kennecott Copper Corp.
Cyprus Pima Co.
White Pine Copper Co.
Duval Slerrlta 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
0.87'
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
over 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-
furlc 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
Chalcopyrite
Copper sulfides
Copper sulfides
Chalcopyrite
Porphyry copper
Copper sulfides
Chalcoclte. chryso-
colla, mlachlte.
azurite
ro
-------
TABLE 2-7 (continued).
Rank
14
15
16
1.7
IB
19
20
21
22
23
24
25
Mine
Mission
Nevada Mines
Yerington
Silver Bell
Copper Cities
Mineral Park
Superior Oiv.
Copper Queen
Continental
Bagdad
Esperanza
Battle Moun-
tain
Property
County
and state
Pima, Arizona
White Pine. Nev.
Lyon, Nevada
Pima, Arizona
Glla, Arizona
Hohave, 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.
ASARCO
Cities Service Co.
Ouval Corp.
Hagna Copper Co.
Phelps Dodge Corp.
UV Industries. Inc.
Cyprus Mines Corp.
Duval Corp.
Duval Corp.
Opera-
tion
C
0
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,
metric 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
Chal copy rite.
molybdenite
Chalcoclte. chalflo
pyrite
Chalcopyrite
Copper, molyb-
denum, silver
Chalcopyrite,
bornite, chal-
cocite
Copper sul fides
and oxides
Porphyry
Copper, gold.
silver
ro
co
A - Mine, concentrator, smelter, refinery
B - Mine, concentrator, smelter
C - Mine, concentrator
0 - Mine, concentrator, smelter, ore leach plant/precipitate plant
£ - Mine, concentrator, hydrometallurgical recovery
F - Mine, concentrator, smelter, refining, Isiching-electro mining plant
G - Mine, concentrator, leach plant
-------
TABLE 2-8. 1976 SURVEY OF MINE AND PLANT EXPANSIONS IN THE UNITED STATES3
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 PeakvAr1z.,
Bagdad, Arizona,
S1err1ta. Ariz..
San Manuel, Ariz.,
RMnelander, Wise.,
Safford, Arizona,
Project
UGmlco
ml/Co
ml/Co
UGmlsm
ex
ex
Pi
ml
OPmt
UGmlco
Planned
10M
120M
9.9M
2M
2M
70M
22. 5M
75M
350M
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.S9MM t 0.351 Cu. 0.041 molyb-
denite
OR 50MM t 1.9SX Cu ore. May be some small
production in 1976.
OR 12MM t 2.21 Cu. Operation expects a 20-
year Hfespan.
OR 300MM t 0.49X Cu. Smelter on site post-
poned.
Chloride leach process recovers electrolytic
copper.
Cu/Zn mine. Dependent on environmental
clearance.
OR 400MM tons of 0.72X Cu. Indefinitely
postooned.
ro
Abbreviations:
U6 - underground tpd - tons per day; tpy - tons per year
•1 - Mine tm - smelter
CO - concentrator ex - complex
pi - plant
8 Data from U.S. Bureau of Mines.
-------
due to the lead time required, the lack of recent innovations, and the
capital intensive nature of the industry (2).
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 concentrating mills. The copper in-
dustry is the largest single employer in Arizona, Montana, Nevada, and
Utah.
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,
especially those that are potentially hazardous to human health. Among the
known trace elements of concern are arsenic and cadmium. 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 recover some of the S02 as a
marketable product. Fifteen percent of the S02 generated by the industry is
fugitive emissions.
The industry makes extensive use of water recycle techniques. 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. Slag from the
industry, which is dumped, contains many elements. Table 2-9 presents ap-
proximate quantities of selected pollutants from the U.S. 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. Smelting followed by electrolytic re-
fining produces about 2.4 kilograms of wastes per metric ton of product.
In 1974 the copper ore mining and concentrating industry produced about
651 million metric tons of solid waste, or about 85 percent of the national
total for metals mining and concentrating. Of this figure, 56 percent was
waste rock, 7 percent was overburden, and 36 percent was concentrator
waste (7).
References
1. Mining Informational Services of the McGraw-Hill Mining Publications.
1975 E/MJ International Directory of Mining and Mineral Processing
Operations. McGraw-Hill, Inc., 1975.
25
-------
TABLE 2-9. ANNUAL GENERATION OF SELECTED POLLUTANTS FROM U.S. PRIMARY COPPER INDUSTRY - 1968
(metric tons)
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
Cadmi urn
Neg.
189
77
222
49
537
Percent
of total*
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 total3
0.13
0.05
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
5
43
Percent
of total3
1.99
0.94
2.57
0.59
6.09
ro
a Percent of total of this pollutant.
Combined total from copper, zinc and lead mining.
-------
2. Energy Penalty Study of the Nonferrous Metals Industry. Arthur D.
Little, Inc. Draft Final Report to Policy Planning Division. U.S.
Environmental Protection Agency. Washington, D.C. August 1977.
3. U.S. Department of the Interior. Bureau of Mines. Mineral Industry
Surveys, Copper in 1976. Washington, D.C. April 15, 1977.
4. U.S. Department of the Interior. Bureau of Mines. Commodity Data
Summaries 1977. Washington, D.C. 1977.
5. Stanford Research Institute. 1977 Directory of Chemical Producers,
United States of America. Menlo Park, California.
6. U.S. Department of the Interior. Bureau of Mines. Copper - 1977.
MCP-3. Washington, D.C. June 1977.
7. Development Document for Interim Final Effluent Limitations Guidelines
and Proposed New Source Performance Standards for the Copper Segment of
the Nonferrous Metals Manufacturing Point Source Category. EPA
440/1-75/032b. U.S. Environmental Protection Agency. Washington, D.C.
November 1974.
27
-------
PYRITE
CHARCOAL 6 CARBON
SMELTING DEPARTMENTS
I "ATER
?A,R
I SOLID
Figure 2-1. Copper industry flow sheet.
28
-------
Figure 2-1. Copper industry flow sheet.
29
-------
INDUSTRY ANALYSIS
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 and their impacts
on the atmosphere are considered especially important.
This industry analysis examines each individual production operation,
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 fol-
lowing 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 are currently
being demonstrated on a large scale at a U.S. facility. Processes believed
to be primarily in the stage of engineering development have been excluded
from this listing. Many of these are considered in Section 8, Emerging
Technology. Figure 2-1 is a flowsheet showing these processes, their inter-
relationships, and their major waste streams.
30
-------
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 (1).
In an open pit mine, holes for placement of explosives are drilled be-
hind 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 transportation 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 some-
times 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 throughout a rock composed of other
minerals. The copper content is between 0.6 and 2 percent (2). The copper
ores of southwestern United States are from porphyry deposits.
Explosives used in copper mining are almost always a mixture of ammonium
nitrite and fuel oil (AN-FO). Some mines add sodium nitrate to make the
explosive slightly more powerful (3). No values are available for the con-
sumption 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 under-
ground mines where more constant temperatures prevail.
4. Utilities - In most mines, electrically operated power equipment is used
for drilling, loading, and hauling. In 1973, 1.02 kilowatt-hours of elec-
tricity was consumed by the mining process per kilogram of copper produced
(4).
A small amount of water is needed for equipment cooling, drill lubrica-
tion, 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.
31
-------
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.
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.
Improper backfill operations may result in acid drainage. The amount of
wastewater from open pit copper mines range 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 (6). Chemical charac-
teristics 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.
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 (7).
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.
Concentrations of most materials do not exceed background levels (8).
6. Control Technology - The only control provided for fugitive dust is the
manual use of water sprays, to be used when needed. Dust from blasting can
be controlled by proper blast design. 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 the dump areas (9).
Mine water wastes and seepage from the spoil dumps are major potential
sources of water pollution from the primary copper industry. Although treat-
ment 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
32
-------
TABLE 2-10. RAW WASTE LOAD IN WATER PUMPED FROM SELECTED COPPER MINES (7)
Parameter
Flow
PH
IDS
TSS
011 & grease
TOC
COD
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
N1
Mo
Sr
Hg
Pb
Underqround mine
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
Raw waste load per unit ore mined
kq/1000 metric tons
17.28 m3/1000 metric tons
7.37a
5,053.9
11.9
<0.173
<0.778
141.5
' 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
(mg/a)
409 m'Vday
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 load per unit ore mined
kg/1000 metric tons
75 m*/1000 metric 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
co
CO
aValue In pH units
-------
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, the practice of leaching low-grade ore is being
practiced. This practice and its effect in the control of mining wastewaters
is described in the section outlining Process No. 32, Sulfide Ore Leaching.
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. EPA-R2-73-274a. U.S. Environmental Protection Agency,
Washington, D.C. September 1973.
3. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 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. PB-210 679, PB-210 478, and PB-210 677. U.S.
Environmental Protection Agency. Research Triangle Park, North
Carolina. May 1972.
6. 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. EPA/1-75/032-6. Environmental Protection Agency, Wash-
ington, D.C. February 1975.
7. Minerals Yearbook. U.S. Department of the Interior, Bureau of
Mines., Washington, D.C. 1973.
8. A Study of Waste Generation, Treatment and Disposal in the Metals
Mining Industry. PB 261-052. Midwest Research Institute for
Environmental Protection Agency. Washington, D.C. October 1976.
9. Dayton, S. The Quiet Revolution in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975.
34
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PRIMARY COPPER PRODUCTION PROCESS NO. 2
Concentrating
1. Function - Sulfide ore from the mine is separated by the concentration
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 stages, the ore 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 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 flotation stages in order to accom-
plish 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
reportedly pine oils, cresylic acid, or long-chain alcohols (1). 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. In
this application it is often used in conjunction with cyanide (1). Various
xanthates or dithiophosphates act as collectors for the valuable sulfide
minerals, and the copper and other recoverable minerals come off with the
froth (2). The gangue does not float and is discarded as "tailings".
After initial separation, the valuable minerals are sent through
stages that further separate them by selective or differential flotation.
By use of proper collectors or depressants, the concentrates may be up-
graded 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; to separate
this fraction, the concentrate must be steam stripped to remove the
collector originally added (3).
Occasionally, a concentrator will batch-treat a copper concentrate with
cyanide to dissolve its silver and gold content. After separating the leached
35
-------
solution from the concentrate, zinc metal is added to reprecipitate the
precious metals.
Concentrates are dewatered by clarification and filtration. They may
be partially dried to simplify handling and shipment, or may be more com-
pletely dried for direct "green" feed to a smelting furnace (see Process No.
5). Ten of the sixteen conventional smelters in this country have concen-
trator plants on-site or nearby (4).
Table 2-11 shows typical composition of copper concentrates; composi-
tions vary with the character of the ore and the amount of processing em-
ployed.
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. 31).
Lime is used for pH adjustment and as a pyrite depressant. Quantities
added vary between 0.9 and 18.0 kilograms per metric ton of ore processed
(5). 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 or cresylic acid required if they are substituted for pine
oil.
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 compounds.
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 temperatures may
approach 100°C (i.e., steam stripping for molybdenite concentration). Occa-
sionally 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 (4).
In the same year, concentrators consumed about 6630 million kilowatt-hours
of electricity, which is about 400 kilowatt-hours per metric ton of
primary refined copper (2,6). The greater part of this electricity was
used to operate the crushing and grinding equipment, with a smaller amount
for production of compressed air.
5. Waste Streams - The handling and milling of dry ore is the principal
source of air pollutants in this process. Items of equipment are always
enclosed, but transitions between pieces of equipment are difficult to seal
tightly. Ore classifiers are not always completely sealed. Dust quantity is
reported as about 1 kilogram per metric ton of ore (3).
36
-------
TABLE 2-11. ANALYSIS OF COPPER CONCENTRATE (7)
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,
% weight
20 - 50
30 - 38
tr. - 0.67
20 - 30
0.2 - 4.0
0.13
31.53
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 (2)
Type
Formula0
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
(C6H5NH)2CS
RCOONa
RS03NA
ROS03Na
RNH3C1
RN(CH3)3C1
RC5H4N-HC1
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
-------
This process produces the largest amount of wastewater in the industry.
The ore flotation water is used to.sluice the tailings into a pond, and
suspended solids are the most critical pollutant in concentrator effluent
(8). 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 wastewater from this process will equal the
water consumption, ranging from 100 to 500 cubic meters per metric ton of
concentrate (3). In many cases in the Southwest, evaporation may equal
consumption.
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 milligram per liter of cyanides, and ranges of metallic elements indi-
cated in Table 2-13 (3).
The water may also contain thiosulfates and thionates, and the materials,
both inorganic and organic, used as flotation additives.
More than 95 percent of all the ore brought from the mine is discharged
from this process as tailings; this quantity totals approximately 241 million
metric tons of waste material each year from the industry (2,6,9). 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 much of the iron pyrite. Production of higher-grade concentrates
to minimize air pollution has increased the proportion of pyrites in the
tailings. In this solid waste, the minerals have been pulverized and inti-
mately 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.
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 economically justifies 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 controlled 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 mate-
rials; each concentrator plant requires separate study. Problems of sec-
39
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TABLE 2-13. METALLIC ELEMENTS IN CONCENTRATOR WASTEWATER (3)
Element
Arsenic
Antimony
Cadmium
Copper
Cobalt
Iron
Manganese
Mercury
Molybdenum
Nickel
Lead
Selenium
Silver
Strontium
Zinc
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
0.01 to 3
0.003 to 0.02
0.1 approximately
0.03 to 2.5
0.05 to 8.50
40
-------
ondary 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 *" H--
1. Hawley, John R. The Use, Characteristics and Toxicity of Mine-Mill
Reagents in the Province of Ontario. Ontario Ministry of the
Environment. Toronto, Ontario. 1977.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. Environmental Protection Agency,
Washington, D.C. September 1973.
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. EPA/1-75/032-6. U.S. Environmental Protection Agency,
Washington, D.C. February 1975.
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. EPA 440/1-75/032-a. U.S. Environmental Protection
Agency, Washington, D.C. February 1975.
6. Dayton, J. The Quiet Revolution in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975.
7. Little, A.D. Economic Impact of New Source Performance Standards
on the Primary Copper Industry: An Assessment. C-76072-20. U.S.
Environmental Protection Agency, Washington, D.C. October 1974.
8. Williams, Roy E. Waste Production and Disposal in Mining, Milling,
and Metallurgical Industries. Miller Freeman Publications, Inc.
San Francisco. 1975.
9. 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
Multiple-Hearth Roasting
1. Function - Roasting is frequently the first of the pyrometallurgical
processes applied to the copper ore concentrate at the copper smelter. The
purpose of roasting is to reduce the sulfur content so that subsequent pro-
cesses operate efficiently. Roasting also removes the water from the con-
centrate, volatilizes some of the arsenic and antimony, and preheats the ore
before it is charged as calcined feed to the reverberatory furnace.
Roasting is accomplished either with a multiple-hearth or fluidization
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 preferentially 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 and other elements in the ore, and these "fumes"
leave the roaster with the SO^ gas.
Multiple-hearth roasting is currently in use at four domestic copper
smelters. The roasters are built to handle from 125 to 650 metric 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 (1). Custom smelters
may require multiple-hearth roasters, as the longer residence time and more
moderate rate of temperature change may be advantageous in the separation
of certin impurities, such as arsenic.
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).
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 percent or more), sufficient heat is released by
the burning sulfur and supplemental fuel is required only to preheat the
42
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roaster at start-up (autogenous roasting). A reported energy requirement is
280,000 kilocalories per metric ton of copper produced (4).
Air is circulated through a hollow shaft that drives the rake arms to
prevent damage to mechanical bearings and seals.
Electricity is used to drive the roaster rakes and for auxiliary mate-
rials handling equipment. Approximately 5700 kilowatt-hours is required for
a plant of 100,000 metric tons of copper per year capacity (5).
5. Waste Streams- Gases leaving most multiple-hearth roasters are too weak
in S02 gas for direct use by a sulfuric acid plant unless the inlet air can
be enriched in oxygen or unless air infiltration is reduced. The S0? con-
centration is variable, being dependent upon 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 S0? generated by
smelter facilities once came from the roaster. Since fuel gas was not re-
quired, the emissions contained from 5 to 10 percent S02 (6). Published
data on current operations are inconsistent and contradictory; recent re-
ports give S02 concentrations in emissions from roasters at 0.5 to 2 per-
cent (2,7,8,9710,11,12). Gas from roasting is a steady stream, however, and
if sufficiently concentrated in SOp 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
is produced per metric ton of copper produced (5,9,10,11,13). Fifteen per-
cent is present in sizes below 10 microns (14). Table 2-14 presents typical
particle-size profiles. Although composition is dependent 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 2-15 gives a
weight analysis of particulate and fume emissions from a multiple-hearth
roaster. Table 2-16 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. There are
significant fugitive emissions of dust and fume at some multiple-hearth
roasters.
The multiple-hearth roaster produces no liquid wastes.
Any organic materials that enter the roaster with the copper concentrates
are vaporized or decomposed in the roaster.
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).
43
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TABLE 2-14. TYPICAL SIZE PROFILE OF MULTIPLE-HEARTH
COPPER ROASTER EMISSIONS (15)
Size, microns
Percent by weight
Entrained particles, carried from the
roaster as solids.
230 - 218
149 - 230
100 - 149
74 - TOO
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
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TABLE 2-15. CONCENTRATION AND WEIGHT ANALYSIS OF PARTICULATE
EMISSIONS FROM A MULTIPLE-HEARTH COPPER ROASTER (15)
Concentration
Entrained participates
1.4 g/Nm3
Sublimed participates
0.6 g/Nm3
Emission
Chemical
Cu
Fe
S
As
Sb
Pb
Zn
Sn
Cd
Ni
Mn
Se
Si02, CaO
CaS04
02 (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 2-16. TYPICAL LEVELS OF VOLATILE METALS IN
DOMESTIC COPPER ORE CONCENTRATIONS (6)
Lead
Zinc
Arsenic
Cadmi urn
Beryl 1 i urn
Vanadium
Antimony
Tin
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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The proposal has 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. Most multiple-hearth roasters cannot mechani-
cally withstand the high temperatures caused by oxygen enrichment. One
company that recovers arsenic from flue dusts maintains that oxygen enrich-
ment causes the arsenic to oxidize to AsgCty and prevents its removal by
current techniques (16).
A variety of devices are used in various combinations for control of
particulate emissions. Larger particulates are occasionally separated 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 (16). Cyclones can remove
80 to 85 percent of the solids, but require an addition of energy to compen-
sate 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, espe-
cially in regard to sublimed particles and fugutive losses.
The collected dusts are returned to the metallurgical processing,
usually to the smelting furnace. One smelter extracts arsenic trioxide be-
fore returning the dusts to the furnace. Excessive accumulation of im-
purities causes dusts to be discarded, but the means of disposal have not been
reported.
7. EPA Source Classification Code - 3-03-005-02
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
2. Compilation and Analysis of Design and Operation Parameters for
Emission Control Studies. Pacific Environmental Services, Inc.
(Individual draft reports.).
3. Control of Sulfur Oxide Emissions in Copper, Lead, and Zinc Smelt-
ing. 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. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
6. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
47
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7. Copper Smelters. In: Compilation of Air Pollutant Emission Fac-
tors, Second Edition. AP-42. U.S. Environmental Protection Agency,
Research Triangle Park, North Carolina. April 1973.
8. Donovan, J.R. and P.O. Stuber. Sulfuric Acid Production from Ore
Roaster Gases. Journal of Metals. November 1967.
9. Exhaust Gases from Combustion and Industrial Processes, Engineering
Science, Inc., Washington, D.C. October 2, 1971.
10. High, M.D., and M.E. Lukey. Exhaust Gases from Combustion and
Industrial Processes. PB-204 861. U.S. Environmental Protection
Agency, Durham, North Carolina. October 1971.
11. Measurement of Sulfur Dioxide, Particulate and Trace Elements in
Copper Smelter Converter and Roaster/Reverberatory Gas Streams.
EPA 650/2-74-111. U.S. Environmental Protection Agency, Washington,
D.C. October 1974.
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., L.J. Shannon, P.G. Gorman, E.W. Lawless, E.E.
Sal less, and M. Reichel. Particulate Pollutant System Study - Mass
Emissions, Volumes 1, 2, and 3. PB-203 128, PB-203 522, and PB-
203 521. U.S. Environmental Protection Agency, Durham, North
Carolina. May 1971.
14. Goldeberg, A.J. A Survey of Emissions and Controls for Hazardous
and Other Pollutants. EPA-R4-73-021. U.S. Environmental Protec-
tion Agency, Washington, D.C. 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.
48
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PRIMARY COPPER PRODUCTION PROCESS NO. 4
Fluidlzation Roasting
1. Function - The function of fliridization 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 con-
verter. Roasting also volatilizes some impurities and preheats the rever-
beratory feed. The roasted concentrate to be smelted is called calcined
feed.
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 discovery that par-
ticles of a solid added to a gas stream moving vertically upward at just the
right velocity take 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 introduced. The solid
is copper concentrate, continuously being fed and overflowing the fluidiza-
tion vessel. Both the fuel and the oxygen are completely consumed; by
elimination of excess air, the $03 content of the flue gas stream is greatly
increased, to a concentration great 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. Operators of
fluidization 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. Fluidization roasters may not pro-
vide sufficient residence time for volatilization of certain substances such
as arsenopyrites.
Three domestic copper smelters have adopted fluidization roasting (1).
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).
2. Input Materials - Copper concentrate is the only input, usually blended
or produced to a quality that the roaster can most economically process.
The concentrate may be pelletized or granulated before being fed to the
roaster.
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%. The pressure in the bed is slightly above atmospheric, and in
portions of the recycle stream the pressure may be 1 kilogram per square
centimeter or more.
49
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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 cyclone
separators that catch the large amounts of dust that rise from the bed. The
dust is returned to the bed. Dust quantities can be as much as 75 percent of
the feed (2,3). The cyclones are most properly considered as part of the
process, and the waste stream considered the outlet of the cyclones. This
waste stream contains particulates and fumes of the same chemical character
as those 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 available; they are likely to be greater
than emissions from a multiple-hearth roaster because of the more complete
separation of smaller particles from the body of the charge.
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 S02 or elemental sulfur, this is the only known technology
with which to dispose of such a concentrated gas stream.
Almost complete removal of particulates is required before the gas is
introduced into a sulfur recovery process. Electrostatic precipitators 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 reverbera-
tory 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 Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency, Washington, D.C. February 1975.
50
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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. EPA-450/2-74-002a. Environmental Protection Agency,
Research Triangle Park, North Carolina. October 1974.
51
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PRIMARY COPPER PRODUCTION PROCESS NO. 5
Drying
1. Function - Copper concentrates that are not to be processed by roasting
must pass through a dryer, whose only function is to decrease the moisture
content. Drying may be practiced to simplify handling of the concentrate.
In recent years it has been practiced to make the concentrate suitable for
direct feed to the reverberatory furnace; this is described in the industry
as "green feed." Excessive moisture in the concentrate may cause minor
eruptions or explosions that may damage the furnace (1).
Some smelters continue to use existing multiple-hearth roasters, oper-
ated 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, although
facilities for substitution of oil are usually provided. One report cal-
culates 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 kilo-
calories per hour of dryer operation (2).
Electricity is used for conveyors and mechanical operation of a dryer.
The report cited above estimates 2700 kilowatt-hours for the same size
plant (2).
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.
6. Control Technology - Dust from a drying operation frequently consists 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 particu-
late 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, collectors (3,4).
52
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Quantities of the organic materials in the concentrate are believed to
be small enough that these materials require no separate treatment.
7. EPA Source Classification Code - 3-03-005-06
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
3. Jones, H.R.. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
4. Systems Study for Control of Emissions Primary Nonferrous Smelting
Industry. Arthur G. McKee & Co. for U.S. DHEW. June 1969.
53
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PRIMARY COPPER PRODUCTION PROCESS NO. 6
Reverberatory Smelting
1. Function - Copper smelting is the process of removing from a roasted or
dried ore concentrate much of its iron and some undesirable impurities,
leaving a molten mixture that can be processed efficiently by a copper con-
verter. This is most often accomplished with a reverberatory furnace. It
may, however, be accomplished by other methods (see Process No's. 7, 8, and
11).
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 11 of the 16 smelters in this country, in one or two modifica-
tions, described as either "deep bath" or "dry hearth." Some reverberatory
furnaces are very large, capable of accepting as much as 1800 metric tons of
material per charge (1).
The reverberatory furnace is a large, arch-roofed, horizontal chamber
into which ore concentrate and flux are charged. The term "reverberatory"
refers to the configuration of the flame which enters the chamber from one
end, reverberates off the roof and strikes the charge from above. As the
temperature of the charge increases, a complex series of reactions takes
place and the charge separates into fractions. One fraction is a gas, con-
sisting of S02 and volatiles, which mix into the combustion off-gases. Two
other fractions are molten liquids, the copper matte and slag, which are not
soluble in each other and therefore separate into layers.
The matte layer consists primarily of copper and iron sulfides and
molten copper metal, which are mutually soluble. Since copper has a weak
chemical affinity for oxygen, very little copper oxide is formed and almost
all of the copper in the charge accumulates in the matte layer. Iron, on the
other hand, combines readily with oxygen to form iron oxides, which in turn
react with silica flux to form iron silicates. These compounds, plus the
calcium, magnesium, and aluminum minerals that were present in the concen-
trate, form a lighter-density slag that floats on top of the matte. Any
sulfur in the charge that is left over from the slag- and matte-forming
reactions reacts with additional oxygen to form S02 gas.
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 (2). 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. Some of the arsenic, selenium, and other trace
elements form volatile compounds and are carried away in the gas stream.
Slag is drained periodically from a skimming bay at one end of the
reverberatory furnace. Matte is also withdrawn periodically through tap
holes in the lower furnace wall. Off-gas from the furnace is usually sent
through waste heat boilers to recover a portion of the excess energy.
54
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2. Input Materials - The primary input material is the roasted or dried
concentrate, not much different from the concentrate received from the mill.
Slags from the converter and anode furnace are added for reprocessing, as are
flue dusts from dust collection equipment throughout the smelter. Precipi-
tates from hydrometallurgical operations or materials from refinery process-
ing may be added at this step. At some smelters, impure scrap copper is re-
processed as part of the change.
Flux normally consists 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.
Composition of one charge to a reverberatory furnace in Arizona is re-
ported as follows in Table 2-17.
TABLE 2-17. COMPOSITION OF CHARGE TO A
REVERBERATORY FURNACE (1)
Ore concentrate
Converter slag
Hydrometallurgical precipitate
Flue dusts
Silica flux
Limestone flux
65%
25%
2%
1%
1%
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 reverberatory furnace usually heats the
mixed charge to at least 1000°C before the matte forms and separates; tem-
peratures 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 requirements
for a smelting operation is consumed in the reverberatory furnace (4). It is
reported that 18 billion kilowatt-hours of energy was used in domestic copper
smelters in 1973 (4). Consumption of energy by this process is very high;
it is usually supplied in the form of natural gas, but pulverized coal or fuel
oil can be used. It is estimated that 500,000 kilocalories of heat is re-
quired 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 is required (5). These values give credit for steam generated
by waste heat boilers, which are almost always installed with a reverbera-
tory furnace. The reverberatory furnace is in itself thermally inefficient,
using more than 4 times the heat theoretically required (6).
Noncontact cooling water is used by copper smelters primarily for the
protection of equipment auxiliary to the roaster, converter, and reverbera-
tory furnace. Data that allocate this cooling load specifically to each
process are not available. Reported data indicate that the total cooling
55
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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 1.7 million liters of
water per day for this purpose (2).
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 opera-
tional changes to reduce this quantity, no recent data are available. The
gas is released as a dilute stream of variable composition, reported as being
from 0.5 to 2.5 weight percent S02 (3,8,9,10,11,12). Other constituents in
the exit gas are shown in Table 2-19, 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 is 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 concentrations of other
elements as shown in Table 2-18.
TABLE 2-18. ANALYSIS OF PARTICULATES EMITTED FROM
A REVERBERATORY FURNACE (15)
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 (10,11,14,16,17,18).
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 granu-
lation. Three complete analyses are shown in Table 2-20. 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, slag constitutes a large quantity of solid
56
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TABLE 2-19. COMPOSITION OF REVERBERATORY FURNACE
EXHAUST GASES (12)
Component
Carbon dioxide
Nitrogen
Oxygen
Water
Sulfur dioxide
Green feed,
% weight
8.4,
69.3
0.25 - 1.0
18.8
1.5 - 2.5
Calcined feed,
% weight
10.2
71.0
0.25 - 1.0
17.7
0.6
57
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TABLE 2-20. EFFLUENTS FROM SLAG GRANULATION (16)
(rag/1)
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
58
-------
waste, as much as 3000 kilograms per metric ton of copper produced (15).
Table 2-21 gives an analysis of trace elements found in a reverberatory
furnace slag. The bulk of the slag is a mixture of iron silicates, as shown
also in Table 2-21.
6. Control Technology - Gases from the reverberatory furnace pass through a
waste heat boiler and then through an electrostatic precipitator for par-
ticulate removal. The gases may pass through spray coolers or baloon flues
before entering the ESP units. The degree of particulate removal ranges from
50 to 99.9 percent. Particulates collected are recycled into the metallur-
gical process, normally as part of the reverberatory furnace charge, but
accumulation 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 SOg emissions from
reverberatory smelters. Intensive studies are under way to develop scrub-
bing techniques that can be applied to large volumes of hot flue gas contain-
ing small concentrations of S02. 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 domestically. Other scrubbing solutions, con-
taining 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-containing 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 being tested. The only one of these processes specific to the domestic
copper industry is DMA absorption, described in Process No. 15.
An alternate method of controlling S02 emission is to increase the
concentration in the off-gas to a level sufficient for sulfuric acid produc-
tion. Such a strategy has been successfully implemented in a number of
Japanese furnaces of conventional design and operation. Methods include fuel-
rich or oxygen-enriched combustion, use of preheated secondary air in order
to achieve rapid smelting and sulfur release, use of high grade concentrates,
instrument controlled combustion and feeding for steady level of operation,
or simply tighter construction and leak control. One Japanese smelter achieves
high S02 concentrations by blending the reverberatory off-gas with the ex-
haust from a continuous furnace that combines the functions of roasting, smelt-
ing, and converting. At another Japanese smelter, the reverberatory exhaust
is blended with the converter off-gas and scrubbed with a magnesium hydroxide
slurry, forming magnesium sulfite, which can be decomposed by calcination to
MgO and concentrated S02 (19).
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 (2,20).
Control of these mixed wastes is discussed in Section 6 of this report.
59
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TABLE 2-21. GENERAL RANGE OF REVERBERATORY FURNACE
SLAG COMPOSITION
Compound
or element
FeO
Si02
CaO
MgO
A1203
Copper
Sulfur
Composition,
percent weight
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
-------
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 reclaim the slag dumping areas, and there are no published
reports on how this could be done. It is generally assumed that the poten-
tial 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
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
2. Hallowell, J.B., et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.,
Washington, D.C. September 1973.
3. Background Information for New Source Performance Standards: Pri-
mary Copper, Zinc, and Lead Smelters. Volume I, Proposed Standards.
EPA-450/2-74-002a. U.S. Environmental Protection Agency, Research
Triangle Park, North Carolina. October 1974.
4. Rejer, 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. Hal ley, J.H., and B.E. McNay. Current Smelting Systems and Their
Relation to Air Pollution. Arther 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.
EPA 650/2-74-111. U.S. Environmental Protection Agency, Washington,
D.C. October 1974.
61
-------
11. Statnick, R.M. Measurement of Sulfur Dioxide, Particulate, and
Trace Elements in Copper Smelter, Converter and Roaster/Reverbera-
tory Gas Streams. PB-238 095. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
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., L.J. Shannon, P.G., Gorman, E.W. Lawless, E.E.
Sallee, and M. Reichel. Particulate Pollutant System Study - Mass
Emissions, Volumes 1, 2, and 3. PB-203 128, PB-203 522, and PB-203
521. U.S. .Environmental Protection Agency, Durham, North
Carolina. 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 Primary and Secondary Nonferrous
Smelting and Refining (Draft). Calspan Corporation, Buffalo, New
York. April 1975.
16. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency, Washington, D.C. February 1975.
17. Davis, W.E. National Inventory of Sources and Emissions: Copper,
Selenium, and Zinc. PB-210 679, PB-210 678, and PB-210 677. U.S.
Environmental Protection Agency, Research Triangle Park, North
Carolina. 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. 502 Control for the Primary Copper Smelter Reverberatory Furnace,
Pacific Environmental Services, Inc. EPA Draft Report. April
1977.
20. Assessment of the Adequacy of Pollution Control Technology for
Energy Conserving Manufacturing Process Options. Industry Assess-
ment Report on the Primary Copper Industry. Arthur D. Little, Inc.
Draft. October 1974.
62
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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 processing.
Smelting requires only the application of heat to produce matte and slag from
a charge of minerals. Three other smelting methods are used (Process No's.
6, 8, and 11). In electric furnace smelting, heat is supplied by electricity.
Electric furnaces for copper smelting are similar to those used in other
metallurgical industries. Three 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 pro-
cessed or blended to a suitable composition. Various fluxing materials are
also required. The charge materials are similar to those outlined for a
reverberatory furnace (Process No. 6).
Electric smelting requires the use of carbon electrodes to conduct
electric current into the layer of slag. Various types of carbon electrodes
can be used. These electrodes are consumed during operation. In one U.S.
smelter, a proprietary paste carbon mixture is consumed at a rate of 2.5
kilograms per metric ton of charge (1).
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 (2). Electric furnaces are normally enclosed in a large
building.
4. Utilities - When hot concentrate is fed from a roasting process, elec-
tric energy is consumed at a rate of 605 kilowatt-hours per metric ton of
total feed (1). Use of cold feed requires 990 kilowatt-:hours 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. Infiltration of air is
also required to ensure complete oxidation of liberated sulfur. 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 transformer capacity (4).
5. Waste Streams - S02 in the gas emitted from an electric furnace is more
highly concentrated and temperatures are lower than in emissions 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 infiltrates to oxidize these combus-
63
-------
tible materials. Very small amounts of hydrocarbons released from the
electrode compounds will also burn.
Because of the lower gas volumes and more uniform gas flow, emissions of
particulate matter would be expected to be lower than with a reverberatory
furnace; no published estimates are available. Particulate 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 three operating electric smelters in this country
use the gases from the furnaces for sulfuric acid manufacture. In each case
the gas stream is first combined with that from another furnace, such as a
fluidized roaster or converter. Acid manufacture is the best available
technology for S02 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. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency,
Washington, D.C. September 1973.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
4. Cole, R.C. Inspiration's Copper Smelter Facilities. Mining
Congress Journal. October 1974. p. 22-32.
64
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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 produce a
molten mixture that can be treated efficiently by a copper converter. Until
about 25 years ago, only two batch-type smelting processes were available,
both inefficient in energy consumption (see Process No's. 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, more easily controlled stream of flue gas, with
a high S02 content (1).
In flash smelting, ore concentrates are injected along with flux and
preheated air into a combustion chamber. Part of the sulfur of the concen-
trate 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 as in a rever-
beratory furnace. 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 an Outokumpu flash smelt-
ing 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. The concentrates 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 80 percent through 14 mesh (2) and must also be
dried. High grade "direct smelting" ores can be substituted if available.
3. Operating Conditions - Temperature in the flash chamber is maintained at
approximately 1100°C (1). Pressure is approximately atmospheric.
4. Utilities - Fuel consumption in the Outokumpu flash smelter is only
about two-thirds of that required by a reverberatory furnace in equivalent
production (2,3). Except for start-up or abnormal operations, fuel is re-
quired only to preheat the combustion air. This is reported as 7,560,000
kilocalories per hour for copper production of 100,000 metric tons per year
(2). Any fossil fuel can be used.
65
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Natural gas or oil is used to heat the furnace to start-up temperature.
Oil can also be used, if required, to maintain smelting temperatures with
some concentrates.
Electrical power is used to operate feed and air injection equipment as
well as the complex instrumentation this process requires. An estimate of
600 kilowatt-hours for a 91,000 metric ton per year copper plant 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 stream of 10 to 20 percent concentration. In Finland the con-
centrates 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 electric furnace 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 sul-
furic acid near the Finnish smelter (5); 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 SO? 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 manu-
facture.
Complete removal of particulates is required for sulfuric acid manufac-
ture, 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
66
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8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
3. Metallurgy Processing in 1974. Mining Congress Journal, February
1975.
4. Personnal Communication with Mr. Paterson. Elken - Spigerverket
a/s, New York, New York.
5. Treilhard, D.G. Copper-State of the Art. Engineering/Mining
Journal. April 1973.
67
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PRIMARY COPPER PRODUCTION PROCESS NO. 9
Peirce-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 country. See also
Process No. 10 for an alternate converting method.
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 principal in-
put. Scrap copper being recycled is also introduced at this step, as is
scrap produced within the smelter from spills or ladle "skulls," and mate-
rials 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 2-22 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°C (1,2). The converter operates at
atmospheric pressure.
4. Utilities - The converting process consumes no fuel, since oxidation of
the remaining sulfur furnishes enough heat to keep the mixture at the proper
temperature. Any excess heat is removed by addition of cold copper scrap. The
proper quantity of sulfur is obtained by carefully controlling the previous
smelting operation.
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.
5. Waste Streams - The converter emits about 120 kilograms of particulate
matter per metric ton of copper produced (3,4,5,6). Tables 2-23 and 2-24
provide data on converter dust from some Arizona smelters, and Table 2-25
gives an analysis of particulates from a smelter in Nevada.
68
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TABLE 2-22. MATERIAL BALANCE ON CONVERTERS -
SMELTER IN ARIZONA (1)
percent, weight
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
69
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TABLE 2-23. COMPOSITION OF CONVERTER DUST (7)
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
70
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TABLE 2-24. 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 + 325
-325 to + 32 microns
-32 microns to + 25 microns
-25 microns
Percent,
weight
0.5
1.5
2.5
3.5
5.0
5.0
16.6
55.6
9.8
TABLE 2-25. 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
Percent,
weight
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).
71
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Peirce-Smith converters are designed to be partially covered by a hood
that catches the S02 and participate 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
at various stages in the processing of a charge, but when diluted by the
excess air the resulting mixture contains from 4 to 10 percent S02
(2,4,5,7,8,9). Some converters may produce gas with S02 content as low as 2
percent (5,7). 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, particulate, and fume occur during those periods.
Table 2-26 gives the composition of converter off-gas from an Arizona
smelter.
Fugitive dust and fumes are generated in considerable quantities in a
converting operation. Measurement of the quantities has not been possible to
date.
There are no solid or liquid wastes from the converter process.
6. Control Technology - Gases from a Peirce-Smith converter are sometimes
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 No's.
3, 4, 6, 7, and 8).
7. EPA Source Classification Code - 3-03-005-04
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc., New York. 1967.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
3. Vandegrift, A.E., L.J. Shannon, P.G. Gorman, E.W. Lawless, E.E.
Sal lee, and M. Reichel. Particulate Pollutant System Study - Mass
Emissions, Volumes 1, 2 and 3. PB-203 128, PB-203 522, and
PB-203 521. U.S. Environmental Protection Agency, Durham, North
Carolina. May 1971.
4. High, M.D. and M.E. Lukey. Exhaust Gases from Combustion and
Industrial Processes. PB-204 861. U.S. Environmental Protection
Agency, Durham, North Carolina. October 1971.
72
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5. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
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. PB-238 095. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. 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 Smelt-
ing. Bureau of Mines Information Circular 8527, 1971.
9. 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.
10. Systems Study for Control of Emissions Primary Nonferrous Smelting
Industry. Arthur G. McKee & Co. for U.S. Department of HE&W. June
1969.
73
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TABLE 2-26. CONVERTER OFF-GAS COMPOSITION (7,11)
Component
N2
°2
so2
S03
Dust
Percent by volume
80
11
6.9 - 9.8
0.05 - 0.07
0.0053 g/lit max.
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 10
Hoboken Converting
1. Function - The Hoboken converter is one type of furnace used to produce
crude blister copper from the matte formed in the smelting process. Its func-
tion 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 construction instead of being a hood
mounted above it. This design minimizes infiltration of uncontaminated air,
should minimize local losses of S02 from the converter mouth, and allows pro-
duction of a gas with a higher and more uniform S02 content.
One smelter in this country uses Hoboken converters; however, optimal
operation has not been achieved. They are used at several smelters in Europe
and South America. The potential operating advantages of this design have
not been clearly documented, and there have been no test programs to character-
ize the level of fugitive emissions.
2. Input Materials - These are the same as for Peirce-Smith converters, con-
si stingleir^ely~of~matte from the smelters, plus silica flux and cold copper
scrap.
3. Operating Conditions - These are also the same as for a Peirce-Smith
unit, 1200°C at atmospheric pressure.
4. Utilities - These are also the same as for Peirce-Smith. No supplemental
fuel is required for the converting process.
5. Waste Streams - Since particulate matter is generated by the air being
blown through the converter charge, particulate emissions should be com-
parable to those of the Peirce-Smith.
The S02 content of the gas stream from this converter could be at least
8 percent if three or more converters are operating, and may reach as high
as 13 percent. It is calculated that with oxygen enrichment, the S02 con-
centration could be increased to 10 to 14 percent (1,2).
There are no solid or liquid wastes.
6. Control Technology - Production of S02 by the Hoboken converter is inter-
mittent, 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. This has been demonstrated by a smelter in Poland in which this
is the normal operating procedure. The one domestic smelter using Hoboken
equipment mixes the gas stream with the emissions from an electric smelter,
and after particulate removal, uses the combined stream for sulfuric acid
manufacture. At the Polish smelter separate fans remove the gases from each
converter, and fugitive emissions are minimized by increasing the draft to
each converter and creating negative pressure during such operations as
charging and pouring. At the U.S. smelter with Hobokens, the converter ex-
hausts are connected in parallel and such individual control is not possible.
75
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7. EPA Source Classification Code - 3-03-005-04
8. References -
1. Cole, R.C. Inspiration's 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. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
76
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PRIMARY COPPER PRODUCTION PROCESS NO. 11
Noranda
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 started up in
Utah. One other unit is operated by the developer in Canada.
The Noranda furnace is a horizontal cylinder about 21 meters long, into
which a mixture of concentrate and flux is continuously fed, along with fuel
and oxygen. The furnace is fired from both end walls. The mixture reacts to
form copper matte and slag, which separate into layers as in the batch smelting
processes. 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 matte are intermittently tapped
from the furnace. Noranda slag contains 10 to 12 percent copper, and is pro-
cessed to recover the copper content (see Process No. 13).
Noranda does not completly eliminate the use of the copper converter.
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
(1,2,3).
2. Input Materials - As normally used, Noranda will be fed with "clean"
concentrates that contain low concentrations of volatile elements, 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
(1,2).
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 (1,2).
4. Utilities - A principal advantage of Noranda is its efficient utiliza-
tion of fuel. Heat losses during transfer of concentrate from the roaster to
the reverberatory furnace are suppressed, as well as heat losses during the
transfer of the matte from the reverberatory furnace to the converter. In
addition, the net heat of oxidation is used for smelting. Fuel is only re-
quired to augment the fuel value of the sulfur and iron in the concentrate.
With oxygen enrichment, about 440,000 kilocalories of heat is 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 (1,2,3).
The fuel consumption reported above was based upon enrichment of combus-
tion 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
77
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use a higher degree of enrichment than in the Canadian plant.
Small amounts of electricity are used for feed injection and auxiliary
services (1,2).
5. Waste Streams - With oxygen enrichment, gas from the furnace contains
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 continuously
injected at high velocity into a moving gas stream, particulate 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 trans-
ferred to Process No. 13 for further treatment (1,2).
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 converter off-gas cannot be established
until data of gas composition are known (1,2).
7. EPA Source Classification Code - None
8. References -
1. Environmental Considerations of Selected Energy Conserving Manufac-
turing Process Options, Volume XIV, EPA 600/7-76-034n. U.S.
Environmental Protection Agency, Cincinnati, Ohio. December 1976.
2. Advertising literature and letter, Noranda Mines Limited, Toranto.
3. Mills, L.S., G.D. Hallett, and C.J. Newman. Design and Operation
of the Noranda Process Continuous Smelter. Extractive Metallurgy
of Copper. AIME. 1976.
78
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PRIMARY COPPER PRODUCTION PROCESS NO. 12
Electric Furnace 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 the
use of an electric furnace. This is the procedure being used with the flash
smelter now operating in the United States. Slag can also be treated by
flotation, as described in Process No. 13.
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 production 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 reverberatory
furnace (Process No. 6), the copper converters (Process No's. 9 and 10), and
the fire refining furnaces (Process No. 18), except with higher copper con-
tent. 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 pro-
cessed (1). Iron pyrites are usually added to the furnace charge to adjust
sulfur content. The coke is similar to that used in electric furnace opera-
tions in other industries. High grade coal can be substituted. No data on
quantities consumed are available.
3. Operating Conditions - Temperatures are maintained somewhere in the
range of 1200° to 1300°C (2). Pressures are approximately atmospheric.
4. Utilities - Electric consumption is reported as 221 kilowatt-hours per
metric ton of slag treated (1). 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
S0£ in the exit gases is negligible. Carbon monoxide and particulates 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.
No liquids are discharged from this process.
Slag discharged from this treatment is primarily iron silicate, similar
to the slag from the reverberatory furnace (Process No. 6). No analyses are
available.
79
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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. Combustion of hot gases prior to particulate removal is
sometimes practiced; the gases are then cooled with water sprays and passed
through an ESP for particulate removal.
Particulates will probably be sluiced into a tailings pond and discarded,
since they should be low in volume. They may contain quantities of trace
elements, however, and their proper disposal warrants further study.
Slags from the slag treatment furnace are discarded, with or without
granulation, as outlined for the reverberatory furnace (Process No. 6).
7. EPA Source Classification Code - None
8. References -
1. Personal Communication with Mr. Paterson Elken - Spigerverket a/s.
New York, New York.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
80
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PRIMARY COPPER PRODUCTION PROCESS NO. 13
Flotation Slag Treatment
1. Function - Slags from a flash or continuous smelter contain a signifi-
cant amount of copper, which can be reclaimed either by electric furnace
treatment, as described in Process No. 12, or by slow cooling, crushing, and
flotation. The flotation method will be used to treat slags from the Noranda
installation now beginning operation in this country (1).
As a molten slag cools, each constituent in the slag will solidify
sequentially in an order determined by the freezing temperatures of the
individual minerals. If the slag is cooled very slowly, crystals of rela-
tively pure materials will form that are large enough to be separated by
conventional concentrating procedures. Copper in the slag will form either
as small particles of metallic copper or as crystals of copper-iron sulfide,
both held in a matrix primarily of iron silicate.
Details of the existing U.S. process have not been released. It is
believed, however, that molten slag from the Noranda furnace is to be trans-
ported while still molten to a series of deep covered pits, where over a
period of days, or perhaps weeks; the slag will cool by natural conduction
through the surrounding earth. When fully cooled, the slag will be reclaimed
by conventional mining techniques, crushed, and concentrated. The resulting
concentrate will be processed in smelting furnaces in the same manner as an
ore concentrate (2).
2. Input Materials - Slag from the Noranda furnace, containing 10 to 12
percent copper, is the only known input.
To reclaim the cooled slag, explosives and concentrating reagents will
be used, as described in Process No's. 1 and 2.
3. Operating Conditions - Slag is withdrawn from the smelting furnace at
approximately 1200°C. The slag will cool to approximately ambient tempera-
ture after an extended period of time.
4. Utilities - It is believed that the molten slag will be transported to
the slag cooling area in specially-designed vehicles, requiring diesel fuel.
Whether or not slag must be heated before being added to the cooling pits has
not been announced.
Reclamation of the slag will require the same utilities used for mining
and concentrating, consisting of electrical energy for crushing and large
quantities of water for concentrating.
5. Waste Streams - A large proportion of the slag will eventually become a
waste in the form of tailings whose chemical composition is similar to slag
from a reverberatory furnace (Process No. 6). Additional wastes will be
created such as airborne particulates from mining and crushing and waterborne
contaminants from concentrating, as described in Process No's. 1 and 2.
81
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6. Control Technology - The flotation treatment of high copper content
furnace slags does not appear to introduce additional requirements for
environmental control beyond those needed for control of mining and con-
centrating wastes. In the existing U.S. application of this process, it is
not known whether special facilities will be built to reclaim the cooled
slag, or whether existing mining and concentrating facilities will be adapted
to this purpose.
7. EPA Source Classification Code - None
8. References -
1. Process Announcement. Kennecott Copper Corporation.
2. Environmental Considerations of Selected Energy Conserving Manufac-
turing Process Options: Vol. XIV, Primary Copper Industry Report.
EPA-600/7-76-034n. U.S. Environmental Protection Agency, Cin-
cinnati, Ohio. December 1976.
82
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PRIMARY COPPER PRODUCTION PROCESS NO. 14
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 also be
controlled by DMA absorption (Process No. 15) or elemental sulfur production
(Process No. 16).
Contact sulfuric acid plants are continuous steady-state processing
units that are operated in other industries using S02 made by burning ele-
mental sulfur. They may be used with waste S02 streams if the gas is suf-
ficiently 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 which oxidizes the S02. All other components of
the plant are auxiliary to this catalytic converter. 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 tempera-
ture, 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 S02 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 metallurgical opera-
tions.
2. Input Materials - Most contact sulfuric acid plants operate most effi-
ciently 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 S02 or as
the flow rate or concentration becomes less consistent. A concentration lower
than 4 percent S02 is extremely inefficient, since sufficient catalyst tem-
perature cannot be maintained (1). 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 particulate
matter, be almost completely dried, and 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 materials. Table 2-27
provides information on the acceptable limits of these impurities.
83
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TABLE 2-27. ESTIMATED MAXIMUM IMPURITY LIMITS FOR METALLURGICAL
OFF-GASES USED TO MANUFACTURE SULFURIC ACID (1)
Approximate limit, (mg/Nm3)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 ^0
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.
84
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Clean water is required to react with the 863 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 kilograms per square centimeter. 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 is
required each day (2).
A small amount of electricity is required for pumps and blowers. This
may be generated on-site in some cases, where recovery of waste heat is
maximized.
In certain patented modifications, heat from combustion of natural gas
is used to provide better efficiency at low S02 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 (3). In addition, it is likely that some SO?
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 most absorber
exit gases contain from 0.01 to 0.5 percent S0£ (4). Total flow rates may
range from 34,000 to 68,000 normal cubic meters per hour (5).
Double-contact acid plants provide a higher percentage of S02 removal if
they are fed gas with a higher S0£ content. Efficiencies higher than 99
percent have been reported. Exit gas S0£ concentration is still usually
within the same range as shown above, although one recently developed process
claims stack emissions of less than 0.005 percent S02 (6).
At a Japanese smelter the exit gas from the acid plant is routed to a
gypsum plant and the SOg concentration is less than 0.002 percent at the
stack exit (7).
In sulfuric acid plants, it is difficult to prevent some loss of SOa, 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 (8). Table 2-28 provides typical analyses
for acid plant blowdown.
85
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TABLE 2-28. RAW WASTE CHARACTERIZATION: ACID PLANT SLOWDOWN (2)
Parameter
pH
TDS
TSS
S04=
Cn-
As
Cd
Cu
Fe
Pb
Hg
Ni
Se
Te
Zn
Oil 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/Rietr-!c 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
O.COOO
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
[138.2]
0.1116
0.2501
0.0002
O.C030
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.
00
Bracketed values not used in averaging computation.
-------
In this industry, most participate matter from gas cleaning equipment is
recycled in dry form or as a water slurry back to the metallurgical 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 large, the
best control may be to increase operating efficiency by adding additional
catalyst stages or by adding heating equipment to maintain proper catalyst
temperature. Changes in the metallurgical operations to produce 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 SC^ absorption
may also be necessary.
Mist eliminators in the form of packed columns or impingement metal
screens can minimize acid mist emissions. Manufacturers claim elimination of
all but 35 to 70 milligrams of mist per cubic meter of gas, and the units at
times perform better. To prevent formation of plumes of mist during periods
of abnormal operations, however, electrostatic precipitators 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 2-29 lists the
practices of the existing smelter acid plants. The practices outlined for
plants 1 and 3 appear to describe the best available control 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 discharge to a pond for
in-perpetuity storage of the resulting gypsum is the only tested and econom-
ical method of disposal.
7. EPA Source Classification Code - None
8. References -
1. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
87
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TABLE 2-29. ACID PLANT SLOWDOWN CONTROL AND TREATMENT PRACTICES (4)
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 I/sec3
Control and/or treatment practice
" Slowdown 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 I/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.
Blowdown 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
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2. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
3. Confidential information from EPA.
4. Control of Sulfur Oxide Emissions in Copper, Lead, and Zinc Smelt-
ing. 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. Browder, T.J. Advancements and Improvements in the Sulfuric Acid
Industry. Tim J. Browder Co. San Marino, California.
7. Evaluation of the Status of Pollution Control and Process Tech-
nology - Japanese Primary Nonferrous Metals Industry. EPA Contract
No. 68-02-1375, Task 36. PEDCo Environmental, Inc. Cincinnati,
Ohio. July 1977.
8. Vandegrift, A.E., L.J. Shannon, P.G. Gormena, E.W. Lawless, E.E.
Sallee, and M. Reichel. Particulate Pollutant System Study - Mass
Emissions, Volumes 1, 2, and 3. PB-203 128, PB-203 522, and
PB-203 521. U.S. Environmental Protection Agency. Durham, North
Carolina. May 1971.
89
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PRIMARY COPPER PRODUCTION PROCESS NO. 15
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 S02 for sale.
Sulfuric acid (Process No. 14) or elemental sulfur production (Process No.
16) may also be used to control S02 emissions.
Waste gases, after first being cleaned of particulate matter and dried,
pass through a scrubber where most of the S02 is absorbed by dimethyl aniline
(DMA). The gases are then scrubbed with sodium carbonate 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 S02 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. 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 (1). Efficiencies of
up to 99 percent removal from a 5 percent gas stream have been reported (2).
Plant capacities are as high as 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. 14).
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.
Sulfuric acid, 98 percent concentration, is used for drying and scrub-
bing. The same smelter reports consumption of 18 kilograms per metric ton of
S0£ produced, as well as a small loss of the expensive dimethylaniline, 0.5
kilogram per ton of S02.
3. Operating Conditions - Feed gas must be cooled to ambient temperatures
prior to DMA absorption. Temperature in some of the regeneration steps may
reach 150°C. Pressures in the waste gas stream are near atmospheric, and may
90
-------
approach 3 kilograms per square centimeter 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 SO?
gas stream requires about 160 kilowatt-hours per metric ton os 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 S02 (3). A small amount of process water is needed to com-
pensate for purge and evaporation.
5. Waste Streams - Exit gas contains 0.05 to 0.3 percent SOg and no partic--
ulate 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 sulfuric
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 dis-
solved solids, 25 milligrams DMA per liter, and 18 milligrams suspended
solids per liter, 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 electrostatic
precipitator is usually installed to eliminate acid mist carryover.
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 metal-
lurgical 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 nonrecoverable 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
9-T
-------
feed blending, which returns it directly to the metallurgical processing.
This last alternative may be the best control technology, with or without
activated carbon absorption. The sodium may then combine into the slag where
it will not increase alkali content of the concentrator water, thus reducing
potential 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. EPA-450/2-74-002a. U.S. Environmental Protection
Agency. Research Triangle Park, North Carolina. October 1974.
2. Fleming, Edward P. and Fitt, T. Cleon. Liquid Sulfur Dioxide from
Waste Smelter Gases. I&EC. Vol. 42, No. 11. pp. 2253-2258.
November 1950.
3. Halley, J.H. and McNay, B.E. Current Smelting Systems and Their
Relation to Air Pollution. Arthur G. McKee and Company. San
Francisco. September 1970.
92
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 16
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. 14) and DMA absorption (Process No.
15). 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 economically competitive with mined sulfur
from the Texas coast.
All variations of the process use high temperatures to oxidize methane
to carbon dioxide and water, while simultaneously reducing the S02 to sulfur.
All use special catalysts and are sophisticated multi-step processes, effi-
cient in energy utilization. Like the ammonia plants or oil refinery units
they resemble, sulfur plants are most efficient as large- capacity installa-
tions.
None of the smelters in this country now include sulfur production
facilities, although one plant is being constructed.
2. Input Materials - Gas must enter the process at constant flow rate and
composition. The gases must contain 5 to 7 percent S02 (1) and must be free
from particulate matter and metal-containing fumes.
Methane in the form of 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 operate between 1000° and 1500°C at
pressures less than 25 kilograms per square centimeter. 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 variation incorporates
electrostatic precipitators as integral components, which require electric
power. No quantitative 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 H2$ gas may occur in some of the process variations.
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.
93
-------
6. Control Technology - The only further control of the waste gases from
this process is wet scrubbing, as described in connection with the rever-
beratory furnace (Process No. 6).
7. EPA Source Classification Code - None
8. References -
1. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation, Park Ridge, New Jersey. 1972.
2. Fleming, E.P. and T.C. Fitt. High Purity Sulfur from Smelter
Gases. Industrial and Engineering Chemistry. Volume 42, No. 11.
March 1950. pp. 2249-2253.
94
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 17
Arsenic Recovery
1. Function - Much 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 much of the U.S. demand for this material. The balance is
met by imports.
The recovery process consists of placing the collected dusts in a
Godfrey roaster, a small special heated enclosure, in which they are heated
until the arsenic vaporizes. The vapors are condensed in chambers, and are
then resublimed and condensed 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 multiple-hearth roasters and rever-
beratory furnaces 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 production.
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°C (1). 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 opera-
tion. Fugitive dusts are generated during the handling of the dry materials.
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 2-30 gives the analysis of this
water.
95
-------
TABLE 2-30. ANALYSIS OF ARSENIC PLANT WASHDOWN WATER (2)
Parameter
As
Cu
Zn
Pb
Cd
. H9
Se
Te
Ni
Fe
so4=
Cn-
Oil and grease
PH
Concentration
mg/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
96
-------
No solid wastes are generated by this process.
6. Control Technology - Fugitive dust emissions in and from the plant
building are controlled by the use of fabric filter baghouses on ventilating
air streams.
Effluent process gas streams are currently routed through an electro-
static precipitator before being vented to a tall stack. After treatment for
arsenic removal, the remaining dusts are returned to the smelting furnace.
In the near future, a fabric filter will be installed to remove particulate
from the waste gas.
The washdown from this plant mixes with another waste stream and dis-
charges to a pond. Table 2-30 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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Development Document for Proposed Effluent Limitations Guidelines
and New Source Performance Standards for the Primary Copper, Lead,,
and Zinc Segment of the Nonferrous Metals Manufacturing Point
Source Category (Draft). Contract No. 68-01-1518. U.S. Environ-
mental Protection Agency. Washington, D.C. December 1973.
97
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 18
Fire Refining and Anode Casting
1. Function - Impure or "blister" copper from the converters must be
refined to remove impurities. This is partially accomplished by fire re-
fining, which is the last major process that occurs at a copper smelter.
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 2-31.
2. Input Materials - Copper from the converting operation (Process No's. 9
and 10) is the principal input, usually charged into the fire refining fur-
nace while still molten.
Various slag-forming fluxes may be added. These include silica sand and
sodium carbonate.
Wooden poles are still occasionally used for the reducing step of the
process. The wood decomposes when thrust into the molten copper, producing 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 (1).
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, addi-
tional 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 kilocalories per hour of operation (1).
Compressed air is used to oxidize the molten mixture in the furnace. No
quantities have been reported.
98
-------
TABLE 2-31. GENERAL RANGE ANALYSIS OF ANODE COPPER3 (1)
Constituent
Copper
Oxygen
Sulfur
Arsenic
Antimony
Bismuth
Lead
Nickel
Selenium
Tel 1 uri urn
Gold
Silver
Platinum
Palladium
Content, % weight
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.
a Extremes omitted
g/metric ton
N.A. - not available
99
-------
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 2-32 for five smelters.
5. Waste Streams - Gases from the fire refining furnace may contain fumes
of zinc and cadmium (2) that probably condense within the exhaust duct.
Concentration of S02 has been estimated at 0.38 kilogram per metric ton of
copper treated (3). 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 (4). 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 2-33 lists the data
reported for one anode cooling operation. The "net change" represents the
difference between inlet and outlet concentrations.
There are no solid wastes from this process. All slags are returned 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. Appar-
ently it is assumed that this is one of the cleaner gas streams from pyro-
metallurgical operations because of the relative purity of the input copper.
Table 2-34 lists the controls of contact cooling water being practiced
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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency. Research Triangle Park, North Carolina. October 1974.
3. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
4. Compilation of Air Pollutant Emission Factors. AP-42. U.S.
Environmental Protection Agency. Research Triangle Park, North
Carolina. March 1975.
100
-------
5. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
6. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA-440/l-74/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
101
-------
TABLE 2-32. WATER REQUIREMENTS FOR COPPER REFINERIES (5)
Plant
A
B
C
D
E
Water intake, liters
per metric ton
of metal produced
4,000
9,000
13,000
3,000
6,000
Water consumed
(Intake minus discharge),
liters per metric
ton of metal produced
4,000
700
1,200
1,900
0
TABLE 2-33. WASTE EFFLUENTS FROM ANODE COOLING WATER (6)
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 2-34. CONTACT COOLING WATER CONTROL AND TREATMENT PRACTICES (6)
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 2-34 (continued).
Plant
code
Discharge
Control and/or treatment practice
13
14
15
18 I/sec
^340 m3/day
(125 I/sec,
45 min/day)
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
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PRIMARY COPPER PRODUCTION PROCESS NO. 19
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 approxi-
mately 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.
Fourteen electrolytic refineries are operating in the United States. Five
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
refining.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. 20).
Various additives are used to ensure a smooth deposit at the cathode.
Chlorides are added to cause silver to precipitate into the slimes.
3. Operating Conditions - Electrolytic cells are normally maintained at
60° to 65°C (1).Pressures are atmospheric.
4. Utilities - Electric power is the only source of energy. Approximately
175 to 220 kilowatt-hours 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 elec-
tricity is required for the auxiliary materials handling equipment.
Water is used for washing the cathodes 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 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 - The only pollution of the air by an electrolytic refinery
is a fine mist of sulfuric acid reported to be created near the electrodes.
Data on this source of pollution are not available.
105
-------
Most refineries reclaim the copper from impure solutions (see Process
No. 20), but two do not, and therefore create a substantial liquid waste
directly from the electrolytic cells. Table 2-35 gives the range of composi-
tion 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 No's. 22 and
24). Because this slime represents a product of considerable 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 quantities,
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 placing it
in a lined pond and allowing it to evaporate to dryness. Climatic conditions
at this site make this procedure workable. The other refinery follows a
practice often used in the copper industry and mixes the solution 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 reclama-
tion and partial recycle of spent electrolyte; this represents the best
available control technology (see Process No. 20).
There are usually no controls specifically designed to handle spills,
leakages, and accidental discharges of electrolyte.
7. EPA Source Classification Code - 303-005-05
8. References -
1. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
106
-------
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
107
-------
TABLE 2-35. GENERAL RANGE ANALYSIS OF ELECTROLYTE, REFINED
COPPER, AND ANODE SLIME3 (2,3)
Constituent
Sulfuric acid
Copper
Oxygen
Sulfur
Arsenic
Antimony
Bismuth
Lead
Nickel
Selenium
Tellurium
Gold
Silver
Platinum
Palladium
Iron
Electrolyte,
9/1
170-230
45-50
0.5-12
0.2-0.7
0.1-0.5
2.0-20.0
Refined copper,
% weight
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,
% weight
(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
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PRIMARY COPPER PRODUCTION PROCESS NO. 20
Electrolyte Purification
1. Function - In operation of the electrolytic cells in a refinery, certain
impurities 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 at the cathode
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 are returned 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 separa-
tion of the acid and produce a stream in which impurities are more concen-
trated. The acid is returned to the electrolyte circulation.
Effluent from the liberator cells or the dialysis equipment may be
concentrated further by removing water in vacuum evaporators. Concentration
of the acid produces a sludge, which has a high concentration of nickel
sulfate and usually 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 refineries may practice all, part, or none of these treatments.
Three refineries recover nickel. One refinery consumes all spent electrolyte
in an associated chemical operation (1).
2. Input Materials - The input is the purge stream from the recirculatinq
electrolyte. The range of analysis is given in Table 2-35 (Process No. 19).
This shows it to be a stream of warm, concentrated acidic copper sulfate
solution, also containing nickel, arsenic, antimony, and bismuth. Smaller
quantities of iron, cobalt, zinc, lead, selenium, tellurium, and other
elements are also found in the stream.
3. Operating Conditions - Temperatures are less than 100°C and pressures
are atmospheric, except in some evaporation operations (2).
4. Utilities - Electricity is used to drive pumps and mechanical equipment,,
and to operate the liberator cells. Since the average concentration 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 kilowatt-hours per metric
ton (1).
109
-------
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
2-31 (Process No. 18) 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 (AsH3) (4). This very poisonous gas can accumulate
to dangerous levels if the liberator cells are not well ventilated. Arsine
will slowly oxidize in the atmosphere to arsenic trioxide and water. Quan-
tities 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 elements must
result in either continuous or occasional disposal of a quantity of "black
acid", regardless of the extent of electrolyte treatment employed.
In some refineries evaporated water from the electrolyte constitutes
another waste stream, which usually is mixed with volumes of steam condensate
and direct cooling water in barometric leg discharge devices. Table 2-36
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 control technology
cannot be evaluated unless the 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.
If it is possible to sell or give away the black acid to the fertilizer
industry, as has been reported, the impurity elements would be transferred
into the gypsum ponds from phosphate rock treatment. This would eliminate
disposition in local tailings ponds, which are already loaded with metal ions
from other sources. The neutral to acidic nature of phosphate ponds may
cause a greater degree of precipitation of arsenic and antimony than would
occur in the alkaline water designed to precipitate copper. On the other
hand, phosphate ponds may already have a heavy load of radium, and are usually
located in an area of greater precipitation than copper smelters. Further
research to establish the compatibility of the wastes of these two industries
is indicated.
110
-------
TABLE 2-36. WASTE EFFLUENTS FROM NiS04 BAROMETRIC CONDENSER (1)
Parameter
PH
Alkalinity
COD
Total Solids
Dissolved Solids
Suspended Solids
Sulfate (as S)
Arsenic
Cadmi urn
Copper
Iron
Lead
Nickel
Zinc
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
111
-------
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, instru-
mented, and operated.
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 Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-74/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
4. Trace Pollutant Emissions from the Processing of Metallic Ores.
PEDCo-Environmental Specialists, Inc. August 1974.
112
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PRIMARY COPPER PRODUCTION PROCESS NO. 21
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 copper. 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 electrolytic
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 phospho-
rous-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" operation (1). This modification is comparable
to fire refining and anode casting (Process No. 18).
3. Operating Conditions - Depending on the process details, temperatures
range from 1150°C to 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 effi-
ciencies; power consumption is rated at 250 to 300 kilowatts for maintaining
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 equipment.
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).
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 2-37 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.
113
-------
TABLE 2-37. ANALYSIS OF WATER USED TO COOL REFINERY SHAPES (4)
(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
114
-------
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
8. References -
1. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
115
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 22
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. 24), 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 cir-
culated. Copper dissolves in the acid as a solution of copper sulfate. This
solution is either mixed with the electrolyte in the refinery cells (1), or
with the electrolyte.purge to the liberator cells, or may be used for copper
sulfate production (Process No. 23).
2. Input Materials - The primary input is the slime from the electrolytic
cells (Process No. 19), which may contain 20 to 40 percent copper (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 circulating
solution, but chemical action may cause a slight temperature rise above
ambient. Pressures are atmospheric to slightly higher, not exceeding one
kilogram per square centimeter.
4. Utilities - A small quantity of electricity is used to power pumps for
acid circulation. Water or steam concentrate is used to wash the leached
slimes prior to transferring them to the roaster.
5. Waste Streams - A minor evolution of $03 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 with other streams (if available) or by scrubbing would be
the best control technology. Since quantities are small, none of the re-
fineries control this emission except by local ventilation.
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 Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-74/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
116
-------
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
3. Leigh, A.M. Precious Metals Refining Practice International
Symposium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 95-110.
117
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 23
Precipitation
1. Function - The function of this process is to precipitate copper sulfate
in crystal form as a marketable by-product. The solution from water or acid
leach constitutes part or all of the source. 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 concentrated liquor either re-
turns 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 the Slime Acid
Leach (Process No. 22), containing copper sulfate and sulfuric acid, or from
the Slime Water Leach (Process No. 25), containing copper sulfate in water.
Copper powder may also be added.
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 (1).
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 auxil-
iaries. 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 S02, 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
process. If quantities of S0£ 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.
7. EPA Source Classification Code - None
118
-------
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Leigh, A.H. Precious Metals Refining Practice International
Symposium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 95-110.
119
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 24
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. 22). Heating the slimes in a strong acid environ-
ment converts the remaining copper to soluble copper sulfate, which can be
removed by a subsequent water leach process (Process No. 25) (1). Roasting
also converts some of the silver and tellurium to soluble salts and volatil-
izes 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 kilogram of slime
treated (2).
Muds from the scrubber (Process No. 27) 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 mineralized
particulates and fumes. Roasting breaks down silver and copper selenides,
releasing Se02 (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. 27), but any loss can represent a hazardous waste. No
analyses of this stream have been reported.
No liquid wastes are generated.
Solids from the roaster contain the most valuable metals, and are
usually carefully transferred to the water leach equipment (Process No. 25).
There is no solid waste.
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
120
-------
8. References -
1. Leigh, A.H. Precious Metal Recovery Practice International
Symposium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 95-110.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
3. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
121
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 25
Slime Mater Leach
1. Function - The objective of this process is to reprecipitate all the
silver and tellurium that has been made water-soluble in the roasting pro-
cess, and to dissolve and separate all the soluble copper (1).
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 fil-
tered. The liquid solution of copper sulfate returns to the electrolytic
cells or is used for copper sulfate production (Process No. 23). The filter
cake is transferred to the Dore" Furnace (Process No. 26) (3).
2. Input Materials - Roasted slime from Process No. 24 is the principal
input. Powdered copper in slightly less than stoichiometric proportions is
added. Water is also added.
3. Operating Conditions - The temperature in the leach tank is less than
100°C, but is not carefully controlled (3). Pressures are atmospheric,
rising to less than 3 kilograms per square centimeter during filtering.
4- Utilities - No external heat is added to this process. The hot solids
from the roaster add incidental heat.
Either deionized water or steam condensate is used to prevent introduc-
tion of foreign elements into the electrolyte solution.
5. Waste Streams - Except for accidental spills, no wastes are generated by
this process.
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
122
-------
3. Leigh, A.M. Precious Metals Refining Practice International
Symposium on Hydrometallurgy. Chicago, Illinois. Feburary 25
March 1, 1973. pp. 95-110.
123
-------
PRIMARY COPPER PRODUCTION
PROCESS NO. 26
Dore" 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, consisting
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
antimony (1). 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 (2) and is further treated (see Process No.
28). An oxidative slag is then formed by blowing air through the molten
metal (3), removing bismuth and any remaining copper. This slag is returned
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. 30). Table 2-38 gives the approxi-
mate range of analysis.
Table 2-38. DORE METAL ANALYSIS (2)
% Weight
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 content;
they contain about 18 percent water (2) and no sulfur.
Silica sand is the first flux, and a 2:1 mixture of sodium carbonate 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).
124
-------
4- Utilities - Gas or oil fuel a Dore" furnace (1,2). Compressed 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 selenium, tellurium,
some arsenic, antimony, and lead. These are normally sent to a wet scrubber
(Process No. 27). The precious metals content of the particulate matter is
high enough that care is taken to collect them, but no analyses have been
reported.
There are no liquid wastes from the Dore* furnace.
This process produces no solid wastes if all the slags are processed 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.
6. Control Technology - Wet scrubbing of the gases for removal of parti -
culates 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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Leigh, A.H. Precious Metal Refining Practice International Sym-
posium on Hydrometallurgy. Chicago, Illinois. February 25 - March
1, 1973. pp. 95-110.
3. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
4. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency,
Washington, D.C. September 1973.
125
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 27
Scrubber
1. Function - Gases from the slimes roaster and the Dore1 furnace contain
particulates in quantities that justify their recovery for further processing.
The gases also contain fumes, especially of selenium, which hydrolyze in the
water scrubber, allowing their separation for sale.
The scrubbers are generally of the water spray type (1), with the water
continuously recirculating. As solid material accumulates, periodic blowdown
is performed. The amorphous selenium is often removed by flotation (2), or
occasionally the blowdown is combined with the soda slag leach liquor (Pro-
cess No. 28). Muds from the scrubber are recycled to the slimes roaster
(Process No. 24).
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,
rnethylamyl alcohol and liquid colloid glue are used as flotation reagents
(2).
3. Operating Conditions - The gases entering the scrubber are extremely
hot, about 1000° to 1300°C. The water sprays are at ambient temperatures
(3). Pressures are near atmospheric.
4. Utilities - Water is used as makeup 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 previous
operations, no liquid or solid wastes are produced.
6. Control Technology - Most refiners find it economical to install electro-
static precipitators on the scrubber effluent to remove the highly metal-
liferous 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
8. References -
1. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
126
-------
2. Leigh, A.H. Precious Metals Refining Practice International
Symposium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 75-110.
3. Particulate and Sulfur Dioxide Emission Control Cost Study of the
Electric Utility Industry. 68-01-1900. U.S. Environmental Pro-
tection Agency. Washington, D.C.
127
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 28
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 because
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
Dor6 furnace for reprocessing. The leached solution is further treated (see
Process No. 29).
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 leach-
ing, to less than 100°C (2,4). Pressures are atmospheric during the leach,
and less than 2 kilograms per square centimeter 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 carefully handled
(3).
6. Control Technology - None is required.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Leigh, A.H. Precious Metal Recovery Practice International
Symposium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 95-110.
128
-------
3. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
4. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
129
-------
PRIMARY COPPER PRODUCTION PROCESS NO. 29
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-products (1). Both are valuable for use in manufacture of
electrical and electronic products, and in xerographic copying machines.
The alkaline solutions are made acidic with sulfuric acid to a pH of
5.5 to 6.5 (2,3). Tellurous acid (^TeOo) precipitates, and is removed by
filtration. Then the solution is treated by bubbling $03 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 precipitates. 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 produce
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. 28), to which are added selenium and tellurium
extracted or floated from scrubber precipitates (Process No. 27).
Laboratory-grade reagents are normally used, such as sulfuric acid,
sodium hydroxide, compressed and liquified SOg, and others. Quantities are
very small.
3. Operating Conditions - Temperatures may reach 450°C during some purifica-
tion steps. Pressures are normally atmospheric.
4. Utilities - A variety of laboratory utilities may be employed. Consump-
tion of each is negligible.
5. Waste Streams - No gas or solid wastes are generated by this process in
anything other than trace amounts.
Liquids are low in volume and normally discharge through standard
laboratory waste systems.
130
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6. Control Technology - None applicable
7. EPA Source Classification Code - None
8. References -
1. Leigh, A.H. Precious Metal Recovery Practice International Sym-
posium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 95-110.
2. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
3. Hallowell, J.B. et al. Water Pollution Control in the Primary
Nonferrous Metals Industry - Volume I. Copper, Zinc, and Lead
Industries. EPA-R2-73-274a. U.S. Environmental Protection Agency.
Washington, D.C. September 1973.
4. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
131
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PRIMARY COPPER PRODUCTION PROCESS NO. 30
Pore* Metal Separation
1. Function - In a series of complex chemical and electrochemical labora-
tory operations, the Dore* metal, a mixture primarily of silver, gold, and the
platinum group metals, is separated into specification grades of each of
these metals (1).
A special small electrolytic cell, the Moebius cell, is used to separate
the silver, which is further processed to produce bullion bars, analyzed at
99.97 percent silver, of 1000 troy ounces each (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.
26).
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 pro-
cessing 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 with other processes in this industry.
5. Waste Streams -
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 electrolyte acids in quanti-
ties of a few gallons at most. No solid wastes are produced, although
residues may occasionally be returned to the Dore" furnace.
(5. Control Technology - Local ventilation is the only control exercised for
this process.
7. EPA Source Classification Code - None
132
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8. References -
1. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Primary
Copper Smelting Subcategory and the Primary Copper Refining Sub-
category of the Copper Segment of the Nonferrous Metals Manufac-
turing Point Source Category. EPA 440/1-75/032-b. U.S. Environ-
mental Protection Agency. Washington, D.C. February 1975.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
3. Leigh, A.H. Precious Metal Refining Practice the International
Symposium on Hydrometallurgy. Chicago, Illinois. February 25 -
March 1, 1973. pp. 95-110.
133
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PRIMARY COPPER PRODUCTION PROCESS NO. 31
Vat Leaching
1. Function - Vat leaching is a simple form of hydrometallurgy in which
copper is dissolved from oxide ores to form aqueous solutions. The leaching
takes place in an arrangement of tanks or vats.
Oxidized copper minerals, occurring as partially weathered deposits in
the mine, cannot easily be processed by conventional smelting processes.
These deposits are selectively mined and crushed to about 1 to 1.25 centi-
meters (1). The crushed ore is then placed in concrete vats of up to 18,000
metric tons capacity, and subjected to alternate flooding with sulfuric acid
and draining. After the copper oxides are converted to soluble copper sul-
fate, the remaining soluble copper is removed by a countercurrent wash of
fresh water. The vat floor is a filter which facilitates upflow and downflow
of wash and leach solutions. The resulting solutions are too dilute for
electrowinning; they are usually treated by cementation (Process No. 33) or
solvent extraction (Process No. 34) (1).
One proprietary process has been developed for vat leaching of a roasted
sulfide ore in which the sulfide ore is converted into sulfates prior to
hydrometallurgical processing (see Process No. 36).
Vat leaching is similar in principal to the leaching of sulfide ores
(Process No. 32), but the vat leaching operations are usually more carefully
controlled and result in lower potential for damage to the environment. Vat
leaching is the most efficient process yet developed for the recovery of
copper values from oxidized copper minerals.
2. Input Materials - The principal input is the oxide ore material, as
described above.
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 extraction of a metric ton of
copper from a 1 percent ore body containing oxidized minerals will require
about 4400 liters of 96 percent acidity (2). »
3. Operating Conditions - The process 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 was 50 to 200 cubic meters per metric ton of copper precipitate (2).
5. Waste Streams - Vat leaching produces a large amount of tailings of
waste rock that is sluiced into a tailings pond. This material is comparable
with the waste from a concentrator plant (Process No. 2). Frequently the same
pond is used for both concentrator and vat-leaching tailings.
134
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The circulating stream of a leaching operation may become so rich in
impurities 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 (3).
6. Control Technology - Most installations mix the discharge of this
process with mining or concentrating wastes. Control is described in con-
nection with Process No's. 1 and 2.
7. EPA Source Classification Code - None
8. References -
1. Williams, Roy E. Waste Production and Disposal in Mining, Milling
and Metallurgical Industries. Miller Freeman Publications, Inc.
San Francisco. 1975.
2. Roberts, R.W. San Xavier Vat Leach Plant Operation. Mining
Congress Journal. December 1974.
3. Davis, W.E. National Inventory of Sources and Emissions: Copper,
Selenium, and Zinc. PB 210-677, 678, and 679. U.S. Environmental
Protection Agency. Research Triangle Park, North Carolina. May
1972.
135
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PRIMARY COPPER PRODUCTION PROCESS NO. 32
Sulfide Ore Leaching
1. Function - Heap and dump leaching are simple forms of hydrometallurgy in
which copper is dissolved from sulfide ores to form aqueous solutions. In
heap leaching, the ore is placed in a pile on the ground. In dump leaching,
the overburden and low grade waste from the mine are leached in the dumps
formed during the mining operation. In situ and thin layer leaching may also
be utilized for sulfide ores.
This process is an accelerated form of natural weathering. It is
usually applied to low-grade ore that contains less than 0.4 percent copper
(1,2,3). Material is placed in an area provided 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 2.5 (5) and
may contain from 1.0 to 18 grams of copper per liter (4,5).
In the leaching of sulfide ores, barren solutions of sulfuric acid from
a copper cementation process are applied originally, after which only makeup
water is required periodically to sustain the leaching process. Water and
oxygen react with pyrite in the dump to generate sulfuric acid and ferric
sulfate; this solution effectively dissolves the copper present. If the ores
contain significant amounts of oxides or carbonates, sulfuric acid must be
added periodically. A dump leaching site is characterized by a grid of ponds
that collect the pregnant leach liquor (6).
In situ leaching involves breaking the ore in place and alternately
circulating air and leach solution through the fractured material. The
pregnant liquor is collected in a system of tunnels.
A recent modification of sulfide ore leaching is the "thin layer"
process, in which still further acceleration of the weathering reactions is
brought about by spreading the ore thinly over a large surface area. This
modification is in use in South America, but has not been reported to be in
use in this country.
These techniques of hydrometallurgy allow the extraction of copper from
low-grade ores without evolution of sulfur dioxide. These can be small
operations that require less capital expenditure than pyrometallurgical
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.
2. Input Materials - The principal input is the ore materials, as described
above. In most cases, these would otherwise be waste materials, unprofitable
to process by conventional techniques.
136
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Sulfuric acid is the leaching chemical. To some extent, leaching opera-
tions are practiced as a means for disposition of excess smelter acid.
Consumption will vary and will depend largely on the composition of the
gangue rock. If there is no limestone in the gangue, very little acid is
consumed.
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).
5. Waste Streams - Wastes from heap leaching include fugitive dusts from
materials handling, and quantities of highly mineralized solid wastes con-
taining 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.
The circulating stream of a leaching operation may become so rich in
impurities 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 (7).
6. Control Technology - Most installations mix the discharge of this
process with mining or concentrating wastes. Control is described in connec-
tion with Process No's. 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. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. October 1974.
2. Copper Hydrometallurgy: The Third-Generation Plants. Engineering
and Mining Journal. June 1975.
3. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
4. Roberts, R.W. San Xavier Vat Leach Plant Operation. Mining
Congress Journal. December 1974.
5. Gardner, S.A. and G.C.I. Warwick. Pollution-Free Metallurgy:
Copper via Solvent-Extraction. Engineering and Mining Journal.
April 1971.
137
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6. A Study of Waste Generation, Treatment and Disposal in the Metals
Mining Industry. PB-261 052. U.S. Environmental Protection
Agency, Washington, D.C. October 1976.
7. Davis, W.E. National Inventory of Sources and Emissions: Copper,
Selenium, and Zinc. PB-210 679, PB-210 678, and PB-210 677. U.S.
Environmental Protection Agency. Research Triangle Park, North
Carolina. May 1972.
138
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PRIMARY COPPER PRODUCTION PROCESS NO. 33
Cementation
1. Function - The cementation process converts soluble copper into a
metallic precipitate through chemical reaction with metallic iron. It is
used to recover copper from strong solutions created by other processes,
especially those from heap and vat leaching.
This process is dependent upon 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 method
(1.2).
Periodically, the trough is cleaned and the cement copper is sent to a
smelter for processing. The cement copper is usually a mixture of copper
with iron compounds and other insoluble minerals. The copper content is
generally around 70 percent and is rarely more than 90 percent (1,3).
In many of its applications, cementation is being replaced by solvent
extraction and electrowinning techniques (see Process No's. 34 and 35).
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 concentrated
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 is in the final steps of develop-
ment (see Process No. 37).
3. Operating Conditions - Cementation processes normally operate at atmo-
spheric 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.
5. Waste Streams - Atmospheric pollution from the cementation process is
negligible. There may be tiny amounts of hydrogen gas created by a side
reaction of acid with the iron.
139
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No liquid wastes can be directly assigned to this process when used as
an auxiliary to a larger liquid handling operation. In the case of a sepa-
rate operation, analysis of a typical treated leach liquor is presented below
in Table 2-39.
TABLE 2-39. ANALYSIS OF TAILINGS EFFLUENT
FROM A PRECIPITATION PLANT (5)
Parameter
Sulfate
Copper
Iron
Lead
Mercury
Selenium
Zinc
Maximum, mg/1
53,000
76.3
3100
0.92
0.0006
0.95
146
Minimum, mg/1
33,000
27.7
2050
0.05
0.0001
0.01
129
Average, mg/1
38,882
52.2
2632
0.67
0.0003
0.12
136
Solid wastes resulting from cementation includes scrap iron partially
used, discarded, or abandoned, causing 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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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. EPA-450/2-74-002a. U.S. Environmental Protection
Agency, Research Triangle Park, North Carolina. 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. EPA/1-75/032-6. U.S. Environmental Protection Agency.
Washington, D.C. February 1975.
5. Personal Communication with J.V. Rouse, U.S. Environmental Protec-
tion Agency. National Enforcement Investigations Center. Denver,
Colorado. 1976.
140
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PRIMARY COPPER PRODUCTION PROCESS NO. 34
Solvent Extraction
1. 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 solvents in 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 mixed with the sol-
vent. 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 hydrometallurgical
processes (1). The concentrated acid solution can be directly treated by
electrowinning (see Process No. 35). This process can also be made con-
tinuous, rather than batch, to adapt it to large-scale operations.
It is reported that about 95 percent of the copper 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 quantities have not been reported.
3. Operating Conditions - No special temperature limits are reported; it is
assumed that ambient temperatures and atmospheric pressure are satisfactory.
4> Utilities - A small amount of electricity is required for agitation and
liquid pumping.
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
141
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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 raf-
finate (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 Journal.
April 1971.
3. Background Information for New Source Performance Standards:
Primary Copper, Zinc, and Lead Smelters. Volume I, Proposed
Standards. EPA-450/2-74-002a. U.S. Environmental Protection
Agency. Research Triangle Park, North Carolina. October 1974.
4. Ion Exchange: The New Dimension in Copper Recovery Systems.
Engineering and Mining Journal. June 1975.
142
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PRIMARY COPPER PRODUCTION PROCESS NO. 35
Electrowinm'ng
1. Function - Electrowinm'ng is a process for the extraction of relatively
pure copper metal from a solution containing copper ions. This is an electro-
lytic process, similar to the cells of an electrolytic refinery (see Process
No. 19), except than an inert anode is used. Copper metal deposits at the
cathode, and the water in the solution is electrolytically decomposed,
liberating oxygen at the cathode, and regenerating the sulfate ion as sul-
furic acid.
If the copper solution is relatively pure, the copper produced by
electrowinning is comparable with the best electrolytic copper, assayed as
99.9 percent plus (1). 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 directly
from vat leaching solutions (Process No. 31), and from purified solution from
solvent extractions (Process No. 34). 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 mate-
rials (see Process No. 38) (2).
Additives to produce a uniform cathode deposit are necessary. They are
the same as for electrolytic refining. One report lists glue for electro-
winning being added at a rate of 0.02 to 0.06 kilogram 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 (3). In a loop with a solvent extraction process, 2.44 kilowatt-
hours are required to produce a kilogram of copper (4). In direct electro-
winning of a vat or heap leach solution 2.79 kilowatt-hours per kilogram of
copper are required (3). 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 removing
them from the cell.
5. Waste Streams - The oxygen produced at the anode of an electrowinning
cell can be considered either as an atmospheric emission or as a by-product.
If it is discarded to the atmosphere, there are no deleterious environmental
effects.
143
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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 accumula-
tion of other elements, was discussed in connection with Process No's. 31,
32, and 34.
There are no solid wastes from this process.
6. Control Technology - No special controls are applicable to the waste-
water 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 Sol vent-Extraction. Engineering and Mining Journal.
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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
4. Ion Exchange: The New Dimension in Copper Recovery Systems.
Engineering and Mining Journal. June 1975.
144
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PRIMARY COPPER PRODUCTION PROCESS NO. 36
Sulfation Roasting
1. Function - One company has developed a hybrid process that will use a
fluidization roaster (Process No. 4) to prepare a calcine especially suited
to vat leaching (Process No. 31). This is the technique of sulfation roast-
ing. 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 operation.
2. Input Materials. - This plant will use an ore concentrate that is pre-
dominantly chalcopyrite (CuFe$2) (1). It is expected that any sulfide con-
centrate could be used. The concentrate is blended with Fe203, which moder-
ates 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 con-
ventional 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 pre-ignite 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. Particulate
emissions following the roaster cyclones have not been estimated.
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 sulfuric
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.
7. EPA Source Classification Code - 3-03-005-02
145
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8. References -
1. Haver, P.P. and M.M. Wong. Lime Roast-Leach Method for Treating
Chalcopyrite Concentrate. U.S. Bureau of Mines, Washington, 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., D.J. Bauer, and R.E. Lindstrom. Copper Recovery
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. EPA-450/2-74-002a. U.S. Environmental Protection
Agency. Research Triangle Park, North Carolina. October 1974.
5. Potter, J. Personnel Communication on Hydrometallurgical Pro-
cesses. Bureau of Mines. Salt Lake City, Utah. 1976.
146
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PRIMARY COPPER PRODUCTION PROCESS NO. 37
Sponge Iron Plant
1. Function - One company is building a sponge iron plant to accompany
its sulfation roasting installation (Process No. 36)(1). After leaching
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 cementation.
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 approximately
1100°C (3).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.
There should be no liquid waste.
The process will generate a solid waste in the form of a slag.
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.
7. EPA Classification Code - None
147
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8. References -
1. Potter, 0. Personal communication on Hydrometallurgical Pro-
cesses. Bureau of Mines. Salt Lake City, Utah, 1976.
2. Hydrometallurgy Makes Advances in Copper Processing. Engineering
and Mining Journal. 1973.
3. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
148
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PRIMARY COPPER PRODUCTION PROCESS NO. 38
CLEAR Reduction
1. Function - An advanced hydrometallurgical process has been developed
with the trade name of CLEAR (Copper Leaching, Extraction, and Refining).
The descriptions for Process No's. 38 through 40 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 (1). Primarily the cupric chloride is
active during this step, reacting with copper ore minerals to form cuprous
chloride, additional ferrous chloride, and elemental sulfur. 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 et
second time in the oxidation process (Process No. 40).
Although leaching of fresh concentrate does occur, the principal purpose
of this step is to prepare the liquor for electrowinning. It may be reduced
with other materials if necessary, then filtered and sent to electrowinning
cells (2). These cells are similar to those described in Process No. 35,
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. 39).
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 (1). It is
received directly from the oxidation leach of the previous batch (see Process
No. 40).
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 temperatures
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.
5. Waste Streams - Some loss of hydrochloric acid vapor from a residual in
the leach liquor 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 corrosive
solutions, however, accidental losses of liquids are likely.
149
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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 C.H. Curtis. 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 Hydrometallurgy.
Chemical Engineering. Volume 83, No. 1: pp. 79-81. January 5,
1976.
150
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PRIMARY COPPER PRODUCTION PROCESS NO. 39
CLEAR Regeneration - Purge
1. Function - Exhausted and stripped solvent from the electrowinning cells
that follow the reduction process (Process No. 38) is oxidized with air to
remove excess iron and to prepare the solution for the pressurized leach
operation (Process No. 40). 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 oxidation (1). 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 kilograms per square centimeter (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 electrolyte (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 materials. This
is probably accomplished by external cooling, condensation, and water scrub-
bing. No details have been disclosed.
The solid wastes from this process are mixed with the wastes from the
oxidation step (Process No. 40) and sluiced into a settling pond. In loca-
tions other than the arid region where this plant is operating, secondary
water pollution could be substantial.
7. EPA Classification Code - None
151
-------
8. References -
1. Potter, J. Personnel Communication on Hydrometallurical Process.
Bureau of Mines. Salt Lake City, Utah, 1976.
2. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometallurgy.
Chemical Engineering. Volume 83, No. 1: pp. 79-81. January 5,
1976.
3. Atwood, G.E., and C.H. Curtis. Hydrometallurgical Process for the
Production of Copper. U.S. Patent No. 3,785,944. January 15,
1974.
152
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PRIMARY COPPER PRODUCTION PROCESS NO. 40
CLEAR Oxidation
1. Function - This is the principal leaching operation of the CLEAR system,
in which partially leached ore concentrate is contacted with freshly regen-
erated 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. 38) (2).
2. Input Materials - Solid residue from Process No. 38 and liquor from
Process No. 39 are the only input materials. The ratios have not been dis-
closed. The liquor at this stage contains about 6 percent cupric chloride
and 15 percent each of ferric and sodium chlorides (1).
3. Operating Conditions - This is a high-temperature, pressurized leaching
operation. Temperatures of 140°C and pressures of 2.7 kilograms per square
centimeter are used (3).
4. Utilities - Source of heat has not been reported; either steam or
direct firing could be applicable. Electricity is undoubtedly required, but
again no information has been published.
5. Haste Streams - Details of the process have not been disclosed in
sufficient detail to establish whether emissions of gases to the atmosphere
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 (1) 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 operating
control of hydrochloric acid emission, since even small concentrations of
this very corrosive gas can damage plant equipment. If needed, scrubbing
with an alkali can further reduce the concentration.
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.
153
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In other locations, however, this waste could cause severe secondary pollu-
tion 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 C.H. Curtis. Hydrometallurgical Process for the
Production of Copper. U.S. Patent No. 3,785,944. January 15,
1974.
2. Potter, J. Personnal Communication on Hydrometallurgical Process.
Bureau of Mines. Salt Lake City, Utah, 1976.
3. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometallurgy.
Chemical Engineering. Volume 83, No. 1: pp. 79-81. January 5,
1976.
154
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SECTION 3
LEAD INDUSTRY
INDUSTRY DESCRIPTION
While U.S. lead mine production decreased slightly in 1976, output from
primary refineries and secondary smelters was higher than in the previous
year (1). Consumption increased about 8 percent in 1976 (1). Some areas of
consumption have decreased, however. 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. Lead has also been
replaced in such applications as joining material for cast iron pipe, and in
plumbing and most other construction applications.
Although new methods are being developed for lead production, 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 especially for lead. Sections of this
deposit produce almost pure galena, analyzed as 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 other smelters are in Idaho, Montana, and Texas. The industry
employs about 7000 people, of which two-thirds are employed in mining and
concentrating operations.
Raw Materials
Lead is most often found in nature as galena (PbSg), the primary sulfide
of lead. Deposits are rarely pure since the lead-bearing compound is usually
mixed with pyrite, sphalerite, and pyrrhotite. Most of these deposits con-
tain very little copper.
Oxidized lead ores also occur and are composed primarily of anglesite
and cerussite, the weathered products of galena. Table 3-1 lists the impor-
tant lead ore minerals, together with others in which the lead is combined
with phosphorus, vanadium, and other elements.
155
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TABLE 3-1. LEAD MINERALS, BY NAME.AND COMPOSITION
Mineral
Galena
Anglesite
Cerussite
Pyromorphite .
Vanadinite
Crocoite
Wulfenite
Linarite
Composition
PbS
PbS04
PbC03
Pb5Cl(P04)3
Pb5ci(vo4)3
PbCr04
PbMo04
PbO-CuO-S03 H20
88.6
68.3
77.5
76.3
73.0
63.9
56.4
156
-------
Most lead produced in this country is from domestic ores. Little is
produced from imported concentrates. General imports of lead represented 15
percent of total consumption in 1976 (2). A considerable quantity is pro-
cessed 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 marketable products. The lead and zinc industries are so inter-
related that allocation of by-products by industry is difficult.
Table 3-2 provides the basic 1976 statistics of this industry.
Companies
In 1976 the domestic lead mining industry consisted of approximately 31
mines in 15 states with production valued at $290 million (2). Twenty-five
mines accounted for 99 percent of this output and the eight leading mines,
all located in Missouri, yielded 80 percent of all ores and concentrates (2).
Four companies own and operate the six lead smelters in this country. 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 3-3 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.
Domestic demand for lead is expected to increase at an annual rate of
about 1.5 percent through 1980 (2). Identified domestic resources should be
adequate to supply the domestic component of primary lead demand (2).
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 metal fumes to the atmosphere.
Two slags produced by these operations may be responsible for the discharge
of heavy metals into watercourses.
Unlike the copper industry, most lead smelters are located in areas
where rainfall is greater than evaporation. Surface evaporation of waste-
waters is therefore not possible, and water pollution from these facilities
is a principal concern.
157
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TABLE 3-2. PRINCIPAL STATISTICS FOR THE
PRIMARY LEAD INDUSTRY IN THE UNITED STATES IN 1976 (1)
Primary lead produced, thousand metric tons
Mine (recoverable)
Refinery (refined lead)
Refinery (antimonial lead, lead content)
Exports, thousand metric tons
Lead materials excluding scrap
Imports, thousand metric tons
Ores and concentrates (lead content)
Refined metal
Consumption, thousand metric tons
Reported
552.98
592.32
5.45
5.33
69.28
134.51
1,144.14
158
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TABLE 3-3. U.S. PRIMARY LEAD PRODUCERS (7)
C71
10
Company
AMAX-Homestake Lead
Tollers
ASARCO, Inc.
The Bunker Hill
Company
St. Joe Minerals
Corporation
Location
Boss, Missouri
East Helena, Montana
El Paso, Texas
Glover, Missouri
Omaha , Nebraska
Ke 1 logg , Idaho
Herculaneum, Missouri
Description
Smelter/refinery3
Smelter
Smelter
Smelter/refinery
Refinery
Smelter/refinery
Smelter/refinery3
Capacity,
metric ton/year
127,000
109,000
109,000
100,000
163,000
118,000
204,000
The smelting and refining of Missouri lead ore is conducted in the same plant.
-------
The sintering process produces most of the sulfur dioxide emissions from
lead smelting; lesser amounts are emitted from blast furnaces. None of the
lead smelters exercise complete control of these emissions, and a few are
without controls for S02.
References
1. Mineral Industry Surveys, Lead Industry Monthly. May 1977. U.S.
Department of the Interior, Bureau of Mines. Washington, D.C.
August 1977.
2. Commodity Data Summaries. 1977. U.S. Department of the Interior,
Bureau of Mines. Washington, D.C. 1977.
160
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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.
161
-------
•««•
?,„
Y sotm
Figure 3-1. Lead industry flow sheet.
162
-------
Figure 3-1. Lead industry flow sheet.
163
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 1
Mining
1. Function - Ore deposits containing economically recoverable amounts of
lead are excavated and transported to an ore concentration plant. Most lead
ore is obtained from underground mines that use normal stoping methods (1,2)
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 composition
and lead content, 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 fre-
quently 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 southeastern Missouri, the ore
deposits are characterized by simple mineralization 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 intrusive igneous masses.
. 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 performed under ambient conditions.
4. Utilities - Electricity is used for operation of equipment in under-
ground mining and transport. Diesel fuel and electricity are required for
ore transport equipment 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.
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 observations from
several types of nonferrous mining.
Wastewater from lead mining results from several sources, the worst of
which is probably seepage of surface water through spoil piles; others
include interception of aquifers, and water sent into the mine for utility
purposes (5,6). The water is pumped from the mine at a rate necessary to
maintain mining operations. The required pumping rate bears no relation to
164
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the ore output and is subject to seasonal variation. The rate can range from
a few cubic meters to thousands of cubic meters 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. Analysis of wastewater from
a Missouri and an Idaho mine are given in Tables 3-4 and 3-5. In general,
chemical characteristics of the water are typical of those from any sulfide
mine in the same geographic area. These characteristics are described in
Section 6.
Substantial amounts of solid waste result from underground mining opera-
tions, the estimated average for 1973 being 0.13 ton per ton of ore mined
(4). This waste material consists of the country-rock surrounding 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.
6. Control Technology - Fugitive dust emissions are controlled by the
manual use of water sprays as needed.
Wastewater is generally treated with lime and impounded as practiced in
.copper mining. Management of wastewater is discussed in detail in Section (5.
Since water from Missouri mines is already basic, liming may hot 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.
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 should not con-
taminate a stream or underground aquifer. Prevention of water seepage is
important in this regard.
7. EPA Source Classification Code - None
8. References -
1. Mineral Facts and Problems. U.S. Department of the Interior,
Bureau of Mines. Washington, D.C. 1970.
2. Minerals Yearbook. U.S. Department of the Interior, Bureau of
Mines. Washington, D.C. 1973.
3. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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 Category Volumes I
and II. Environmental Protection Agency. Washington, D.C.
EPA-1-75/032-6. February 1975.
165
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5. Hawley, J.R. The Problem of Acid Mine Drainage in the Provence of
Ontario. Ontario Ministry of the Environment. Toronto. 1977.
6. Williams, R.E. Waste Production and Disposal in Mining, Milling,
and Metallurgical Industries. Miller Freeman Publications, Inc.
San Francisco. 1975.
7. Wixon, B.G., et al. An Interdisciplinary Investigation of Environ-
mental Pollution by Lead and Other Heavy Metals from Industrial
Development in the New Lead Belt of Southeastern Missouri. Univer-
sity of Missouri, Rolla and Columbia, Missouri. June 1974.
8. Ha11 owe!1, 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.
166
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TABLE 3-4. ANALYSIS OF A MISSOURI MINE WATER (7,8)
Component
Concentrate, mg/1
Mercury
Cadmi urn
Chromium
Manganese
Iron
Sulfate
Chloride
Fluoride
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
TABLE 3-5. ANALYSIS OF AN IDAHO MINE WATER (6)
Constituent
pH
Sulfate as S04~
Total iron
Zinc
Nickel
Copper
Manganese
Aluminum
Lead
Cadmi urn
Concentration
2.2
63,000.0
16,250.0
14,560.0
4.8
13.4
2,625.0
347.0
0.8
22.5
Constituent
Magnesium
Calcium
Potassium
Sodium
Chromi urn
Chloride
Nitrate as NOf
Electrical con-
ductivity
(micromhos @ 25°C)
Concentration
1,500.0
31.6
0.7
0.5
0.3
38.0
77.5
48,000.0
167
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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, followed by separation into two or
more fractions. The fractions rich in desired minerals are called concen-
trates, 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 flotation method
achieves separation by the use of compressed air and chemical additives that
create a froth in which finely divided mineral particles 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 are considerable differences in specific
gravities of the ore minerals and the gangue. Since the milled ore particles
need not be as small as those required for flotation, 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 specific gravity between that of the lead mineral and the
gangue. The mineral particles sink, while the gangue floats to the top for
removal 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 produc-
tion. 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.
Ore concentrate from the flotation cells requires dewatering before
shipment to smelters. The slurry is fed to thickeners, and flocculating
agents such as alum are added to improve the settling rate and fines collec-
tion. The thickened slurry of about 50 percent solids is vacuum filtered and
dried to a product containing 6 to 15 percent moisture (1).
Typical analyses of southeastern Missouri and western lead ore concen-
trates are presented in Tables 3-6 and 3-7, respectively.
168
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TABLE 3-6. TYPICAL SOUTHEASTERN MISSOURI LEAD CONCENTRATE ANALYSES (1)
(percent by weight)
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
MqO
0.90
0.75
vo
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TABLE 3-7. WESTERN LEAD CONCENTRATE ANALYSES (4)
Constituent
Pb
Zn
Au
Ag
Cu
As
Percent
45-60
0-15
0-0.05 kg/ ton
0-1.4 kg/ton
0-3
0.01-0.40
Constituent
Sb
Fe
insolubles
CaO
S
Bi
Percent, weight9
0.01-2.0
1.0-8.0
0.5-4.0
tr-3.0
10-30 .
tr-0.1
tr = trace.
170
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2. Input Materials - Lead content of the sulfide ores fed to concentrating
plants ranges from 3 to 8 percent, except for the high-grade Missouri ores in
which lead content exceeds 70 percent (2).
Table 3-8 lists the flotation chemicals and amounts required for pro-
cessing lead ore; included also are some of the less commonly used agents.
3. Operating Conditions - All concentrating operations take place at
atmospheric pressure and ambient temperatures.
4. Utilities - Water usage varies with the degree of processing and is
approximately 4 cubic meters per metric ton of ore processed (3).
Electricity is used to operate grinding equipment and generate com-
pressed air.
5. Waste Streams - Fugitive dust emissions are the only type of atmospheric
pollutant warranting consideration. Compositions of the dust are not speci-
fied. Crushing operations generate, on the average, 3.2 kilograms of particu-
late emissions per metric ton of ore processed; 0.9 kilogram is attributable
to the crushing and grinding operations and 2.3 kilograms to material trans-
port and storage (3,4,5).
Liquid waste from the concentrating operation is in the form of a
tailings slurry discharged to the tailings pond. Approximately 4 cubic
meters of tailings slurry is discharged per metric ton of ore processed
(3,6).
Flotation and conditioning chemicals are present in the wastewater
either as a floating layer or a solute. In general, lead sulfide flotation
is run at an elevated pH level (8.5 to 11) requiring frequent pH adjustments
with hydrated lime or sodium carbonate (7). This alkaline wastewater dis-
solves only small amounts of heavy metals, but can carry mineral particles in
suspension.
Wastewaters leaving a concentrating operation contained metals as shown
in Table 3-9. These were the only metals investigated; others may have been
present in greater than normal concentrations. Concentrations of calcium,
magnesium, sodium, and potassium in mill waters are significantly higher than
those in surface water.
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 reagents. Typical
quantities are 0.9 to 1.1 tons per ton of ore milled. The main component is
dolomite, with small quantities of such constituents as lead, zinc, copper,
mercury, cadmium, manganese, chromium, and iron.
6. Control Technology - Control methods used by copper producers in con-
centrating plants are also used at lead concentrators. Details are given in
Process No. 2 of Section 2, and in Section 6.
171
-------
TABLE 3-8. FLOTATION CHEMICALS (8)
Chemical
Amount used,
kg/metric ton of ore
NdpCOg (conditioner)
CaO (conditioner)
CuS04 (activator)
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
172
-------
TABLE 3-9. LEAD MILL WASTEWATER ANALYSIS (6)
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
173
-------
To preserve the Ozark area of Missouri where the New Lead Belt is
located, special attention has been given to wastewater treatment. Flotation
reagents in the wastewater are biologically degraded by algae growth. Use of
meandering streams before final discharge to receiving waters increases
exposure to the algae and provides good conditions for algae growth. The
algae sink to the bottom of the stream and act as a solids collector, as well
(6). Disposal of algae in event of excessive growth is not discussed in the
literature.
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/l-75/032-a.
February 1975.
2. Mineral Facts and Problems. U.S. Department of the Interior,
Bureau of Mines. Washington, D.C. 1970.
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 Category Volumes I
and II. Environmental Protection Agency. Washington, D.C.
EPA-1-75/032-6. February 1975.
4. PEDCo-Environmental Specialists, Inc. Trace Pollutant Emissions
from the Processing of Metallic Ores. August 1974.
5. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
6. Wixon, B.G., et al. An Interdisciplinary Investigation of Environ-
mental Pollution by Lead and Other Heavy Metals from Industrial
Development in the New Lead Belt of Southeastern Missouri.
University of Missouri. Rolla and Columbia, Missouri. June 1974.
7. Hawley, J.R. The Use, Characteristics and Toxicity of Mine-Mill
Reagents in the Provence of Ontario. Ontario Ministry of the
Environment. Toronto. 1977.
8. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
174
-------
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 amenable to
smelting;
3) To drive off volatile oxides such as SO?, SO,, As90.,, and Sb90.,;
and •* * 6 * J
4) To produce a firm, porous clinker that is easily fed to a blast
furnace (1).
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 impurity
levels of zinc, copper, arsenic, antimony, and bismuth. 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 in diameter.
The pellets are spread evenly over a horizontal metal belt which takes
them through either an updraft or downdraft sintering machine. The two
varieties of sintering machines differ primarily in their combustion air
methods of circulating. As the pellets proceed through the sintering machine,
they are heated and undergo oxidizing reactions that convert sulfides
to oxides and sulfates. 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 millimeters
diameter (2). The crushed sinter is screened for removal of fines, which are
recycled to the charge blending step. The screened product is stored for
blast furnace reduction. Table 3-10 gives typical ranges of components in
the sintered product.
175
-------
TABLE 3-10. 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
176
-------
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 description (Process No. 2). Col-
lected flue dusts, recycled sinter, and smelter residues also are part of the
charge for sintering. 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 as needed. Coke fines, in the amount of 1 percent by weight of the
total charge, are mixed with the charge (5).
Typical feed to a sinter machine is given in Table 3-11.
TABLE 3-11. SINTER MACHINE FEED (10)
Ore concentrate
Misc. lead materials
Flux diluent
Sinter recycle
Weight, percent
31.47
12.44
19.86
36.20
3. Operating Conditions - Temperatures in both updraft and downdraft sinter
roasting machines reach approximately 600°C. Pressure is atmospheric.
4. Utilities - In both updraft and downdraft sintering machines gas- or
oil-fired burners are used to ignite the charge. Energy consumed in the
sintering process amounts to 0.5 millionkilocalories per ton of lead pro-
duced. A breakdown allocates 40 percent to coke consumption and 60
percent to gas or oil consumption, gas being used more than oil (6).
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 general
operating equipment. Approximately 20 percent less power is required for the
updraft fans than for downdraft (7).
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 shows roughly 40 to 70 percent lead, 10 to 20 percent zinc,
and 8 to 12 percent sulfur (8). Depending upon concentrate composition, the
flue dust contains various amounts of antimony, cadmium, germanium, selenium,
tellurium, indium, thallium, chlorine, fluorine, and arsenic (7). Tables
3-12 and 3-13 give weight analysis and size distribution of particulate
emissions.
177
-------
TABLE 3-12. GRAIN LOADING AND WEIGHT ANALYSIS OF INPUT FEED AND
EMISSIONS UPDRAFT LEAD SINTERING MACHINE (11)
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 3-13. TYPICAL SIZE PROFILE OF EMISSIONS,
UPDRAFT LEAD SINTERING MACHINE (11)
Size,
micron
20-40
10-20
5-10
< 5
% weight
15-45
9-30
4-19
1-10
178
-------
Sintering is the only step in the lead smelting process that emits
enough S02 to create a serious air pollution control problem. About 85
percent of the sulfur is removed from the concentrate during sintering.
Approximately 50 percent of the remainder is discharged as S02 from subsequent
operations; the balance goes into the slag as sulfates (7,9).
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 machine,
little SOo is being emitted. If the exit gases are removed in a single
stream, the S0£ concentration is about 2 percent (7,8). The exit gases can
be split into two streams, one predominantly from the front and the other
from the rear. This procedure produces both a weak and a strong S02 stream,
0.5 and 5.7 percent S02 respectively (7). Only updraft sintering machines
have incorporated this engineering modification.
Off-gases also contain organic vapors from flotation reagents or their
combustion products. 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 and ranges from 0.25 to 0.50
normal cubic meters per minute per square meter of bed area (8). Tempera-
tures of the gases normally range from 150 to 400°C (5,8). Flow rates may
vary between 58,000 and 66,000 standard cubic meters per hour (8).
Table 3-14 gives a typical analysis of gases from a sintering machine.
A small amount of arsenic trioxide in the gaseous form is included in these
exit gases; quantities are unknown.
6. Control Technology - The particulate control devices discussed in
Section 2, Process No's. 3 and 6, are also used for treating sintering pro-
cess exhaust gases. 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
3-15 lists current atmospheric controls on lead sintering processes.
The strong gas stream collected from separation of updraft exit gases is
the only stream amenable to sulfuric acid production. In both the weak
stream and the combined single stream, concentrations are too low for such
treatment. At least one smelter has installed a recirculation system for the
weak stream which allows S02 to be removed in the strong stream exit for sub-
sequent treatment. Certain foreign operations have successfully made use of
this technique for some time. A detailed discussion of sulfuric acid produc-
tion is presented in Section 2, Process No. 14.
Three plants are now controlling S02 in sinter machine off-gases by use
of single-contact sulfuric acid plants, which reduce total $62 emissions by
70 to 80 percent.
179
-------
TABLE 3-14. ANALYSIS OF SINTER MACHINE EXHAUST GASES
(MISSOURI LEAD OPERATING COMPANY) (12)
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 .
180
-------
TABLE 3-15. ATMOSPHERIC CONTROL SYSTEMS ON PRIMARY LEAD
SINTERING MACHINES (7)
Plant
Control system
Bunker Hill/Kellogg, Idaho
.AMAX/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 machines. Gases
treated by scrubbers, a spray chamber,
and a baghouse, then out stack.
181
-------
Currently no lead smelters practice control on weak S02 streams. The
best available control technology for these streams would be chemical scrub-
bing, as described in Section 2, Process No. 6.
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/l-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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
4. Hallowell, J.B., R.H. Cherry, Jr., and 6.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. 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.
6. Copper Hydrometallurgy: The Third-Generation Plants. Engineering
and Mining Journal. June 1975.
7. 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.
8. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
9. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry. Volume II - Primary and
Secondary Nonferrous Smelting and Refining. Draft. April 1975.
10. Arthur G. McKee & Co. Systems Study for Control of Emissions
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. June 1969.
182
-------
11. 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.
12. PEDCo-Environmental Specialists, Inc. Trace Pollutant Emissions
from the Processing of Metallic Ores. August 1974.
183
-------
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 descrip-
tion of this process is given in Section 2, Process No. 14.
2. Input Materials - Inputs are the same as those discussed in Section 2,
Process No. 14.
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 scrubbers are part
of the input gas cleaning system.
Liquid waste effluents are treated with lime and by settling in cooling
ponds. Overflows are recycled to slag granulation or discharged (1). Tables
3-16 and 3-17 give the treatments now practiced for control of acid plant
blowdown and scrubber wastewaters, by 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/l-75/032-a.
February 1975.
184
-------
TABLE 3-16. 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
0
273 m3/day
(72,000 GPD)
TABLE 3-17. 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
0
Undetermined
185
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 5
Blast Furnace
1. Function - Sintered feed is reduced in the blast furnace 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 re-
sulting carbon monoxide reduce most of the metallic oxides to yield a molten
mass of metal. Some of the metallic impurities interact with the flux 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, if present in 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 charge 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 3-18.
Upon completion of the process, the crude bullion is charged to dressing
kettles (2), 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
3-19.
The capacity of most large blast furnaces is 1360 metric 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 operations. The
relative amounts of these materials are presented in Table 3-20.
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.
3. Operating Conditions - Temperatures in a blast furnace range from 215°C
for the charge 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 operates at
a pressure slightly above atmospheric.
186
-------
TABLE 3-18. LEAD BULLION COMPOSITION (2,3,4)
Component
Ag
Au
Cu
S
Pb
Fe
Zn
Sn
As
Sb
Bi
Wt. percent
0.13-0.31
1.6-3.1*
1.0-2.5
0.25
94-98
0.6-0.8
tr.
tr.
0.7-1.1
1.0-1.75
0.01-0.03
Value for Au in g/metric ton.
tr. - trace
187
-------
TABLE 3-19. TYPICAL BLAST FURNACE SLAG ANALYSIS (2,3,4)
Component
Ag
Cu
Pb
FeO
CaO
Zn
insol
MnO
As
Sb
Cd
F
Cl
Ge
S
Weight percent
l.'56-4.69a
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
3 Values for Ag in grams per metric ton.
Variable, depending on the furnace charge.
c tr = trace.
Insolubles include MgO - A10 - Si02 phases.
188
-------
TABLE 3-20. TYPICAL BLAST FURNACE CHARGE (2)
Component
Sinter
Coke
Miscellaneous products
(zinc plant residues)
Slag (dross)
Silica
Lime 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
189
-------
4. Utilities - Air is injected into the charge at a pressure of 0.1 to 0.3
kilogram per square centimeter (2). Consumption of 140 to 175 cubic meters
per hour is required for a charge of 1360 metric tons per day (2).
Cooling water circulates through jacketed shafts to control furnace
temperatures. Quantities are unreported.
5. Waste Streams - Particulate emission rates in blast furnace exhaust gas
range from 125 to 180 kilograms per metric ton of bullion produced (3,5).
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 (5,6).
Undiluted gas temperatures are estimated to be 650° to 750°C (3,5), with
theoretical flue gas rates of 170 to 400 normal 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 reported in
Table 3-21.
TABLE 3-21. EXHAUST GAS ANALYSIS AFTER
AIR DILUTION AND CO COMBUSTION (5,7)
Component
co2
°2
C0a
so2
N2
Percent by volume
15
15
5
0.05
Remainder
a CO concentration estimated to be 25 to
50 volume percent prior to combustion.
Other reports indicate 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 concentrate
is removed in the blast furnace, half emitted as S02 and the rest retained in
the slag or matte (8).
Slag from the blast furnace can be discharged and granulated with
cooling water, eliminating the need for a slag fuming furnace.
6. Control Technology - The dilute S0£ emissions from the blast furnace are
not controlled at lead smelters. The best available control technology is
chemical scrubbing.
190
-------
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 3-22 describes controls for blast furnace gases. See Section
2, Process No. 6 for details.
Current practice for slag disposal is to convey it hydraulically with
the granulating water stream to a dump or tailings pond. Recommended tech-
nology includes use of concrete settling pits, ground sealing of disposal
area, and diversion of runoff to a water treatment lagoon.
7. EPA Classification Code - 3-03-010-02
8. References -
1. Arthur G. McKee & Co. Systems Study for Control of Emissions
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. June 1969.
2. PEDCo-Environmental Specialists, Inc. Trace Pollutant Emissions
from the Processing of Metallic Ores. August 1974.
3. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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/l-75/032-a.
February 1975.
5. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
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. August 1962. pp. 657-668.
7. Arthur G. McKee & Co. Systems Study for Control of Emissions
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. 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.
191
-------
TABLE 3-22. ATMOSPHERIC CONTROL SYSTEMS ON PRIMARY
LEAD BLAST FURNACES (7)
Plant
Control system
Bunker Hi 11/Kellogg, Idaho
AMAX/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.
192
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PRIMARY LEAD PRODUCTION PROCESS NO. 6
Slag Fuming Furnace
1. Function - A slag fuming furnace is used to recover metal values other-
wise lost in the slag. Blast furnace slag contains appreciable concentra-
tions of zinc; most lead ores contain this metal and many smelters add it to
the blast furnace charge in the form of residues containing as much as 12
weight percent zinc.
The slag is fed to the fume furnace in a molten state, and pulverized
coal is added to maintain temperature by combustion. By maintaining 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 charged to the
fume furnace is shown in Table 3-19 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 1200°C.
Atmospheric pressure is used (1).
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 design
of cooling water circuit at a given plant. A typical analysis is given in
Table 3-23.
5. Waste Streams - The exhaust gas from the furnace typically has a low SC>2
concentration. The literature (2) cites a value of 0.02 volume percent for a
flow rate of 5,660 normal cubic meters per minute. Gas stream temperature
is about 1200°C.
The exit gas also contains high concentrations of particulate and fume
composed of the volatile components of the blast furnace slag. The litera-
ture does not cite quantities or composition.
The dumped slag and water used for granulation constitute the major
waste stream from the process. The slag is made up of various compounds of
iron, calcium, silicon, aluminum, magnesium, and other elements. The water-
193
-------
TABLE 3-23. 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
Met
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.
Production: 695 metric tons/day.
Source: This contract and 1971 RAPP data.
Notes:
a NLC = no load calculable.
194
-------
soluble portions are leached by the cooling water. Table 3-23 presents
analyses of the intake and outflow streams of water used for slag granula-
tion.
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. Section 2, 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.
The varied treatment and disposal practices for slag granulation water
are summarized in Table 3-25. 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 technology for wastewater treatment is a combination of
neutralization and clarification; the resulting effluent concentrations are
presented in Table 3-24.
TABLE 3-24. EFFLUENT CONCENTRATIONS WITH
NEUTRALIZATION AND CLARIFICATION
Component
Cadmium
Lead
Mercury
Zinc
Concentration,
mg/1
0.5
0.5
0.005
5.0
These values are currently being met by five of the six lead smelters.
Further information is given in Water Management (Section 6).
7. EPA Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Arthur G. McKee & Co. Systems Study for Control of Emissions
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. June 1969.
3. Hallowell, J.B., et al. Water Pollution Control in the Primary
Nonferrous Metals Industry. Volume I, Copper, Zinc, and Lead
195
-------
TABLE 3-25. PRIMARY LEAD SLAG GRANULATION
WASTEWATER TREATMENT (5)
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
196
-------
Industries. Environmental Protection Agency. Washington, D.C.
EPA-R2-73-274a. September 1973.
Vandegrift, A.E., L.J. Shannon, P.G. Gorman, E.W. Lawless, 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.
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/l-75/032-a.
February 1975.
197
-------
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 oxida-
tion. The kettle and molten bullion are cooled to a temperature at which
lead is still a liquid but oxides of the common impurities 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 dressing
kettle. This sulfur combines with the remaining copper forming cuprous
sulfite (CU2S), which floats and is skimmed off with the rest of the dross.
E5y dressing, a bullion with copper content as high as 2 percent can be reduced
to approximately 0.005 percent copper (1).
The dross is sent to a dross reverberatory furnace for further treatment
and recovery of marketable products. Dross may typically contain 90 percent
lead oxide, 2 percent copper, and 2 percent antimony, in addition to 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 is
shown in Table 3-26.
2. Input Materials - During the dressing procedures sulfur is added in a
ratio of approximately 1 kilogram per ton of bullion from the blast and dross
reverberatory furnace. Various amounts of coal or coke, 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 temperature of
-370° to 500dC and maintained within that range. Pressures are atmospheric
4. Utilities - Most of the dressing kettles are heated with natural gas.
About 1.1 million kilocal cries per metric ton are consumed during this stage,
of which 90 percent is allocated to gas and 10 percent to oil (2). Conveyors,
agitators, pumps, and similar equipment are powered by electricity. Air is
injected by submerged lances for supplementing oxidation and agitation (1,5,6).
Quantities of electrical and air consumption are not given in the literaure.
5. Waste Streams - The dressing operation generates small amounts of air
pollutants and slag. The air pollutants are S02 and volatile components of
the lead bullion. A typical analysis of the bullion was presented in Table
3-18. Varying quantities of copper, iron, arsenic, zinc, cadmium, antimony,
198
-------
TABLE 3-26. LEAD BULLION ANALYSIS (1,3,4)
Basis: As drossed
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.
199
-------
and bismuth may be volatilized; it is believed that the quantities are very
small because of the low temperatures. 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 normal cubic meters per
meters per minute. Temperatures of these gases are low, approximately 200°
to 300°C (6). The particulate leading has been quantified by one source as
being between 1.0 and 21.7 grams per cubic meter of off-gas (7). Another
source cites the emission rate as 10 kilograms per metric ton of lead produced.
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.
Particulates and fumes from the dressing kettles 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. Additional information is given in Section 2, Process No.
6.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. 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.
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/l-75/032-a.
February 1975.
4. PEDCo-Environmental Specialists, Inc. Trace Pollutant Emissions
from the Processing of Metallic Ores. August 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. Arthur G. McKee & Co. Systems Study for Control of Emissions
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. June 1969.
7. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
200
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 8
Dross Reverberatory Furnace
1. Function - Dross removed from the lead bullion requires further treat-
ment 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 pro-
duce matte and speiss containing high ratios of copper to lead. Whether
treated or not, the dross is charged into a reverberatory furnace along with
pig iron, silica, and lime rock (optional).
Smelting of the charge separates the products into four layers: slag on
top, matte and speiss intermediate, and molten lead at the bottom. Through
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 copper, 38 percent lead, 11 percent zinc, 11 percent FeO, and 16
percent Si02 (1). 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 copper, 38 percent lead, 16 percent sulfur, 1 percent iron, and 0.6
percent arsenic, plus small amounts of zinc, rare earths, and precious metals
(1). 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, the process requires the addi-
tion 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 atmo-
spheric pressure.
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 pollution
from the dross reverberatory furnace. Particulate emission rates are 10
kilograms per metric ton of lead produced (2). The reference does not
indicate whether this emission includes condensed fume. 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 normal cubic meters
per minute (2,3). The S02 content of this gas is usually below 0.05 percent.,
.Exhaust gases are about 760° to 980°C (2).
201
-------
Water used in matte and speiss granulation is evaporated before trans-
port.
All solids are recycled or marketed.
(5. Control Technology - No control methods are now used for the weak S02
stream emitted from the dressing reverberatory. The best available tech-
nology is chemical scrubbing.
Particulates and fumes from the dressing reverberatory are combined with
the blast furnace off-gas at all plants. As mentioned earlier, all plants
use baghouses. Control efficiency is estimated at 95 to 99.8 percent.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
2. Jones, H.R. Pollution Control in the Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
3. Arthur G. McKee & Co. Systems Study for Control of Emissions
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. June 1969.
202
-------
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 (1).
Roasting, performed in a Godfrey roaster, a special small heated en-
closure, (see Section 2, Process No. 17), causes cadmium, 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 - Operating temperature and pressure are not
reported.
4. Utilities - Equipment can be oil- or gas-fired (1). No quantities are
specified in the literature.
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.
There are no liquid or solid wastes from cadmium recovery.
6. Control Technology - The flue dust and fume emitted from the roaster can
be contained by further cooling with water sprays and collection in a bag-
house.
7. EPA Source Classification Code - None
8. References -
1. Arthur G. McKee & Co. Systems Study for Control of Emission
Primary Nonferrous Smelting Industry. U.S. Department of Health,
Education, and Welfare. June 1969.
2. Hallowell, J.B., et al. Water Pollution Control in the Primary
Nonferrous Metals Industry. Volume I, Copper, Zinc, Lead Indus-
tries. Environmental Protection Agency. Washington, D.C.
EPA-R2-73-274a. September 1973.
203
-------
PRIMARY LEAD PRODUCTIONS PROCESS NO. 10
Reverberatory Softening
'1. Function - The drossed lead bullion is further purified by a process
termed "softening", which entails removal of the elements that make lead hard,
notably arsenic, antimony, and tin. Several other softening processes can be
used (Process No's. 11 and 12) (1,2).
The reverberatory method is similar to the dressing procedure and is
particularly applicable to 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 increase the rate of impurity oxi-
dation (1).
Furnaces with capacities of 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 further treated to recover contained values. The softened lead is tapped
from the bottom of the furnace and pumped to the desilverizing process. Table
3-27 presents typical analyses of the softened bullion product and the soft-
ened 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
3-26 (Process No. 7). Litharge is added only when especially hard bullion is
processed. Coke or coal may be added to inhibit the oxidation of lead.
3. Operating Conditions - Temperatures during softening reach 700°C.
Pressures are atmospheric (1).
4. Utilities - Electricity is the power source for mechanical agitators,
pumps, and conveyors. Most of the heat is supplied by the exothermic oxida-
tion of impurities. Gas or oil is used to begin the reaction and maintain the
temperature. Air is injected through lances or pipes into the bath. Air
consumption 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.
There are no liquid or solid wastes from reverberatory softening.
6. Control Technology - No controls of atmospheric fume emissions are
reported.The exhaust gas could be routed to blast furnace baghouses.
7. EPA Source Classification Code - None
204
-------
TABLE 3-27. TYPICAL COMPOSITIONS OF SOFTENED LEAD BULLION AND SLAG (1)
(AMOUNTS IN WEIGHT PERCENT)
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.
205
-------
8. References -
Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Lead
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Environmental Protection Agency. EPA-440/l-75/032-a.
February 1975.
206
-------
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) (1,3). A slag containing the oxidized impuri-
ties results. 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 3-27 (Process No. 10);
residual hardness is less than 0.03 weight percent antimony equivalent.
2. Input Materials - In addition to the drossed lead bullion, caustic soda
and sodium nitrate (niter) are required for fluxing. Amounts depend upon the
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 atmospheric (1).
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 - Atmospheric emissions containing oxides of nitrogen are
released during kettle softening. Details are unreported.
There are no liquid wastes from this process.
The slag containing 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.
6. Control Technology - The controls used for atmospheric emissions are not
known.
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 softening (Process No. 12) is recommended.
207
-------
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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 Guide-
lines and Proposed New Source Performance Standards for the Lead
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Environmental Protection Agency. EPA-440/l-75/032-a.
February 1975.
208
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 12
Harris Softening
1. Function - In addition to reverberatory and kettle softening, removal of
arsenic, antimony, and tin from drossed lead bullion can also be accomplished
by the Harris process. 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 pyrometallurgical
step is the same as the kettle softening process. The drossed bullion is
charged to a kettle, melted, and agitated. Sodium nitrate and sodium hydrox-
ide 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 process (1).
The second operation is a 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 discarded, and the
solution is cooled to preferentially precipitate sodium antimonate. 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 stoichiometric amounts
for fluxing. Unspecified amounts of process water are fed to the operation
for slag leaching. Lime is required in quantities sufficient to precipitate
salts of arsenic and tin.
3. Operating Conditions - Maximum temperatures are 700°C for the pyrometal-
lurgical operation, 100°C for hydrometallurgical processing, and more than
200°C for sodium hydroxide evaporation (2). Atmospheric pressure is main-
tained during all operations.
4. Utilities - Process equipment, such as agitators, pumps, and conveyors,
are powered by electricity.
Gas or oil is utilized for kettle heating, temperature maintenance, and
sodium hydroxide recovery.
5. Waste Streams - Other than the atmospheric emissions noted for kettle
softening (Process No. 11), the leached slag constitutes the only waste stream
of the process. It is not water soluble (1). Composition and quantity of
slag are not reported.
209
-------
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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
210
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 13
Antimony Recovery
1. Function - Slags from the softening processes (Process No's. 10, 11, 12)
are treated in order to separate and recover the contained mineral values,
notably arsenic, tin, and antimony. Two methods are commonly used to recover
antimony, depending upon the product desired.
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 poten-
tial of the other minerals in the charge is higher, the oxides of lead and
antimony are preferentially reduced. Slag is formed and skimmed off while the
metallic mixture of lead and antimony 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 (1).
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 morz volatile, it is driven off first and is sepa-
rated from the antimony trioxide by selective condensation (2). Collection of
the oxides consists of condensing the volatilized fume and capturing it in an
electrostatic precipitator or a 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, containing appreciable lead values, is returned to the blast furnace
or sintering process.
2. Input Materials - Slag from the softening processes is the main input; it
contains primarily lead, arsenic, antimony, and tin. A typical analysis is
presented in Table 3-27 (Process No. 10). For production of hard lead, coke
or charcoal is used as a reductant, the quantity dependent on feed slag com-
position. 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. Elec-
tricity is required for operation of conveyors. In antimony trioxide produc-
tion, cooling water is required for fume condensation. No quantitative data.
are furnished in the literature.
5. Waste Streams - In the volatilization process, the air stream carrying
the oxide fume is released to the atmosphere after condensation and collec-
tion.
There are no liquid wastes from antimony recovery.
211
-------
The arsenic trioxide, if it cannot be sold, represents the only solid
waste from the process.
6. Control Technology - In the volatilization and condensation process, the
fume stream passes through an electrostatic precipitator or baghouse (3).
Details on these collection devices are given in Section 2, Process No. 3.
No established control for excess arsenic trioxide has been developed.
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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
3. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Lead
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Environmental Protection Agency. EPA-440/l-75/032-a.
February 1975.
212
-------
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. These metals do not oxidize easily
and 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,
forming 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 (1,2). Following removal of this crust, the second addi-
tion 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. Arsenic content
in the bullion must be less than 0.10 weight percent (1).
2. Input Materials - Softened lead bullion is required for the process.
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 3-27, 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 540°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 pumps and agitators. Utility quantities are not given in the
literature.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
213
-------
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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 Guide-
lines and Proposed New Source Performance Standards for the Lead
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Environmental Protection Agency. EPA-440/l-75/032-a.
February 1975.
214
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 15
Retorti ng
1. Function - Crusts from the Parkes process are retorted to recover zinc
for reuse in the desilvering process and to form a purified Dore" metal. Dross
is placed in graphite crucibles of 600 to 640 kilograms capacity; these are
heated in a special Faber du Faur type furnace, and the zinc is vaporized.
The zinc vapor is condensed in a cooling chamber and then tapped to bars or
blocks for recycle. The remaining retort metal assays 140 to 400 kilograms
Dor6 per metric ton of dross (1). Table 3-28 presents a typical analysis of
this retort metal.
Table 3-28. TYPICAL RETORT ANALYSIS (1)
Constituent
Zinc
Arsenic
Antimony
% weight Constituent
15.6-43.8 Copper
1.5-2.5 Tellurium
0.4 Bismuth
1.0 Lead
% weight
1.5-4.0
0.2
0.25
Remaining
percentage
2. Input Materials - Crusts from the desilverizing process are the only
input materials 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 (1). 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. Destruction of baghouse fabric at high temperatures
is prevented by routing the gases through tubular cooling chambers prior to
entering the baghouse. 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.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
215
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 16
Cupelling
1. Function - Retort metal is purified in the process called cupelling. In
a furnace called a cupel, the molten metal charge is successively blown with
air and slagged to remove impurities and produce a relatively pure Dore". The
difference in oxidation potentials of the impurities allows sequential removal
of slags with distinct characteristics.
Zinc, arsenic, and antimony are oxidized first and removed; 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 Dore* until the final stages
of cupelling, when they oxidize to a slag called "coppery litharge" because
the 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 Dor6 output. Usual capacities range from
2,850 to 11,300 kilograms 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 3-28, Process No. 15.
Sodium nitrate and silica flour are added to remove the last traces of
copper and tellurium from the Dore*. Amounts depend upon residual levels of
the impurities.
3. Operating Conditions - Temperature of the cupel reaches 1150°C (1).
Pressure 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 impuri-
ties. A cold air stream is also blown across the face of the bath at a pres-
sure of 70 to 87 grams per square centimeter to cause rippling, which hastens
oxidation. No additional quantitative data are given.
5. Waste Streams - Process exhaust gases range in temperature from 1000° to
1100°C and contain metallic vapors (fume) as well as particulates. Zinc,
lead, arsenic, and antimony comprise the fume. Particulates may contain any
of the components listed in Table 3-28 (Process No. 15). Emission data are
not present in the literature.
6. Control Technology - Several smelters control exhaust gases, cooling them
by passage through tubular cooling chambers before routing them to 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.
216
-------
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
217
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 17
Vacuum Dezinclng
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 surface of the bell.
The product bullion typically is analyzed as less than 0.001 weight per-
cent zinc, 0.0003 weight percent antimony, and 0.15 weight percent bismuth (1),
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°C (1,2) with an operating pressure of 50 to 60 microns absolute (1).
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 temperature requirements.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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.
218
-------
Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Lead
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Environmental Protection Agency. EPA-440/175-032-a.
February 1975.
219
-------
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 Betterton process, can be used, as well as vacuum dezincing (Process No.
17) and Harris dezincing (Process No. 19).
Molten lead is pumped from a heated kettle to 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 analyzed as 99 percent zinc
chloride is obtained. The dezinced lead bullion is sent for debismuthizing 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 con-
tains 0.005 weight percent zinc (1).
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 kilograms per hour into
molten lead recirculated at a rate of 7 to 11 metric tons per minute (1).
3. Operating Conditions - A temperature of 370°C maintains the lead in a
molten condition. Pressure is atmospheric (1).
4. Utilities - The kettle is heated with oil or gas. Electricity is used
for pumps, agitators, and conveyors. No quantitative data regarding consump-
tion are cited.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
220
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 19
Harris Dezincing
1. Function - 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).
The pyrometallurgical equipment is the same as for Harris softening
(Process No. 12), i.e., a charging kettle, a reaction cylinder, 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 in turn reacts with caustic to form sodium zincate.
After 30 minutes of lead recirculation, pumping is stopped, fresh caustic is
added to the cylinder to maintain salt fluidity, and the contents of the
cylinder are emptied into a granulator tank. Fresh molten caustic is again
pumped to the cylinder and 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 complete, the product con-
tains less than 0.001 weight percent zinc and 0.0003 weight percent antimony.
The dezinced lead is pumped from the kettle for debismuthizing or casting.
The spent salts from the granulation tank are treated hydrometallurgi-
cally. After granulation and solution in hot water, sodium zincate hydrolyzes
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 evaporated and the resulting anhydrous
caustic recycled (1).
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.
3. Operating Conditions - A temperature of 540°C is required for the pyro-
metallurgical operation. The temperature for hydroextraction is 100°C.
Evaporation of the sodium hydroxide solution requires temperatures above
200°C. All operations are performed at atmospheric pressure (1).
4- Utilities - Kettle heating and soda evaporation of caustic require gas or
oil as fuel. Electricity is used to operate pumps, agitators, and conveyors.
5. Waste Streams - None
221
-------
6. Control Technology - None
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
222
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 20
Debismuthizing
1. Function - When the dezinced lead bullion contains 0.15 weight percent or
more bismuth, it must be debismuthized before the final refining and casting
process. The debismuthizing procedure is called the Betterton-Kroll process.
Calcium and magnesium are added to the molten lead to form ternary
compounds (e.g., CaMg2Bi'2) with the bismuth (1). The compounds have a higher
melting point than lead, and a lower density. Therefore, when the temperature
of the mixture is reduced to just above the melting point of lead, the metal-
lic compounds solidify to form a dross that can be skimmed from the lead. To
enhance physical separation, antimony or organic agents are sometimes added.
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
is returned to smelting.
2. Input Materials - Dezinced lead bullion fed to the process typically
assays 0.001 weight percent zinc, 0.0003 weight percent antimony, and 0.15
weight percent bismuth (2). The quantities 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 bismuth separation. The literature does not
specify the organic compounds or the amounts.
3. Operating Conditions - The molten lead bath is maintained at 500° to
550°C (1,3) for calcium-magnesium addition and is 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 powered.
5. Waste Streams - None
6. Control Technology - None
7. EPA Source Classification Code - None
223
-------
8. References -
1. Development Document for Interim Final Effluent Limitations Guide-
lines and Proposed New Source Performance Standards for the Lead
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Environmental Protection Agency. EPA-440/l-75/032-a.
February 1975.
2. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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.
224
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 21
Bismuth Refining
1. Function - The dross generated by debismuthizing is processed for bismuth
recovery. The material is placed in a furnace, where it is melted. Chlorine
gas is injected, and the calcium, magnesium, zinc, and lead combine with the
chlorine to form chlorides more readily than does bismuth. The chlorides form
a solid slag that is skimmed from the surface of the molten bismuth. Air is
then blown through the bismuth and a caustic soda flux is added to oxidize any
residual impurities. After slag removal, the metal, which is now 99.99
percent bismuth, is cast into marketable shapes and sold (1).
2. Input Materials - The slag from the debismuthizing process is chiefly
composed of ternary compounds of calcium, magnesium, and bismuth. Chlorine
constitutes about 25 weight percent of the slag charged to the furnace.
Caustic soda flux is used in varying 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
atmospheric (1).
4- Utilities - Gas or oil is used for heating and maintaining temperature.
Electricity is used to run pumps and agitators. Compressed air is furnished
to oxidize impurities. The literature does not state the quantities required.
5. Haste Streams - Exhaust gases to the atmosphere contain chlorine and
fume. No emission quantities were found.
There are no liquid wastes from the process.
Slag composed of chlorides of calcium, magnesium, zinc, and lead is 40
weight percent of the dross fed to the process. Final oxidation generates a
soda slag in unspecified amounts.
6. Control Technology - Control of atmospheric emissions is not practiced,,
Slags are discarded with those generated in smelting. Further informa-
tion is given in Process No's. 5 and 6. The chloride salts contained in the
slag are very water-soluble and easily leached into adjacent water supplies,.
Existing practice does not represent good control technology.
7. EPA Source Classification Code - None
8. References -
1. Encyclopedia of Chemical Technology. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 1967.
225
-------
PRIMARY LEAD PRODUCTION PROCESS NO. 22
Final Refining and Casting
1. 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 skimming, the lead,
assayed as 99.999 percent purity, is reheated and sent to the casting opera-
tion, where it is formed into ingots or pigs. Most casting is performed by
fully automated machines. Slag is recycled to the blast furnace (1).
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 impurity of the lead (1,2). 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°C (1,4). Pressure is atmospheric.
4' Utilities - Gas or oil is required for heating. Pumps and agitators are
electrically operated. The literature gives no data on quantities.
5. Waste Streams - Small amounts of atmospheric emissions are released
during refining and casting operations. Emission factors and constituents
have not been reported.
Direct-contact cooling water becomes contaminated with particulate
matter, including lead and lead oxides.
6. Control Technology - There are no controls on the atmospheric emissions.
Several methods are used to handle the contaminated 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. Interscience Publishers, a
division of John Wiley and Sons, Inc. New York. 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/l-75/032-a.
February 1975.
226
-------
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. PEDCo-Environmental Specialists, Inc. Trace Pollutant Emissions
from the Processing of Metallic Ores. August 1974.
227
-------
SECTION 4
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 this analysis, the
zinc industry is considered in segments: pyrometallurgical zinc production,
electrolytic zinc production, zinc oxide production, and cadmium recovery.
Production of other by-products such as germanium, thallium, gallium,
and indium is not considered a part of this industry because it does not take
place 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 1976, approximately 6,700 people were employed in lead-zinc mining
and milling operations. Zinc smelters, including secondary smelters, em-
ployed 4,100 people in 1976, including workers involved in by-product pro-
cesses. Table 4-1 shows mining, production, and consumption totals for zinc
and cadmium in the years 1972 to 1976. These data indicate that zinc and
cadmium production have remained fairly constant over the past few years.
Consumption has exceeded production; imports and shipments from government
stockpiles have made up the difference. Reduced rate of economic activity
has led to a 24 percent drop in slab zinc consumption in 1973-1976 (1).
Imports of cadmium metal have been increasing.
The leading states for mine production in 1976 were Missouri and Tennes-
see, 17 percent each; New York, 16 percent; Colorado and Idaho, 10 percent
each; and New Jersey, 7 percent. The 25 largest U.S. mines accounted for 91
percent of the zinc ore mined in 1976, and the five largest alone accounted for
43 percent (1). Available information on these mines is given in Table 4-2.
A total of 43 zinc mines operated in 1975 (2).
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.
228
-------
TABLE 4-1. MINING, PRODUCTION, AND CONSUMPTION OF ZINC AND CADMIUM (1)
(metric tons)
1976 (est)
1975
1974
1973
1972
Mine production
480
469
500
479
478
Primary slab
zinc production
530
438
555
583
633
Slab zinc
consumption
1150
925
1288
1504
1418
Cadmium production
(includes secondary)*
2.2
2.2
3.3
3.8
4.1
Cadmi urn
consumption
5.8
3.3
6.1
6.3
6.3
PO
r\>
IJD
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 STATES (2,3)
Mine
Balmat
Bulck
Sterling
Bunker Hill
Eagle
Zinc Mine Works
Blue Hill
Friedensville
New Market
Burg in
Austinville &
Ivanhoe
Star Unit
Leadville
Ground Hog
Jefferson City
Edwartis
Shullsburg
Location
St. Lawrence. N.Y
I t*on Mo
Sussex, N.J.
Shoshone, Id.
Eagle, Co.
Jefferson, Tn.
Hancock, Maine
Lehigh, Pa.
Jefferson, Tn.
Utah, Utah
Wythe, Va.
Sh.osho.ie, 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.
ASARCC
ASARCO
New Jersey Zinc Co.
St. Joe Minerals Corp.
Eagle-Richer Ind. Inc.
Opera-
tions
B
B
B
C
B
3
B
B
B
-
8
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
-
-
Prod.ucts
Lead, zinc
Lead, zinc cone.
refined lead,
sulfuric acid
Zinc ore.iron-
mangenese ore
Lead, zinc, silver
Zinc cone., lead
cone .copper-
silver ore
Zinc
Zinc, copper
Zinc cone.,
agricultural
1 imestone
Zinc cone.
•
Zinc cone. .lead
cone.
Zinc, lead, silver
Lead, zinc, si Tver
Zinc, lead, copper
silver
Zinc conc.,agric.
1'imestone
Lead zinc
-
Mineralization
-
Galena, Sphalerite.
Chalcopyrite
Zinc oxide, and
silicates
Lead-zinc-silver
Marmatite, Galena
-
-
Sphalerite
-
-
Sphalerite, Gelena
-
Sul fides
Sphalerite, Gelena
Sphalerite
-
-
CO
o
-------
TABLE 4-2 (continued).
Mine
Idarado
Bruce
Sunnyside
Rend Oreille
Ozark
Young
Magmont
Viburnum 129
Location
San Miguel, Co.
Yavapai, Az.
San Juan, Co.
Pend Oreille, Wa.
Reynolds, Ho.
Jefferson, Tn.
Iron, Mo.
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
1.042.0CO
-
Products
Zinc, lead, copper
gold, silver,
cadmium
Gold, lead, zinc,
silver .copper,
cadmium
Zinc, lead, silver
Lead & zinc
cone.
Zinc cone., cop-
per cone, silver^
-
Mineralization
Zinc, copper, lead,
gold, silver,
cadmium
Galena, Sphalerite
Lead, zinc
Galena, Sphalerite
Galena, Sphalerite
Chalcopyrite
-
ro
CO
-------
All domestic zinc smelters produce cadmium. Collected flue dusts from
roasting and sintering operations and cadmium-containing materials from
refining and precipitation processes are treated hydrometallurgically 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
primary zinc smelters in the U.S. Waste materials containing these metals
are shipped as intermediate products or are disposed of as waste if the by-
product content is not sufficient for recovery.
U.S. primary zinc demand is expected to increase at an annual rate of
2.6 percent through 1980. The decline in primary zinc production that began
in 1969 started to reverse itself in 1976, with a 21 percent increase over
tge previous year (1).
Primary domestic smelting capacity, however, has declined 47 percent
since 1968, with the closing of eight smelters due to outdated equipment and
environmental problems (4).
Jersey Miniere Zinc is constructing a new electrolytic refinery in
Clarksville, Tennessee, with start-up scheduled for the second half of 1978;
full capacity will be about 90,000 metric tons per year. Most concentrates
will come from the company's new Elmwood and Gordonsville mines in Tennessee;
a third mine is to be developed near Stonewall.
Raw Materials
Zinc is usually found in nature as the sulfiide called sphalerite, which
has a cubic lattice structure and is commonly referred to as zinc blende,
blende, or jack. Zinc content can be 67.1 percent in the pure state. 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 franklinite and zincite are mined for their
zinc content at the New Jersey Zinc Co. mine (5).
What may prove to be one of the five largest massive zinc-copper sulfide
deposits in North America has been discovered near Crandon, Wisconsin. Tests
indicate the presence of about 70 million metric tons of ore, analyzed as 5
percent zinc and 1 percent copper (6).
Iron is the most common impurity or associated metal of zinc, owing to
the chemical similarities and relative ease of substitution in their respec-
tive lattices. 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 sulfides, but the cadmium
content rarely exceeds 1 or 2 percent.
232
-------
TABLE 4-3. COMMON ORES MINED FOR THEIR ZINC CONTENT (5)
ZnO
ZnS04 • 7H20
ZnC03
Zn4Si207(OH)2 •
(ZnJln)O • Fe90,
£ Cf O
2ZnCO, • 3Zn(OH)9
0 £
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
233
-------
In zinc ores, commonly associated non-zinc minerals are calcite (CaCOs),
dolomite (Ca,Mg)C03, pyrite and marcasite (Fe$2), quartz (SiOo), chalcopyrite
(CuFeS2), and barite (BaS04).
Zinc ores are processed at the mine to form concentrates containing
typically 52 to 60 weight percent zinc, 30 to 33 weight percent sulfur, and
4 to 11 weight percent iron (7). Roasting at the plant lowers the sulfur
content to about 2 percent. Other raw materials are required at the smelter
for producing metallic zinc. Coke or coal and sand along with inert mate-
rials are required during pyrometallurgical sintering, in quantities depend-
ing 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 pyrometal-
lurgical 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 kilocalories depending on the process. In 1972, 9.6
billion kilocalories were used in the manufacture of primary zinc slabs, an
average of 15.3 million per metric ton (8). This represents a 9 percent
increase in efficiency of energy utilization over 1967, primarily due to the
closing of inefficient horizontal retorts.
Products
The principal 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 compounds, for
alloying, and for grinding into zinc dust. Table 4-4 shows U.S. consumption
of slab zinc for 1976. 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.
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 (10).
Companies
Capacities of the six primary zinc smelters and one zinc oxide plant
that use primary concentrate feed are listed in Table 4-6. The three pyro-
metallurgical plants range in capacity from 57,000 to 227,000 metric tons per
year. U.S. primary zinc capacity is about equally divided between pyrometal-
lurgical and electrolytic processes. The single largest plant accounts for
35 percent of the total domestic primary capacity. All of the primary
234
-------
TABLE 4-4. U.S. SLAB ZINC CONSUMPTION - (1.976) (9)
Galvanizing
Brass and bronze product;
Zinc-base alloy
Rolled zinc products
Zinc oxide
Other
Total
Metric tons
342,893
150,817
387,403
27,088
35,405
35,287
1,028,893
Percent
38
15
38
3
3
3
100
TABLE 4-5. GRADES OF COMMERCIAL ZINC
Consumption, percent weight
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
Cadmi urn
0.003
0.03
0.40
0.50
0.50
235
-------
TABLE 4-6. PRIMARY ZINC PROCESSING PLANTS IN THE UNITED STATES (12,13,14)
Company
Location
Description
Capacity
metric
ton/year
Date plant
built
CO
AMAX, Inc
ASARCO, Inc.
The Bunker Hill Co.
National Zinc Co.
New Jersey Zinc Co.
St. Joe Minerals Corp.
East St. Louis, Illinois
Corpus Christi, Texas
Columbus, Ohio
Silver King, Idaho
Bartlesville, Oklahoma
Palmerton, Pennsylvania
Monaca, Pennsylvania
Electrolytic
Electrolytic
Pyrometallurgical
Electrolytic
Electrolytic
Pyrometal1urgi cal
Pyrometallurgical
76,000
98,000
20,000
98,000
45,000
103,000
227,000
1929
1942
1967
1928
1977
1899
1930
-------
smelters produce sulfuric acid as a by-product. In addition, three of these
plants were responsible for over 75 percent of primary cadmium production in
1976 (1). Salient statistics for the domestic primary zinc industry in 1976
are presented in Table 4-7.
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 smelters. In
1969, zinc emissions to the atmosphere totalled 65 metric tons from mining
and milling, and 45,350 metric tons from metallurgical processing (11).
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 noncontact and contact. Non-
contact 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 pollutants of concern are primarily
zinc and sulfates, accompanied normally by such elements as cadmium and lead,
and small amounts of arsenic and selenium (9).
Limitations on liquid effluent and atmospheric emissions from new
sources have been promulgated (12,13). The solid waste problem is currently
under investigation.
Bibliography
1. Commodity Data Summaries, 1977. U.S. Department of Interior.
Bureau of Mines. Washington, D.C. 1977.
2. 1975 E/MJ International Directory of Mining and Mineral Processing
Operations. Mining Informational Services of the McGraw-Hill
Mining Publications. New York. 1975.
3. McMahon, A.D., et al. Bureau of Mines Minerals Yearbook, 1973.
U.S. Department of the Interior. U.S. Government Printing Office.
1973.
4. Deane, G.L., et. al. Cadmium: Control Strategy Analysis. GCA
Corporation Report No. GCA-TR-75-36-G. U.S. Environmental Protec-
tion Agency. Research Triangle Park, North Carolina. April 1976.
p. 157.
237
-------
TABLE 4-7. PRINCIPAL STATISTICS FOR THE PRIMARY ZINC
INDUSTRY IN THE UNITED STATES IN 1976 (9)
(metric tons)
Production:
Mine, recoverable zinc
Smelter, slab zinc
Imports:
Ores and concentrates
(dutiable zinc content)
Slab zinc
Consumption:
Slab zinc
Ores (recoverable zinc content)
Exports:
Slab zinc
439,550
510,218
88,103
648,184
1,028,893
91,846
3,187
238
-------
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. U.S. Department of the Interior, Bureau of Mines. Minerals and
Materials/A Monthly Survey, 55. July 1977.
7. 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.
8. Fejer, M.E., and D.H. Larson. Study of Industrial Uses of Energy
Relation to Environmental Effects. EPA-450/3-74-044. U.S. Environ-
mental Protection Agency. Research Triangle Park, North Carolina.
July 1974.
9. U.S. Department of Interior, Bureau of Mines. Mineral Industry
Surveys. Zinc Industry in July 1977. October 4, 1977.
10. Jones, H.R. Pollution Control in the Nonferrous Metals Industry
1972. Noyes Data Corporation. Park Ridge, New Jersey.
11. W.E. Davis and Associates. National Inventory of Sources and
Emissions Section V, Zinc. Report No. APTD-1139. National
Technical Information Service, U.S. Department of Commerce.
Springfield, Virginia. May 1972.
12. Background Information for New Source Performance Standards:
Primary Copper, Lead, and Zinc Smelters. Office of Air and Waste
Management. U.S. Environmental Protection Agency. Research
Triangle Park, North Carolina. EPA-450/2-74-002a. October 1974.
13. 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 Cate-
gory. U.S. Environmental Protection Agency. Washington, D.C.
EPA-440/1-75/032. November 1974.
14. American Bureau of Metal Statistics. Nonferrous Metal Data 1975.
New York, New York. 1976.
239
-------
\AT
2
1
xy
FLOTATION
ADDITIVES
Figure 4-1. Zinc industry flow sheet.
240
-------
COAL
SAND
j- OTHER ZIHC-BEARING
MATERIALS
ELECTROLYSIS
13
• COLLOID
• SODIUM CARBONATE OR
BARIUM HYDROXIDE
A
SB
s~
V
- ALLOYING MATERIAL
- FLUXES
tf
CADMIUM
LEACH )5
- HjSOl
- SODIUM CHLORATE
-SPENT ELECTROLYTE
-CAUSTIC
-PRECIPITATORS
'-SODIUM BICHROMATE
A
_ i§§ „ CADI
§">5 PRECIP
-S T
v$P A ^r
UUM -^
TATION *g
\ ~ / h ZINC OUST \ /
\J *- PRECIPITATORS \J
CADMIUM
PURIFICATION
AND CASTING 17
a
— •
- LEAD SMELTER
BAGHOUSE DUST
-LIME
- COAL OR COKE
- H2S04 OR RETURN
ELECTROLYTE
- SODIUM BICARBONATE
- HYDROGEN SULFIDE
L PRECIPITATORS
. BY-PRODUCT RECOVERY
Figure 4-1. Zinc industry flow sheet.
241
-------
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 organizations. This industry
segment analysis examines each production operation, to define its industrial
purpose and its potential and practice in affecting 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. Figure 4-1 is a flowsheet that shows these
processes, their interrelationships, and their major waste streams.
242
-------
PRIMARY ZINC PRODUCTION PROCESS NO. 1
Mining
1. Function - Mining involves excavation and treatment of zinc ore deposits.
Most zinc is mined underground using open shrinkage, cut-and-fill, or square-
set stoping methods. 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 supported and the entire ore body extracted.
A few mines, particularly in early stages of operation, use open pit methods,
which closely resemble those used in copper mining.
Mining operations consist of drilling, blasting, and removing the broken
rock. After removal from the mine, the ore is loaded onto trucks for trans-
portation to concentrating facilities. Mining operations range in size from
those 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 utilize "trackless" mining techniques, 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 some-
times feasible to mine ore containing 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. Condi-
tions depend upon type of mining, i.e., open pit or underground.
4. Utilities - Fuels and electricity are used for operating mining equipment
and transporting ore to concentrating facilities. One estimate for electrical
usage for mining ore in 1973 was 2.7 x 1011 kilocalories (1). With the 1973
mine production level of 434,000 metric tons (2), electrical usage was about
720 kilowatt-hours per metric ton of zinc moved. Unspecified amounts of
water are pumped into mines for machinery and hydraulic backfill operations.
5. Waste Streams - 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 estimated at 0.1
kilogram per metric ton of ore mined (5). Total emissions of cadmium to the
atmosphere were thus 240 metric tons for 1968 (5) and 220 metric tons for
1973 (6).
Mine waste water results from infiltration of ground water, water
pumped into the mine for machines and hydraulic backfill operations, and
infiltration of surface water. Quantities of effluent are not necessarily
related to the quantity of ore mined. The water required to maintain opera-
tions 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
243
-------
hydraulic fluid. Dissolved solids found in the wastewater are generally
lead, zinc, and associated minerals (7).
Conditions compatible with solubilization of certain metals, particu-
larly 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.
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 quantitative or
qualitative estimate was found concerning these spoils.
6. Control Technology - Fugitive dusts from drilling, conveying, and
crushing can be reduced by wetting and control systems (8).
Control of water from this process is described in Section 6, Water
Management. 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 facil-
ity, or inability to reuse all of the mine waste water at a particular mill.
Small 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 for later use 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 limestone.
Acid mine water containing solubilized metals may be effectively treated in
the mill tailings pond. The water may be further treated by lime clarifica-
tion and aeration.
The solid wastes from mining operations are disposed of in a spoil pile
or pond. These wastes can be used as mine backfill.
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.
244
-------
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. W.E. Davis and Associates. National Inventory of Sources and
Emissions - Cadmium, Nickel, and Asbestos - 1968. Cadmium, Section
I. 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 Protec-
tion Agency. Research Triangle Park, North Carolina. April 1976.
7. Development Document for Interim Final and Proposed Effluent Limita-
tions 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.
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.
Draft. April 1976.
245
-------
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 conditioned 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 differences 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 (1). 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. Upon 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 concentrates, 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 flotation. 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 smelters. Treatment of such ores
was the primary reason for development of electrolytic zinc recovery methods
in North America.
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 quantities of lead,
cadmium, copper, and other elements (2). Table 4-8 lists some of these
elements of ore concentrate.
246
-------
TABLE 4-8. RANGE OF COMPOSITIONS OF ZINC CONCENTRATES (6,7,8)
Constituent
Pb
Zn
Au
Ag
Cu
As
Sb
Fe
Insolubles
CaO
S
Bi
Cd
Percent, weight
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
247
-------
Additional details regarding concentrating equipment and procedures are
given in Section 2, 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 concen-
trate produced varies with the zinc content of the ore. A common range is 5
to 10 tons of ore per ton of zinc concentrate (1).
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 encour-
ages 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 4-9 lists commonly used reagents.
Additives are an estimated 1.9 kilogram of lime, 0.4 kilogram of copper
sulfate, 0.04 kilogram of Z-5 xanthate, and 0.02 kilogram of pine oil per
metric ton of concentrate (3).
3. Operating Conditions - Flotation machines are operated at ambient
temperatures and pressures. Flotations are generally run at elevated pH
values of 8.5 to 11 (1).
4. Utilities - Electrical energy is consumed during milling operations.
One estimate for electrical usage in the milling of zinc ores in 1973 is, 290
million killowatt-hours (4). At the 1973 mine production level of 434,000
metric tons (5), electrical usage was about 477 kilowatt-hours per metric ton
of zinc produced.
The water requirement ranges from 330 to 1,100 cubic meters per metric
ton of ore processed per day (1). Feed water for the mills is usually taken
from available mine waters.
5. Waste Streams - It is estimated that 1 kilogram of particulate is
emitted per metric ton of ore processed during crushing and grinding opera-
tions. 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 concen-
tration, 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.
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 cubic meters per
metric ton (1). Tailings slurry discharge is about 4 cubic meters per
metric ton of ore processed.
The raw wastewater from a lead/zinc flotation mill consists principally
of the water used in the flotation circuit, along with any housecleaning
water. The waste streams consist of the tailings streams (usually the under-
flow of the zinc rougher flotation cell), the overflow from the concentrate
thickeners, and the filtrate from concentrate dewatering.
248
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TABLE 4-9. TYPICAL FLOTATION REAGENTS USED FOR ZINC
CONCENTRATION UNITS (1)
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
249
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The principal characteristics of the waste stream from mill operations
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 concen-
trates.
(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 cubic
meters per metric ton (1). The main component of the waste is dolomite.
Table 4-10 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 mineralogy, grinding practices, and reagents.
6. Control Technology - Participate 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.
250
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TABLE 4-10. RANGES OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE LOADS FOR FIVE SELECTED MILLS (1)
ro
tn
Parameter
PH3
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and grease
NBAS 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,
kg/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 limit
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.
-------
Water separated from the concentrates is often recycled in the mill, but
it may be pumped to the tailings pond, where primary separation 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 effluent, is reuse
of the water in other facilities as either makeup water or process water.
Usually, some treatment is required before reuse of this decanted water.
Treatments include secondary settling, phosphate or lime addition, pH adjust-
ment, flocculation, clarification, and filtration.
The most frequently used control technology is the use of a settling or
sedimentation pond system consisting of primary tailings pond and 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 (1).
Control of wastewater is further discussed in Section 6, Water
Management, which presents data typical for most of the zinc industry.
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 backfilling
(1). The most effective means of control is impoundment with isolation of
disposal sites from surface flows. Techniques include the following (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 ground-
water system.
(2) Compaction of waste material to reduce infiltration.
(3) Maintenance of uniformly sized refuse to enhance good compaction
(which may require additional crushing).
(4) Construction of a clay liner over the material to minimize infil-
tration. 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 disposal
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.
252
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7. EPA Source Classification Code - None
8. References -
1. Development Document for Interim Final and Proposed Effluent Limita-
tions Guidelines and New Source Performance Standards for the Ore
Mining and Dressing Industry. Point Source Category Vol. 1.
Publication EPA-440/1-75/061. Effluent Guidelines 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 Metallurgical
Industries. EPA-450/3-73-002. Office of Air and Water Programs,
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. December 1973. pp. 269-309.
3. Denver Equipment Company. Mineral Processing Flowsheets. Denver,
Colorado. 1962.
4. Dayton, S. The Quiet Revolution in the Wide World of Mineral
Processing. Engineering and Mining Journal. June 1975.
5. Commodity Data Summaries 1976. U.S. Department of the Interior,
Bureau of Mines. Washington, D.C. 1976.
6. Water Pollution Control in the Primary Nonferrous - Metals In-
dustries. Volume 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.
7. 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.
8. Katari, V., et al. Trace Pollutant Emissions from the Processing
of Metallic Ores. EPA-650/2-74-115. U.S. Environmental Protection
Agency, Office of Research and Development. Research Triangle
Park, North Carolina. October 1974.
253
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PRIMARY ZINC PRODUCTION PROCESS NO. 3
Ferroalloy Production
1. Function - One U.S. zinc smelter recovers a ferro-manganese by-product
from zinc oxide ore by means of reduction, vaporization, and oxidation in
Waelz kilns. The ore, from a New Jersey mine, is not amenable to flotation.
The zinc oxide fume produced in the kilns is collected and sintered in the
same manner as roasted sulfide concentrates.
The Waelz kilns in use are rotary kilns similar to those used in the
cement industry. Their sizes range from 3.0 meters in diameter by 42.7
meters long to 3.7 meters in diameter by 48.8 meters long. Gas flow is
countercurrent to the solids flow, with the zinc oxide fume and kiln gases
withdrawn at the feed end. Automatic 10-compartment dust tube collectors
remove zinc oxide from the gas streams, which have been previously cooled by
dilution and radiation loss. Solid residues from the kiln contain 20 to 24
percent iron and approximately 10 percent manganese; they are processed in
open-arc electric furnaces to produce a nominal 20 percent manganese iron
(Spiegeleisen) (1).
2. Input Materials - The charge to the Waelz kilns consists of the oxide
ores which are mixed with anthracite coal to raise the carbon content and
limestone to stiffen the charge. Feed to the kilns ranges between 9 and 15
metric tons per hour (1).
3. Operating Conditions - Waelz kilns operate at atmospheric pressures (2).
Operating temperatures are not known.
4. Utilities - Heat generated by the oxidation reactions is usually suf-
ficient to sustain reduction. Additional heat if required for reduction of
the zinc can be supplied by the combustion of coal, oil, natural gas, or
electric furnace gas. No information is available as to the quantities of
additional heat that may be supplied by combustion of fuel.
5. Waste Streams - Unknown quantities of combustion products are emitted
from the kilns and vented to the atmosphere.
There are no liquid or solid wastes from the process.
6. Control Technology - There are no known environmental controls on the
Waelz kilns used for ferroalloy production.
7. EPA Source Classification Code - None
254
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8. References -
1. The New Jersey Zinc Company. Manufacturing Operations. Palmerton,
Pennsylvania.
2. Schlecten, 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.
255
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PRIMARY ZINC PRODUCTION PROCESS NO. 4
Multiple-Hearth Roasting
1. Function - Multiple-hearth roasting is a high-temperature process that
removes sulfur from the concentrate and converts zinc to an impure zinc oxide
called calcine. Roasting may also be accomplished by suspension (process No.
5) or fluidized-bed (Process No. 6) units. In a multiple-hearth roaster, the
concentrates drop from hearth to hearth. As much as 20 percent of the cad-
mium present in the zinc concentrate may be vaporized (1). Any mercury
present is volatilized and enters the gas stream. Multiple-hearth roasters
are the oldest type of roaster in use in the United States.
At present, only two domestic primary zinc smelters use multiple-hearth
roasters. At one of these plants, the multipe-hearth roaster is used in con-
junction 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. Mercury is recovered from the
flue dusts captured in the electrostatic precipitators used to clean the gas
stream from these roasters.
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 + 2S02
2S02 + 02 -»• 2S03
ZnO + S03 •»• ZnSO.
In pyrometallurgical reduction only the oxide state is desired, whereas in
electrolytic reduction, small amounts of the sulfate state are acceptable
(3).
Elimination of the remaining small percentages of sulfur requires a long
residence time. Hence a multiple-hearth-type roaster that eliminates 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 concentrate per day to about 2 percent
sulfur.
Because concentrations of SO? 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.
The roaster consists of a brick-lined cylindrical steel column with 9
or more hearths. A motor-driven central shaft has two rabble arms attached
256
-------
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 spiral fashion and are discharged at the bottom. Additional fuel must
be added to maintain combustion.
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 cadmium are removed preferentially. This is help-
ful when cadmium is to be recovered from the flue dust, since there is less
zinc dust contamination (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 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 required 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 sin-
tering. Specific quantities have not been reported.
3. Operating Conch'tions - 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 initially to
bring the charge up to reaction temperature. About 280,000 kilocalories per
metric ton of ore are required for ignition. The primary fuel is natural
gas. In some multiple-hearth furnaces, when concentrations of less than 1.0
percent of sulfide sulfur are reguired, about 1.1 million kilocalories 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. Electric-
ity powers the rabble arms.
257
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5. Waste Streams - In a zinc smelter the roasting process is typically
responsible for more than 90 percent of the potential SOg emissions; 93 to 97
percent of the sulfur in the feed is emitted as sulfur oxides. Concentrations
of SOg in the off-gas vary with the type of roaster operation. Off-gases
contain up to 6 to 7 percent sulfur, depending on the sulfur content of the
feed. The volume of off-gases ranges from 140 to 170 cubic meters per minute
(6) for furnaces currently in use (approximately 180 metric tons per day).
Typical S02 concentrations range from 4.5 to 6.5 percent (3). 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 dis-
tribution, and equipment configuration. Particulate emissions consist of
fumes and dusts composed of the zinc concentrate elements in various combina-
tions. 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;
arid 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 parti-
culate carryover, depends primarily on the concentrate composition and opera-
ting 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.
Information on liquid and solid wastes is not available at this time.
6. Control Technology - Gases leaving the roaster are routed directly to
"hot" Cottrells, operating at 200° to 220°C, and then to the acid plant (2).
7. EPA Source Classification Code - Roasting/multiple-hearth 3-03-030-02.
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 Cate-
gory. EPA-440/1-75/032. Effluent Guidelines Division Office of
Water and Hazardous Materials. U.S. Environmental Protection
Agency. Washington, D.C. 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.
258
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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. 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.
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. Environ-
mental Protection Agency. Research Triangle Park, North Carolina.
July 1974.
6. Field Surveillance and Enforcement Guide for Primary Metallurgical
Industries. EPA-450/3-73-002. Office of Air and Water Programs,
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. December 1973.
259
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PRIMARY ZINC PRODUCTION PROCESS NO. 5
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. Roasting in suspension promotes better heat transfer than multiple-
hearth roasting and thereby increases reaction rates for desulfurization.
The chemical reactions occurring in the two processes are the same, and the
S0;2 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 desulfuriza-
tion. 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 sulfide content of the ore
is too low.
The roaster consists of a refractory-lined cylindrical steel shell with
a large combustion space at the top and two to four hearths in the lower
portion, similar to those of a multiple hearth furnace. Because the feed
must be carefully sized, additional grinding may be needed for proper prepara-
tion. In more recent models of flash furnaces, concentrate is introduced
into the lower one or two hearths to dry before final grinding in an auxiliary
ball mill and introduction into the combustion chamber.
About 40 percent of the roasted product settles out on a collecting
hearth at the bottom of the combustion chamber. This coarser material 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 product (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 remainder (1).
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 con-
centrate per day (1).
260
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2. Input Materials - Zinc concentrate is the input material for suspension
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 (1). 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 maintains opera-
ting temperatures. About 280,000 kilocalories per 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, pumps, and rabble arms.
5. Waste Streams - The S02 concentration in the off-gas from suspension
roasting is higher than that in multiple-hearth processes, averaging 10 to 13
percent (5). It also contains oxygen, nitrogen, carbon dioxide, and water
vapor. The higher SOg content increases the efficiency of sulfuric acid
production. Emission of particulates depends on operating conditions,
averaging about 6 percent of the feed. These emissions consist of dust and
metal fumes, depending on the concentrate composition and operating condi-
tions. The volume of off-gases ranges from 280 to 420 cubic meters per
minute (5).
There are no process water or solid wastes. There is a boiler blowdown
from the waste heat boiler. Solids are recycled to recover by-product metals.
6. Control Technology - The S02 stream is concentrated enough to allow
sulfuric acid production, discussed in detail in Section 2, Process No. 14.
The roaster off-gases are cooled to about 400°C by heat transfer 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 to 700 cubic meters per minute, and the S0? concentration is 6 to 8
percent (6).
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.
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 Protection
Agency. Research Triangle Park, North Carolina. October 1974.
261
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3. 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 Cate-
gory. EPA-440/1-75/032. Effluent Guidelines Division Office of
Water and Hazardous Materials. U.S. Environmental Protection
Agency. Washington, D.C. 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, North Carolina. July 1974.
5. Field Surveillance and Enforcement Guide for Primary Metallurgical
Industries. EPA-450/3-73-002. Office of Air and Water Programs,
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. December 1973.
6. Jones, H.R. Pollution Control in The Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
262
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PRIMARY ZINC PRODUCTION PROCESS NO. 6
Fluidized-Bed Roasting
1. Function - The fluidized-bed roaster is the newest method for removing
sulfur and converting zinc to calcine. In this roaster, finely ground sul-
fide 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 pro-
duces 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 arseno-
pyrite gold ores; several North American zinc smelters have adopted it in
different forms for use in pyrometallurgical and electrolytic 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, feed to the roasters is pelletized. The calcined
product is also pelletized, eliminating the need for further agglomeration by
sintering. Fluid column roasting operates at slightly higher temperatures
than fluidized-bed.
In the fluidized-bed process, no additional 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 manually or automatically,
via water injection. Relatively low, uniform operating temperatures appear
to lessen the formation of ferrite. The temperatures 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. Efficiency of the operation is maximum when 20 to 30 percent
excess air is supplied over stoichiometric requirements for oxidation of both
sulfur and metals of the charge (1).
Dust carryover into the dust collecting system is somewhat less than in
flash roasting. Particulates emitted average 50 to 85 percent for fluidized-
bed and only 17 to 18 percent for fluid column (2). Amounts vary with the
feed rate (and consequently with the air rate), and with the size of the
material being roasted.
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 S02 concentration achievable is 14.6 percent
when roasting 100-percent zinc sulfide concentrates to completion, unless the
air supply is enriched with oxygen (1).
Although the preferred method of regulating temperature within the bed
is by water injection with automatic thermocouple-operated control, water
263
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injection and slurry feeding can be eliminated when it is desirable to mini-
mize the water content of the gas.
The major advantage of the fluidization roaster is its ability to
process 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 processes; 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 (3).
Feed capacities can range up to 320 metric tons of concentrate per day
(3).
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 usually finely ground at one plant to
90 percent minus 0.044 millimeter (4). However, in some modifications, the
feed is pelletized to provide longer retention time in the roaster (5).
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 kilo-
calories per metric ton of ore is required for ignition (6). 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. Haste Streams - Typical SOp 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. The volume of
off-gas produced ranges from 170 to 280 cubic meters per minute (5). Temper-
atures are approximately 950°C.
As with the other processes, the amount and composition of particulates
and metal fumes depend on the concentrate composition and operating condi-
tions.
As with other roasting processes, solids are recycled to recover by-
product metals and there are no solid or process-water wastes.
264
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6. Control Technology - Emission controls are the same as for the flash
roasting process.The SO2 stream is used for sulfuric acid production.
Since SC^ 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 captured by the
cyclones, and 3 percent entered the hot Cottrell electrostatic precipitator
with the flue gases (!)• Flue dusts are not recirculated to the reactor,
being sufficiently low in sulfide sulfur. Roasters with pelletized feed
yield about 80 percent to the overflow and 20 percent carry-over 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. Encylcopedia of Chemical Technology. Inter-
science Division of John Wiley and Sons, Inc. New York. 1967.
2. Field Surveillance and Enforcement Guide for Primary Metallurgical
Industries. EPA-450/3-73-002. Office of Air and Water Programs,
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. December 1973.
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 Zinc:
Segment of the Nonferrous Metals Manufacturing Point Source Cate-
gory. Battelle Columbus Laboratories. EPA Contract Environmental
Protection Agency. Research Triangle Park, North Carolina.
December 1973.
5. Van Den Neste, E. Metallurgie-Hoboken-Overpelt's Zinc Electro-
winning Plant. CIM Bulletin. Sixth Annual Hydrometallurgical
Meeting. 1977.
6. 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.
265
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PRIMARY ZINC PRODUCTION PROCESS NO. 7
Sintering
1. Function - Sintering has two purposes: first, to volatilize lead and
cadmium impurities and discharge them into the off-gas 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 indus-
try. 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 sustained by sup-
plying air to the pellets. Temperature control is achieved by limiting the
coke and coal content and sulfur content of the sinter mix. Once oxidation
is started, it becomes self-sustaining. Air flow regulation provides addi-
tional 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 4-11.
Estimates are that 80 to 90 percent of the cadmium and 70 to 80 percent of
the lead are removed from the sinter feed (2). 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 cadmium.
The lower portion of the bed, which was not removed by the scalper, is
discharged to a set of crushers. 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 (2).
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 a great amount
of confining pressure from the overburdening charge. Mechanical collapse is
prevented by close chemical control of the nonvolatile ingredients and addi-
tion of silica to increase hardness and strength of the sinter mass. In
horizontal-reduction-type retorts, no longer in use in this county, a soft,
friable sinter was usually desirable. Structural strength was not needed
because of the lack of heavy overburdening pressures. Cadmium content of the
dust is usually 1 percent, allowing profitable cadmium recovery.
A process known as "sinter slicing" may be utilized in Dwight-Lloyd
machines where higher grades of zinc are sought or recovery of impurities 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 lead and cadmium sulfides at the bottom of the
266
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TABLE 4-11. PRODUCT SINTER COMPOSITION (3)
(percent weight)
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
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
ro
cr>
-------
sinter cake. The finished sinter is sliced off for further refining in zinc-
reduction retorts. The lower section of the sinter cake, with its high
cadmium content, passes on to the discharge end of the machine. After
crushing it is returned to the charge.
Dust collected in a baghouse still contains incompletely oxidized mate-
rials such as submicron 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 han-
dling a more finely divided charge. Such a charge consists of about 2.8-
percent sulfur, of which about 2.4 percent is sulfate. These 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 structure which resists
disintegration as well as keeping the reductant and zinc oxide in close
contact. Heat is generated by burning 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 reduc-
tion (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 kilograms of sand and 80 kilograms of coke breeze (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.
In typical zinc sintering operations, charging capacities range from 220
to 550 metric tons per day; 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 kilocalories per metric
268
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ton of concentrate processed (8), although one plant reported use of 51,000
kilocalories per 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 pollutants in
the form of particulates and SO?. The SO? concentration in off-gas from the
sinter machine is very low, 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, sulfur emissions from a sinter machine are pri-
marily 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 off-gas stream contains approximately 1000 ppm
SO?; the concentration, however, can range from 400 to 3000 ppm SO?, depending
upon the total sulfur content of the feed stock (7).
All of the particulate matter is less than 10 microns. Solids loadings
range from 9 to 100 grams per cubic meter (9). Emissions consist of dusts
and metallic fumes of a composition similar to that of the calcine. Typical
chemical composition of particulates is 5 to 25 percent zinc, 30 to 55 percent
lead, 2 to 15 percent cadmium, and 8 to 13 percent sulfur (9). Other con-
stituents include copper, arsenic, antimony, bismuth, selenium, tellurium,
and tin. 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. An
analysis of particulate emissions from a sinter machine is presented in Table
4-12.
In the briquetting step used at one plant, the coker combustion products
are released through an uncontrolled stack and contain sizable quantities of
metallic fume (10). Data on particulate emissions from this process are
presented in Table 4-13.
Gas rates vary widely. With a calcined feed, exhaust gas rates vary
from 42 to 72 cubic meters per square meter of grate. The gas rate with a
concentrate feed is 5.4 to 6.5 cubic meters per square meter (9). Tempera-
ture of combined exit sinter gases varies from 160°C to 380°C in different
plants (9). In addition to sulfur oxides, the gases contain air, water
vapor, carbon dioxide, and traces of other gases.
The sinter process includes no sources of process liquid or solid
wastes. 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 SO? stream from the sintering operation. Best available control tech-
nology is application of chemical scrubbing techniques. These are described
in Section 2, Process No. 6, which discusses scrubbers used for control of
SO? from reverberatory smelters and similar processes. Exit gases may be
cooled by air dilution and water sprays in preparation for gas cleaning.
269
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TABLE 4-12. PARTICULATE EMISSIONS FROM A ZINC SINTERING PROCESS (10)
Cyclone and
ESP control
Cyclone
control
only
Emission factors
(kg/metric ton zinc)
Emission rates
(kg/hr)
Emission factors
(kg/metric ton zinc)
Emission rates
(kg/hr)
Size
>2 ym
<2 ym
Total
>2 ym
<2 ym
Total
Total
Total
Cd
0.432
0.648
1.08
3.43
5.14
8.57
3.16
17.1
Pb
0.310
0.435
0.745
2.45
3.45
5.90
2.18
6.43
Zn
0.846
0.794
1.64
6.71
6.29
13.0
4.80
38.8
Cu
0.002
0.001
0.003
0.017
0.011
0.027
0.010
0.118
Total
parti culates
4.97
3.28
8.25
39.4
25.9
65.3
24.1
191
ro
^j
o
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TABLE 4-13. PARTICULATE EMISSIONS FROM A ZINC COKING FURNACE (10)
to
Emission factors
(kg/metric ton zinc)
Emission rates
(kg/hr)
Size
>2 ym
<2 ym
Total
>2 ym
<2 ym
Total
Cd
0.421
0.749
1.17
0.670
1.20
1.87
Pb
0.250
0.375
0.625
0.400
0.600
1.00
Zn
10.3
2.48
12.8
16.5
3.98
20.5
Cu
0.005
0.004
0.010
0.009
0.007
0.016
Total
parti culates
12.2
11.7
23.9
19.6
18.7
38.3
-------
Ideal control for S02 emissions from sintering would be to eliminate as
much sulfur as technically possible during roasting. Based on the capability
of fluid bed and suspension roasters, a calcine averaging 1.5 percent total
sulfur could be produced, rather than the current typical calcine with
approximately 3 percent total sulfur. Roasting 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 accomplish sintering, elimination of sulfur in the
calcine could increase costs.
Various combinations of settling flues, cyclones, ESP's, and baghouses
are used on sintering machines. Efficiencies range from 94 to 99 percent.
Table 4-12 includes comparative data for different control devices.
Sinter crushing and screening operations have enormous potential 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. 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 Cate-
gory. EPA-440/1-75/032. Effluent Guidelines Division Office of
Water and Hazardous Materials. U.S. Environmental Protection
Agency. Washington, D.C. November 1974.
2. Field Surveillance and Enforcement Guide for Primary Metallurgical
Industries. EPA-450/3-73-002. Office of Air and Water Programs,
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. December 1973.
3. Lund, R.E., et al. Josephtown Electrothermic Zinc Smelter of St.
Joe Minerals Corp. AIME Symopsium on Lead and Zinc, Vol. II.
1970.
4. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer. Ency-
clopedia of Chemical Technology. Interscience Division of John
Wiley and Sons, Inc. New York. 1967.
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.
6. 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 Cate-
gory. Battelle Columbus Laboratories. EPA Contract No.
68-01-1518. Draft data.
272
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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. Environ-
mental Protection Agency. Research Triangle Park, North Carolina.
July 1974.
9. Jones, H.R. Pollution Control in The Nonferrous Metals Industry.
Noyes Data Corporation. Park Ridge, New Jersey. 1972.
10. Jacko, Robert B., and David W. Nevendorf. Trace Metal Emission
Test Results from a Number of Industrial and Muncipial Point
Sources. Journal of the Air Pollution Control Association.
27(10):989-994. October 1977.
273
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PRIMARY ZINC PRODUCTION PROCESS NO. 8
Vertical Retorting
1. Function - The vertical retort process is a continuous reduction/vola-
tilization method for producing high-purity zinc from zinc oxide by reduction
with carbon at elevated temperatures in vertical silicon-carbide retorts. An
alternate retorting process is the electric retort (Process No. 9). The
older horizontal retorting process is no longer in use in the United States.
Because of the relatively low boiling point of zinc (906°C), reduction
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 pyrometallurgi-
cally, there is virtually no alternative. Even at 857°C (the lowest tempera-
ture 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 commercially
is carbon monoxide. The zinc-reduction cycle consists of the following
reactions:
ZnO •*• CO -»• Zn(vapor) + C02 (actual reduction step)
C02 + C + 2 CO (reneneration 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 reduction 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 (1). The following additional reactions occur during
retorting:
Fe203 + 3C -»• 2Fe + 3CO
Fe203 + C •*• 2FeO + CO
ZnS + Fe -»• Zn + FeS
Zn + C02 -> ZnO + CO (blue powder formation)
2
Zn2Si04 + 2C •»• 2Zn + 2CO + SiO
FeO + Si02 •*• 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 smelting column or,
if zinc oxide is the product, by the danger of contaminating the product with
soot. Carbon in excess of stoichiometric reduction requirements is normally
274
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provided to furnish extra reaction surface; this compensates for the slowness
of reduction of carbon dioxide by carbon.
The charge to a vertical retort must be in the form of a hard sinter or
briquette. The reduction fuel is mixed with 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 zinc (1).
Although briquetting is an expensive operation, the ability to use
briquetted sinter feed could be turned into an advantage since it permits the
use of a soft sinter produced either by coke-sintering or roast-sintering.
Such a process is being used in England by the Imperial Smelting Corporation,
Limited, but, with one exception, 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 kilo-
grams 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 (1).
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 sidewalls. Gases from the combustion chambers
are used to preheat incoming air for combustion by means of recuperators.
The briquettes maintain their shape throughout 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
condensed to liquid zinc (2). The residue, containing approximately 10
percent 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 products 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 nitro-
gen, and some carbon dioxide (4). These gases exit near the top of the
charge column, pass through a zinc condenser, and then to a venturi scrubber.
275
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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.
When the zinc vapor is not cooled quickly enough from reduction tempera-
tures to condensation temperatures, some of the carbon monoxide present in
the mixture decomposes into carbon dioxide. This carbon dioxide, and the
small amount already present in the vapor mixture, oxides a portion of the
vaporized zinc at temperatures below 1100°C. The resulting zinc oxide
adheres to fine globules of condensed zinc forming an undesirable by-product
called "blue powder" (1). The blue powder 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
supplementary fuel or flared.
About one-half of the zinc produced in vertical retorts is refined to
99.99 percent purity by means of continuous fractional distillation. 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 cadmium dust
depending on how the process is operated.
2. Input Materials - Feed consists 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, 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 typi-
cally 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 operate
from 1300 to 1400°C (1). Internal pressures are slightly below ambient.
Exact operating temperatures for fractional distillation or other purifica-
tion steps vary, since separation is based on differences in boiling points.
276
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4. Utilities - Heating is done in vertical retort furnaces by combustion of
gas surrounding the retort side walls. One source eliminates that 4,000 to
5,000 kilocalories per kilogram of zinc are required (6). Another source
reports energy efficiency to be about 10 percent, with typical energy consump-
tion 5.0 million kilocalories per metric ton of zinc produced. One company
using vertical retorts finds that 9.1 million kilocalories of coal and coke
per metric ton of zinc produced is required for the briquetting process.
However, this company claims an energy consumption by retorts of only 2.8
million kilocalories per metric ton of distilled zinc, for a total energy
consumption of 12.8 million kilocalories per metric ton (7). Energy require-
ments for refining are about 834,000 kilocalories per metric ton of zinc pro-
duced (8).
5. Waste Streams - Emissions are minor when compared with other steps in
the smelting process such as roasting and sintering. SO, emissions average
less than 50 ppm (9). Flow rate for the carrier gas is 23,000 cubic meters
per metric ton of product, with 2.5 to 3.0 percent carbon dioxide (3).
Particulate emissions are evident only during charging for approximately
one minute. High-efficiency metal recovery is possible from these metal and
metal oxide fumes. Information on particulate emissions, including size
distribution data is presented in Table 4-14. Blue powder 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.
The gas washing water contains zinc and metal oxides, possibly hydro-
carbons, various particulates (as suspended solids), and the corresponding
products of hydrolysis.
Residues are also generated in vertical retorting operations. Amounts
produced are around 1050 kilograms 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.
The germanium and gallium contents originally 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 gallkim, using residues
from zinc and aluminum production. One refinery in Oklahoma produced all the
primary domestic germanium from zinc smelter residues in 1976 (11).
An analysis of some constituents in the residue from one vertical
retort furnace is given in Table 4-15.
277
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ro
-vl
oo
TABLE 4-14. UNCONTROLLED PARTICULATE EMISSIONS FROM A
ZINC VERTICAL RETORTING FURNACE (10)
Emission factors
(kg/metric ton zinc)
Emission rates
(kg/hr)
Size
>2 ym
<2 ym
Total
>2 ym
<2 ym
Total
Cd
0.006
0.027
0.033
0.002
0.012
0.014
Pb
0.038
0.131
0.169
0.016
0.057
0.073
Zn
0.920
2.28
3.20
0.393
0.967
1.36
Cu
0.0006
0.0004
0.001
0.0002
0.0002
0.0004
Total
parti culates
2.11
5.04
7.15
0.890
2.14
3.03
-------
TABLE 4-15. ANALYSIS OF VERTICAL RETORT FURNACE RESIDUES (3)
Constituent
Cadmi urn
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 (12). The residue also con-
tains 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 composition
depends 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 condensation 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 in open slag dumps or processed to
recover their metal values. The best control technology for the slag dump is
sealing the soil and routing runoff to a waste treatment lagoon.
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.
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. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry - Volume II, Primary arid
Secondary Nonferrous Smelting and Refining. Draft. April 1975.,
279
-------
4. Field Surveillance and Enforcement Guide for Primary Metallurgical
Industries. EPA-450/3-73-002. Office of Air and Water Programs,
U.S. Environmental Protection Agency. Research Triangle Park,
North Carolina. December 1973.
5. 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 Cate-
gory. EPA-440/1-75/032. Effluent Guidelines Division Office of
Water and Hazardous Materials. U.S. Environmental Protection
Agency. Washington, D.C. November 1974.
(5. 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.
7. Battelle Columbus Laboratories. 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-68-01-1518. Draft data.
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. Environ-
mental Protection Agency. Research Triangle Park, North Carolina.
July 1974.
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 Protection
Agency. Research Triangle Park, North Carolina. October 1974.
10. Jacko, Robert B., and David W. Neuendorf. Trace Metal Particulate
Emission Test Results from a Number of Industrial and Municipal
Point Sources. Journal of the Air Pollution Control Association.
27(10):989-994. October 1977.
11. Commodity Data Summaries 1977. U.S. Department of the Interior,
Bureau of Mines. Washington, D.C. 1977.
12. 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. Environmental Protection
Agency. Washington, D.C. September 1973.
280
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PRIMARY ZINC PRODUCTION PROCESS NO. 9
Electric Retorting
1. Function - Electric retorting is a continuous reduction/volatilization
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. Vertical retorting (Process No. 8) is the
other reduction method in use at zinc smelters. This newest zinc-smelting
furnace was designed to overcome the difficulties involved in heating the
charge externally. As in the other retorting process, 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 reduc-
tion 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 St. Joe
Minerals Corporation, which began commercial operation in 1930 (1).
The St. Joe electrothermic furnaces are basically vertical, refractory-
lined cylinders. The largest furnaces now in use have an inside diameter of
1.5 meters and are 15 meters high, with a production capacity 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 conductor, 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 produc-
tion of over 90 metric tons per day from a single unit. If necessary,
further 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 pro-
cesses. First, the increased thermal efficiency (compared with external
heating methods) results in cost savings in fuel consumption. Larger quanti-
ties of charge can be treated, and the continuous operation is amenable to
281
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automation. The furnace can readily process secondary zinc scrap and zinc
residues. Because of special deleading by heat treatment in multiple-hearth
roasters followed by desulfurization in fluidized-bed roasters, electrothermic
furnaces emit practically no S02 or particulates.
2,. Input Materials - The feed 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, thereby 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 mercury 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 condenser cooling well
maintain a constant batch temperature of 480° to 500°C (2). 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-centimeter diameter
graphite electrodes may range as hiqh as 800 amperes. A furnace contains
eight individual single-phase circuits, each typically carrying 770 to 1190
kilowatts. 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 2800 kilowatt-hours per metric ton
of metal. Power input per electrode circuit ranging up to 1255 kilowatts,
corresponding to a maximum of 10,000 kilowatts per furnace has been'reported
(5). Energy consumption for subsequent refining steps is the same as that
used fn other retorting processes.
Another source reports 25 to 30 percent energy efficiency for electro-
thermic retorting, with the process consuming about 2.8 million kilocalories
per metric ton of zinc. With an additional 5.8 million kilocalories per
metric ton of zinc consumed as coke, a total of 8.6 million kilocalories per
metric ton of zinc is used. If the energy for electric generation is also
considered (assuming 33 percent efficiency), total energy consumption in-
creases to 14.2 million kilocalories per metric ton (6).
5. Waste Streams - As with other types of retorting, emissions are minor
relative to those from roasting or sintering.
282
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Particulate emissions consist primarily of metal and metal oxide fumes.
Participate emission levels for electrothermic retortinq average about 10
kilograms per metric ton of zinc produced. As in the other retorting pro^.
cesses, blue powder is the principal constituent of these emissions, along
with cadmium, copper, chromium, lead, and iron. Also as with the other pro-
cesses, 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 averages
850°C. Gas composition is approximately 45 percent zinc vapor and 45 percent
carbon monoxide; the balance is nitrogen, carbon dioxide, and hydrogen (5).
The gas washing water contains the same impurities as in 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 vertical 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 kilocalories per cubic meter,
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 recycled. Besides permitting recovery
of residue containing enough zinc and carbon to make retreatment worthwhile,
a minimum of power is consumed in unproductive melting of residue. 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 Protection
Agency. Research Triangle Park, North Carolina. October 1974.
283
-------
2. 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 Cate-
gory. EPA-440/1-75/032. Effluent Guidelines Division Office of
Water and Hazardous Materials. U.S. Environmental Protection
Agency. Washington, D.C. November 1974.
3. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry - Volume II, Primary and
Secondary Nonferrous Smelting and Refining. Draft. April 1975.
4. Schlechter, A.M., and A. Paul Thompson. Zinc and Zinc Alloys. In:
Kirk-Othmer. Encyclopedia of Chemical Technology. Volume 22.
Interscience Division of John Wiley and Sons, Inc. New York 1967.
5. Lund, R.E., et al. Josephtown Electro-thermic Zinc Smelter of St.
Joe Minerals Corporation. AIME Symposium on Lead 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. Environ-
mental Protection Agency. Research Triangle Park, North Carolina.
July 1974.
284
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PRIMARY ZINC PRODUCTION PROCESS NO. 10
Oxidizing Furnace
1. Function - In the direct, or American, process for zinc oxide production,
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 discussed 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 subsequently 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 grate-type furnaces, the coal-sinter feed (usually as briquettes) 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 atmo-
sphere 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 rpm (1). The additional heat for reduction of the
zinc is supplied by the flow of gaseous fuel through the kiln. The vaporized
zinc then is ducted to a combustion 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 evolved zinc vapor and carbon monoxide exit the charge. Preheated coke
and zinc-bearing sinter are continously fed to the furnace. Coke serves as
the principal electrical conductor. Electricity introduced through the
electrodes develops the heat energy required for smelting. Further details
on the electrothermic furnace are given in Process No. 9.
Analysis of an American process zinc oxide from one plant is presented
in Table 4-16.
285
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TABLE 4-16. ZINC OXIDE IMPURITIES AND BRIGHTNESS (2)
Lead as PbO
Cadmium as CdO
Iron as F0203
Manganese as MnO
+325 mesh screen residue
Brightness; Hunter D-40
0.009%
0.010%
0.015%
0.002%
<0.03%
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 (!)• 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 desired (1).
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 approximately 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 (1); 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 produc-
tion, with temperatures of 1200°C at the vapor ring elevation, 1400°C in the
main smelting zone, 900'°C near the walls, and 1300°C at the bottom electrode
elevation (1). Further details on operating conditions 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 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.
286
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The main factor influencing production in electrothermic furnaces is the
quantity of electricity introduced. Josephtown's largest furnace operates at
10,000 kilowatts. Energy consumption averages 2800 kilowatt-hours per metric
ton of metal (2).
5. Waste Streams - Combustion products and some zinc oxide are emitted from
all three furnaces; again, quantities are unspecified.
There are no liquid wastes from the process.
A solid residue is produced, estimated at 350 kilograms per metric ton of zinc
oxide. This residue contains approximately 6.2 percent zinc and 0.09 percent
of other metals 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 presented in Table
4-17.
TABLE 4-17. SELECTED CONSTITUENTS OF OXIDIZING FURNACE RESIDUE (3)
Constituent
Cadmium
Chromium
Copper
Lead
Zinc
Concentration,
ppm
10
17
810
68
62,000
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 No's. 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.
2. Lund, R.E., et al. Josephtown Electrothermic Zinc Smelter of St.
Joe Minerals Corp. AIME Symposium on Lead and Zinc. Volume II.
1970.
3. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry - Volume II, Primary and
Secondary Nonferrous Smelting and Refining. Draft. April 1975.
287
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PRIMARY ZINC PRODUCTION PROCESS NO. 11
Leaching
1. Function - Electrolytic production of zinc is an alternative to pyro-
metallurgical processing. The first step is to separate zinc from gangue
minerals by leaching roasted calcine in recycled electrolyte solution. The
zinc dissolves, and the insoluble gangue is separated from the solution by
decantation, thickening, and filtration. The solution is purified (Process No.
12) and the waste solids are either discarded or, if their concentration
warrants, further processed to recover lead and precious metals (1,2).
Two general leaching methods can be employed, described as either a single
or a double leach. In a single leach, recycled electrolyte, which is a solu-
tion containing principally sulfuric acid, is brought only once into contact
with the calcine. Zinc oxide in the calcine reacts with sulfuric acid to form
soluble zinc sulfate and water. The single leach is not often practiced, how-
ever, since losses of sulfuric acid are excessive and recovery of zinc is poor.
Double leaching is used most often. In several variations, the calcine
is leached first in a solution that is neutral or slightly alkaline, then in
an acidic solution, with the liquid passing countercurrently to the flow of
calcine. In the neutral leach, the readily soluble sulfates from the calcine
dissolve, but only a portion of the zinc oxide enters into solution. The
second acidic leach solubilizes the remainder of the zinc oxide, but also
dissolves many impurities, especially iron. Recycle of the liquor to the first
neutral stage causes much of the iron to reprecipitate, so the neutral leach
acts also as an initial stage of solution purification. In some of the more
complex process variations, considerable overlap occurs between leaching and
purification steps, and the calcine may be subjected to as many as four
leaching operations in progressively stronger or hotter acids to bring as much
of the zinc as possible into solution.
The leaching process is conducted most often in a series of agitated
tanks. Batch operation is common, since it is thereby possible to compensate
for variations in calcine composition. A few smelters, especially those pro-
cessing ore of consistent quality, are equipped with continuous leaching
equipment. In all of the leaching operations, the pH, temperature, and
solution composition at each step are carefully regulated.
2. Input Materials - The principal input to the leaching process is calcined
zinc concentrate from roasting (Process No's. 4, 5, and 6). If a pelletized
feed is used in the roaster, the calcine must be ground prior to leaching. Air
classification and grinding is occasionally practiced also with unpelletized
material. One foreign plant grinds calcine to 95 percent through a 200-micron
screen (about 70 mesh) (3). Wet grinding with a portion of the leach liquor
is also practiced in some smelters.
The only other input specific to the leaching process is recycled spent
electrolyte from Process No. 13, containing about 200 grams of sulfuric acid
per liter (2).
288
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3. Operating Conditions - Batch leaching processes usually operate at atmos-
pheric pressure, whereas continuous processes may include pressurized steps
up to 2.5 kilograms per square centimeter (1). Most leach operations take
place around 50°C; exothermic chemical reactions maintain this temperature
without requiring additional heat energy. Some hot acid process variations
may employ temperatures up to 90°C(3).
4. Utilities - Electricity is required for pumping the solution and convey-
ing the calcine and final residue. No estimates of energy consumption for
the leaching process are given. In batch leaching, air at about 6.3 kilo-
grams per square centimeter must be available to clear out accumulating de-
posits of coarse material at the bottom of the tanks. In continuous leach--
ing, it is seldom necessary to use air at pressures higher than the normal
operating pressure of 1.4 to 2.5 kilograms per square centimeter (1). Con-
tinuous leaching requires a larger total volume of air per tank, because
agitation must be continuous. During agitation in either process, con-
sumption ranges from 2.8 to 4.2 cubic meters of air per minute (1).
5. Waste Streams - As in the closely associated purification process, emissions
of acid mist occur from the leach tanks. This waste is described more com-
pletely in Process No. 12.
Leach solution may be cooled in open towers, as described for electrolyte
cooling in Process No. 13. Atmospheric mist may therefore be created from
these towers, as described in connection with the electrolysis process.
Except for leaks and spills, no liquid waste is produced from this pro-
cess .
After all leaching, the solid residue is filtered from solution and the
filter cake is rinsed with fresh water. This cake will contain all the lead
originally present in the concentrate, and also other acid-insoluble trace
elements such as indium, gold, and the platinum-group metals. Other minerals
present will be silica, alumina, and silicates of iron, aluminum, and calcium.
The quantity and composition will vary with the characteristics of the ore
concentrate; one measurement reports a quantity of 360 kilograms per metric
ton of zinc produced (1).
6. Control Technology - Control of acid mist emissions is described in con-
nection with the similar mists produced in Process Nos. 12 and 13.
The solid residue will frequently be sent to a lead smelter for recovery
of the lead and other reclaimable elements. Alternatively, the residue may
be batch-treated at the zinc smelter with cyanide for recovery of gold. In
other cases, the concentrations of recoverable metals may be so small that
the residue will be discarded in a dump. Ore residue from a double leach
process should be inert; however, as generally practiced, the residue will
also contain trace elements from Process No. 12. Single leach residues may
also contain zinc oxide or zinc ferrites. The potential of this waste for
secondary water pollution is unreported.
7. EPA Source Classification Code - None.
289
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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.
2. 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.
EPA-440/1-75-032. Effluent Guidelines Division, Office of Water and
Hazardous Materials. U.S. Environmental Protection Agency. Wash-
ington, D.C. November 1974.
3. E. Van Den Neste. Metallurgie Hoboken-Overpelt's Zinc Electrowinning
Plant. CIM Bulletin. August 1977.
290
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PRIMARY ZINC PRODUCTION ' PROCESS NO. 12
Purifying
1. Function - The leaching of zinc calcine causes other elements in addition
to zinc to dissolve. Unless impurities are removed from the solution, they
will either contaminate the zinc product or interfere with the proper operation
of the electrolysis process. The solution from leaching is therefore purified
to remove metallic ions that are more electropositive than zinc.
Purification is usually conducted in large agitated tanks. A variety of
reagents is added in a sequence of steps that causes impurities to precipitate.
The precipitates are separated from the solution by filtration. The purifi-
cation techniques are among the most advanced applications of inorganic
solution chemistry in industrial use, and vary from one smelter to another.
Iron is often removed in conjunction with the leaching process (Process No.
11) by precipitation as a hydrated oxide (goethite) or a complex sulfate
(jarosite). Some of these processes, which are patented, also reduce the
concentration of arsenic and other elements. Almost all smelters then add
zinc dust, often in the form of "blue powder" from the pyrometallurgical pro-
duction of zinc. This addition causes a cementation reaction that precipitates
cadmium, copper, and several other elements. The final steps remove all but
trace quantities of a group of metals that includes arsenic, antimony, cobalt,
germanium, nickel, and thallium. These metals severely interfere with
electrolytic deposition of zinc, and their final dissolved concentrations are
limited usually to less than 0.05 milligrams per liter.
2. Input Materials - The principal input is the filtered, acidic, mineral-
rich solution from Process No. 11. Reagents are mostly inorganic, primarily
finely divided metals such as zinc, arsenic, and antimony. Fresh sulfuric
acid may be added in small quantities, and lime may be used to remove excess
sodium carbonate or sodium hydroxide may be used for iron precipitation.
Organic materials such as l-nitroso-2-naphthol may be used to remove cobalt (1).
Inorganic salts such as copper sulfate may be added to catalyze or promote
some precipitation reactions. Except for zinc dust, the quantities of these
additives are small.
3. Operating Conditions - The purification process takes place at tempera-
tures ranging from 40° to 85°C, and pressures ranging from atmospheric to 2.5
kilograms per square centimeter (2). The conditions at each step of pre-
cipitation are carefully regulated.
4. Utilities - Electricity is required to pump solutions and drive equipment
agitators. Steam and noncontact cooling water may be used to heat and cool
solutions. Quantities are not reported.
5. Waste Streams - The atmospheric emission from this process consists of a
mist that develops from the ventilation of the leach and purification tanks.
Ventilation is necessary since side reactions can cause evolution of small
quantities of explosive hydrogen gas which must not be allowed to accumulate.
The mist contains sulfuric acid and smaller quantities of zinc, calcium, and
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arsenic (3). There is no estimate of the quantity of waste products released
from this source.
Except for leaks and spills, no liquid waste is created by this process.
Precipitated solids consisting of impurity elements and excess reagent
metals are accumulated as cakes from pressure filters. The composition of the
cakes is variable, depending on the characteristics of the zinc concentrates
and the details of the processing. All are highly metalliferous.
In some process modifications, much of the iron, and part of the arsenic
and antimony originally present in the concentrate, are precipitated into and
discarded with the insoluble residue from Process No. 11.
6. Control Technology - Ventilation of leach and purification tanks is
usually controlled with impingement or centrifugal demisting equipment. The
efficiency of these devices in this service has not been reported.
The disposition of most of the solid residues has also not been reported.
With some zinc concentrates, a filter cake rich in copper is produced which is
sold to copper smelters. Some residues are recycled to the roaster or leach
tanks, or are separately treated to reclaim zinc and cadmium. Two companies
are reported to have refined indium from some of the residues (4). Filter
cakes rich in cobalt are apparently being stockpiled by a foreign refinery,
and similar disposition may be practiced by some U.S. refineries (2). It is
not known whether any of these materials is being discarded.
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.
2. E. Van Den Neste. Metallurie-Hoboken-Overpolt's Zinc Electrowinning
Plant. CIM Bulletin. August 1977.
3. Privileged communication, EPA files.
4. Commodity Data Summaries 1976. U.S. Department of Interior, Bureau
of Mines. Washington, D.C. 1977.
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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. Electrolysis takes place in rectangular tanks, or
cells, each of which holds a number of closely spaced rectangular metal plates.
Alternate plates are made of lead containing 0.75 to 1.0 percent silver; these
are the anodes that are electrically connected to a positive potential. The
remaining plates are made of aluminum, and -are connected with a negative
electrical potential.. Purified electrolyte from Process No. 12 is circulated
slowly through the cells, and water in the electrolyte dissociates, releasing
oxygen gas at the anode. Electrode voltage is maintained sufficiently high so
hydrogen is not released at the cathode; instead, zinc ions absorb electrons
and deposit zinc metal. Hydrogen ions remain in solution, and thereby regen-
erate sulfuric acid for recycle to the leach process (1,2).
Zinc smelters contain a large number of electrolytic cells, often several
hundred. They are most often made of concrete with a lead, plastic, or
vitreous lining, and are electrically connected in series banks. A portion of
the electrical energy is converted into heat, which increases the temperature
of the electrolyte. Therefore, a portion of the electrolyte is continuously
circulated through cooling towers. These are usually open towers, in which
the electrolyte falls through a rising stream of air drawn through the tower
by fans. This method both cools the electrolyte and evaporates the excess
water (3). The cooled and concentrated electrolyte is repumped to the cells.
Every 24 to 48 hours, each cell is shut down and the zinc-coated cathodes
are removed, rinsed, and the zinc is mechanically stripped from the aluminum
plates. Stripping is accomplished manually in some smelters, while others are
specialized automated equipment. The aluminum cathodes are then chemically
cleaned, replaced in the cells, and the cell is restored to normal operation.
Stripped zinc is sent to Process No. 14 for melting and casting.
2. Input Materials - The principal input is purified electrolyte, which
is a water solution containing about 70 grams of zinc per liter and about 200
grams of sulfuric acid per liter (4). Barium hydroxide or manganese sulfate
may be added to the electrolyte in order to form insoluble coatings on the
cell anodes, thereby minimizing both anode corrosion and lead contamination of
product zinc. Colloidal materials are also usually added to prevent uneven de-
position of zinc on the cathodes; materials used include glue, goulac, gum
?!
arabic, and a mixture of agar-agar, sodium silicate, and cresylic acid (2).
3. Operating Conditions - Electrolytic cells operate at 30° to 35°C and
atmospneric pressure (4).
4. Utilities - Electrolysis consumes the major amount of energy in an
electrolytic zinc smelter. Most plants use direct current at a current
density of about 600 amperes per square meter of cathode surface. Voltage
drop is 3.3 to 3.5 volts per cell, and current efficiency is 85 to 94 per-
cent. Approximately 3300 kilowatt-hours of power is needed to electrolyze a
metric ton of cathode zinc (4).
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Electricity in much smaller quantity is used to operate pumps and fans.
5. Waste Streams - There are two sources of atmospheric emissions from this
process. Escape of oxygen at the cell anodes causes the formation of a mist
of approximately the same composition as the electrolyte that escapes into the
air of the cell room. Another mist of the same approximate composition is re-
leased from the atmospheric cooling towers. Cell emission has been estimated
to be 3.3 kilograms per ton of zinc produced (1). No estimate of cooling tower
emissions has been reported.
If the ore concentrate contains quantities of sodium or halogen compounds,
a portion of spent electrolyte must be routinely removed and discarded. This
is not known to be occurring regularly in U.S. smelters. In general, there is
no loss of electrolyte or other liquid waste other than leaks and spills.
A sludge accumulates in the cells which is periodically removed.
4-18 provides reported analyses of this material.
TABLE 4-18. ANALYSES OF ANODE SLUDGES FROM
ELECTROLYTIC ZINC REFINING (5).
Table
Concentration, ppm
Fresh anode sludge
Old anode sludge (from dump)
Cd
12
1,400
Cr
10
8
Cu
85
1,900
Pb
170,000
89,000
Zn
12,800
39,200
6. Control Technology - Cell rooms must be well ventilated to avoid accumu-
lation of oxygen. This ventilation also serves to remove mist from the room.
There are no reports of treatment being applied to this air stream, although
this is apparently the largest source of air pollution from the electrolytic
process (6).
Treatment of mist from the electrolyte cooling towers is also not docu-
mented. An opinion has been expressed that the concentration of these
emissions is so diluted by the large volumes of air that it is not a nuisance
to nearby residents (3).
Disposition of anode sludge is also unreported. The material is a rich
source of recoverable minerals, and is probably reprocessed or recycled, as
are the filter cakes from Process No. 12.
7. EPA Source Classification Code - 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
294
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Air and Waste Management, U.S. Environmental Protection Agency. Re-
search Triangle Park, North Carolina. October 1974.
2. 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.
3. E. Van DenNeste. Metallurie-Hoboken-Overpelt's Zinc Electrowinning
Plant. CIM Bulletin. August 1977.
4. Krupkowa, D., and A. Udrycki. New Electrolytic Zinc Plants. Chemical
Age of India. August 1975.
5. Calspan Corporation. Assessment of Industrial Waste Practices in the
Metal Smelting and Refining Industry - Volume II, Primary and Secon-
dary Nonferrous Smelting and Refining. Draft. April 1975.
6. Privileged communication, EPA files.
295
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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. The molten zinc is cast into 27-kilogram
slabs on an in-line casting machine. Some zinc is cast into 640- to 1100-
kilogram blocks in stationary molds (1).
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 purification 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 can be
melted in a reverberatory furnace with the formation of 1 percent or less of
dross; electrolytic cathodes lose 6 to 17 percent under these circumstances,
depending on the presence of glue, cobalt, and the like in the electrolyte
(1).
2. Input Materials - Inputs are zinc, ammonium chloride flux, and various
alloying materials to meet special requirements. 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.
Heat requirements 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 the form of metal oxides, mold washes, and lubricants from casting
equipment.
The only solid waste is the dross produced if the slab zinc is melted in
an electrolytic furnace.
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.
296
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The dross is treated 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.
297
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PRIMARY ZINC PRODUCTION PROCESS NO. 15
Cadmium Leaching
1. Function - Leaching selectively dissolves as much cadmium as possible
from various cadmium-bearing dusts, fumes, and sludges from primary zinc
plants, precipitating lead and other impurities without precipitating any of
the dissolved cadmium. There is no separate primary cadmium industry in the
United States. Cadmium recovery processes use by-products of other opera-
tions, all involving zinc, in four major categories (1):
(1) Fumes and dusts from roasting and sintering of zinc concentrates.
(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. The second category involves
recycled zinc metal, and 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 chlorate to ensure
complete dissolution of cadmium sulfide. Cadmium goes into solution by the
following reaction:
CdO + H2S04 -»• CdS04 + H20
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 significant quantities of gold, silver,
and indium (2). More sulfuric acid is then added to the solution in order to
bring the concentration up to 10 percent and to raise the temperature.
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).
298
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Another element in the flue dusts from roasting and sintering that is
economically recoverable is thallium; this processing is not done at primary
zinc smelters. One refiner in Colorado is the sole domestic producer of
thallium (4).
Another source of high-cadmium residues for leaching is electrolytic
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 in stages making possible rough separations such as
high-copper and high-cadmium precipitates. Copper sulfate and arsenic tri-
oxide may be added during these stages. The cadmium precipitate is high in
zinc and contains virtually all the cadmium originally present in the elec-
trolyte. 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 sul-
furic 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 grams per
liter sulfuric acid and 65 grams per liter 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 reported 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 acid (1).
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 grams per 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 exhaust gas temperature of 60°C from leaching
was reported at one electrolytic plant, which also reported exhaust gases
from pre-leaching purification stages ranging from 60 to 90°C (3).
299
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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. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry - Volume II, Primary and
Secondary Nonferrous Smelting and Refining. Draft. April 1975.
2. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer. Ency-
clopedia of Chemical Technology. Interscience Division of John
Wiley and Sons, Inc. New York. 1967.
3. Battelle Columbus Laboratories. 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-68-01-1518. Draft data.
4. Commodity Data Summaries 1977. U.S. Department of Interior,
Bureau of Mines. Washington, D.C. 1977.
300
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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 precipitate
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 electrolyte. The
filtrate is then precipitated with zinc dust in two or three stages to mini-
mize 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 purifica-
tion 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 (1).
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, potassium xanthate, and
potassium chromate or dichromate. Quantities for these inputs are not speci-
fied.
3. Operating Conditions - Precipitation takes place at atmospheric pressure.
Temperatures range from ambient to 40°C (1).
4. Utilities - Electricity is used to pump liquids, and natural gas or oil
is used to heat the stripped liquor. Quantities are not cited.
5. Waste Streams - Water vapor is released from both purification and
precipitation operations at temperatures ranging from 45° to 60°C.
301
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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 (2). The purification precipitates may contain arsenic and mercury,
both of which are relatively low in concentration in ores and concentrates.
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 (3).
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 settling ponds, where the solids can 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. Ency-
clopedia of Chemical Technology. Interscience Division of John
Wiley and Sons, Inc. New York. 1967.
2. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry - Volume II. Primary and
Secondary Nonferrous Smelting and Refining. Draft. April 1975.
3. 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 Cate-
gory. EPA-440/1-75/032. Effluent Guidelines Division Office of
Water and Hazardous Materials. U.S. Environmental Protection
Agency. Washington, D.C. November 1974.
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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 collect-
ing 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 conventional hori-
zontal-type retort, where the cadmium is reduced and collected 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 redistillation. Typical impurities in the pro-
duct 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 antimony (1). 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 recovery 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. Contin-
uing 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 precipitated
with hydrogen sulfide, then dissolved in sulfuric acid. This sulfate solu-
tion can be electrolyzed for recovery of pure sponge thallium. The sponge is;
washed, pressed into blocks, melted, and cast.
Processing of concentrated lead smelter baghouse dust in an electrolytic
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 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
303
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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 elec-
trolyte, 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 (1).
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 pressurized.
Operating temperatures of 455°C have been reported for the gas melting and
ammonium chloride stages, and 790° to 910°C in the furnace itself. Electroly-
sis also takes place at atmospheric pressure, with the temperature held at
about 30°C (1).
4. Utilities - Electricity, oil, or natural gas is used for melting the
cadmium product. Air is brought into the furnace. Specific quantities were
not found. In electrolysis, current density may range from 140 to 360
amperes per square meter (2).
5. Waste Streams - Particulates are released at several stages in the
process"! One plant reported the release of 20 kilograms per hour of ammonium
chloride from fluxing and 108 kilograms per hour of cadmium from the retorting
furnace (1).
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 copper (2). Thallium is not
regarded as a serious impurity since it is removed 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 - 21.5 percent, zinc - 3.9
ipercent, and thallium - 40 ppm. The filter cake amounts to 1.8 kilogram
per metric ton of zinc product (3).
6. Control Technology - Particulates can be controlled with fabric filter
systems^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 pro-
304
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cessing, the spent electrolyte is returned to various leaching stages in the
circuit. In some cases the electrolytic cadmium process cycle is completely
closed with no discharge.
7. EPA Source Classification Code - None
8. References -
1. Battelle Columbus Laboratories. 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 Contract No. 68-01-1518.
Draft data.
2. Howe, H.E. Cadmium and Cadmium Alloys. In: Kirk-Othmer. Ency-
clopedia of Chemical Technology. Interscience Division of John
Wiley and Sons, Inc. New York. 1967.
3. Calspan Corporation. Assessment of Industrial Waste Practices in
the Metal Smelting and Refining Industry - Volume II, Primary and
Secondary Nonferrous Smelting and Refining. Draft. April 1975.
305
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SECTION 5
AIR MANAGEMENT
The most noticeable form of pollution from the primary copper, lead, and
zinc industries is the discharge of materials to the atmosphere. Although
mining and concentrating operations produce considerable fugitive dusts,
there are no direct atmospheric discharges. Electrolytic copper refineries
cause very little atmospheric pollution. It is the smelters themselves that
produce trace metal-bearing particulates, metallic fumes, and S02 gas, all of
which may be discharged into the air or lost to the surroundings as fugitive
emissions. Many of the metal compounds that are volatilized by high-tempera-
ture operations escape either as fugitive emissions or by passing through
dust collection systems as fume which condenses upon leaving the stack.
Emissions of arsenic alone can amount to hundreds of pounds per hour where
high-temperature dust collectors are used. Table 5-1 summarizes sources,
quantities, and characteristics of atmospheric emissions from mining, con-
centrating, and smelting operations of the primary copper, lead, and zinc
industries.
EMISSION CHARACTERISTICS
Almost every smelting process produces atmospheric emissions, and
despite the variety of processes, many of the emission streams have common
characteristics. In production of these metals, an ore concentrate 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 metallic elements in the con-
centrate, including antimony, arsenic, boron, cadmium, lead, mercury, sele-
nium, tellurium, vanadium, and zinc. The composition of these emissions
varies widely and depends mainly on the specific ore and the type of smelting
operation. The metallic constituents in the emission stream are characterized
by their small particle sizes; usually less than one micron. Little quantita-
tive data on particulate composition 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 5-2 for
samples collected in a flue serving a copper ore roaster and arsenic plant
after large particulate has settled out. These data are not necessarily
representative of other plants and processes since this ore had a high arsenic
content, but the trend to higher metallic content in the fine fraction is
evident.
306
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TABLE 5-1. ATMOSPHERIC EMISSIONS FROM MINING, CONCENTRATING, AND SMELTING OPERATIONS
Process
COPPER
Mining
Concentrating
Multiple hearth roasting
Fluidization roasting
Drying
Reverberatory smelting
Electric smelting
Flash smelting
Peirce-Smith converting
Hoboken converter
Noranda
Slag treatment
Contact sulfuric acid
plant
DMA S02 absorption
Waste
Dust
Dust
Roasting gas
Roasting gas
Dust
Smelting gas
Off-gas
Off-gas
Off-gas
Off-gas
Off-gas
Off-gas
Off-gas
Exit gas
Characteristics
Characteristics of ore being mined
Characteristics of ore being mined
S02, participates
S02, participates
Same composition as ore
SO,, participates, Zn, Ca, Mg, etc.
fugitive dust
SO.,, some particulates
S02, particulates
Particulates, S02 (see Table 2-26)
S02. particulates
S02, particulates
Particulates, Zn and other element
fumes
S02, fine mist H2S04
S02, dilute H2S04 mist
Quantities
~110 kg/metric ton ore
~1 kg/metric ton ore
0.5 to 2% S02
12 to 16% S02
-
0.5 to 2.5% S02, between
14 and 40 kg participate/
metric ton copper. Largest
amount of dust produced
of processes
3 to 8% S02
10 to 20* SO,, 6 to n
(part.) of feed
4 to 10% S02
At least 8% S02
Low SO- content
-
0.01 to 0,5S of 34,000 to
68,000 m3/hr
-
Control
-
Cyclone separators, fugitive
dust uncontrolled
Cyclone collectors, preclpl-
tators, H20 spray
SO, for production of H-SO.,
w6t scrubbers, ESP ' *
Wet scrubbing
ESP, no control for SO,
H2S04 production
H2S04 production, cyclones,
ESP, etc.
H2SO. production
H2S04 production
ESP
-
None for SO? mist elimina-
tors or ESP for H2S04 mist
None usually required for
S02 ESP for acid mist
o
(continuod)
-------
TABLE 5-1 (continued).
Process
Elemental sulfur production
Arsenic recovery
Fire refining and anode
casting
Electrolytic purification
Slime acid leach
CuS04 precipitation
Slimes roasting
Dore" furnace
Scrubber
Heap and vat leaching
Sulfatlon roasting
Sponge Iron plant
(not yet in operation)
CLEAR reduction
CLEAR regeneration
CLEAR oxidation
Waste
Exit gas
Exit gas
Off-gas
Fugitive gas from
liberator cell
Gas evolution from Cu
and H2S04
Gas evolution from Cu
and H2S04
Off-gas
Off-gas
Exit gas
Fugitive dust
Off-gas
Off-gas
Off-gas
Off-gas
Off-gas
Characteristics
Possibly slight HzS, SO? possibly
fugitive dust from sulfur
Fugitive dust, small amounts of SOo
and arsenic fumes, particulates in
off-gas
Zn, Cd fumes, S02, particulates
AsH3 oxidizing to As03
so2
S02
Mineralized particulates and fumes,
Se02. S02
Particulates, Se, As, etc.
Particulates, furnaces not removed.
Selenium major constituent
-
Particulates, metallic fumes, SO?
Particulates, volatile metal fumes
expected
Some HC1 acid vapor, fugitive dust
HC1 acid vapor, entrained materials
HC1 acid vapor could be produced
(details not disclosed)
Quantities
No more than 0.7% S02 .
~
0.38 kg S02/metr1c ton Cu
treated, 5 to 20 kg partlcu-
1 ate/metric ton Cu produced
-
Small
Small
-
-
-
-
Considerably less than
conventional roaster
-
-
-
Control
Wet scrubbing
Fabric filter for fugitive
dust. ESP
None
Can be scrubbed
None
None
Scrubber
Wet scrubber
ESP
-
Acid plant
Unknown
Scrubbing with alkaline
solution
-
Scrubbing with alkaline
solution
CO
o
CO
(continued)
-------
TABLE 5-1 (continued).
Process
LEAD
Mining
Concentrating
Sintering
Acid plant
Blast furnace
Slag fuming furnace
Dressing
Dross reverberatory furnace
Cadmium recover
Reverberatory softening
Antimony recovery
Retorting
Cupelling
Bismuth refining
Waste
Dust
Dust
Flue dust
Effluent gas
Effluent gas
Off-gas
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Exhaust gas
Exhaust gas
Characteristics
Typical of ore being mined
Typical of ore being crushed
Pb, Zn, S plus Sb, Cd, Ge, etc.,
S02, traces HF, SiF4
Same as Copper
Particulate (oxides, sulfides,
sulfates of metals present) dilute
so2
Low SO;;, high participate, fumes of
volatile components
Low S02, volatile components of lead
bullion
Particulate, SO?, 803, nitrogen
compounds
Particulate, fume, data not reported
No data reported
Oxide fumes
Probably metallic fume (Zn, As, Pb)
particulate. No data were found
Metallic vapors (Zn, Pb, etc.)
particulate. Data not given
Cl, metallic fume
Quantities
110 g/metric ton ore
3.2 kg/metric ton ore pro-
cessed
See Table 3-12
100 to 250 kg particu-
late/metric ton ore
produced
125 to 180 kg particulate/
metric ton produced
20.02 volume percent SO;
1.0 to 21.7 g/m3 off -gas
(7)
S02 <0.05*
-
-
-
-
-
-
Control
Water sprays (manual )
Cyclone separators; fugitive
dust uncontrolled
Acid plants, particulates
collected by baghouses or
ESP
See Table 3-21
Baghouse
Baghouses or ESP
Baghouses or ESP
Recycle, baghouse
None reported
ESP or baghouse
Baghouse
Baghouse
None
CO
o
IO
(continued)
-------
TABLE 5-1 (continued).
Process
Zinc
Mining
Concentrating
Multiple hearth roasting
Suspension roasting
Fluidized bed roasting
Sintering
Horizontal retorting
Vertical retorting
Electric retorting
Oxidizing furnace
Electrolysis
Melting and casting
Cadmium purification and
casting
Waste
Oust
Dust
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Effluent gas
Fugitive
Fugitive
Flue dust
Characteristics
Characteristics of ore mined
Characteristics of ore mined
SO?, particulate (fumes and dust)
S02, particulate (fumes and dust)
SOj, particulate (fumes and dust)
SO,, particulate (fumes and dust)
Particulate (zinc oxide, cadmium,
copper, etc. ), S02
Particulate (zinc oxide, cadmium,
copper, etc. ), $03
Particulate (zinc oxide, cadmium,
copper, etc. ), S02
Combustion produced, Zn oxide
H.SO. mist, particulate (thought
to Include Zn, Pb, Cd, As)
Zn oxides, chloride compounds
Participates (Cd, Cu, Pb, etc.)
Quantities
0.1 kg/metric ton ore
1.0 kg/metric ton ore
6.3 m3 to 11 m3 S02/mine
Higher than above
Higher than above
Very low,S02; 9 to 100 g
part./m
8 kg particulate/metric ton
Zn produced, very low S02
50 kg particulate/metric Zn
produced. Very low $03
10 kg particulate/metric Zn
produced. Very low SO.
Similar to above
Est. 3.3 kg particulate/
metric ton Zn
-
514 kg/ day
Control
-
Cyclones
To Cottrell and acid plant
ESP, add plant
ESP, acid plant
No S0? control, settling
flues ESP's, baghouses for
particulates
Wet scrubber
Met scrubber
Impingement scrubbers
Similar to above
Electrolyte covers, addi-
tives for H2S04 mist
Scrubbers or fabric filters
Fabric filter
co
__j
o
-------
TABLE 5-2. ELEMENTAL ANALYSIS OF PARTICIPATE SIZE FRACTIONS
Average
partial late
size, y
14.3
7.5
4.6
2.3
1.3
0.58
0.11
<0.11
% by weight of sample3
Total
42.6
3.4
3.8
16.7
13.5
10.7
3.3
5.8
As
0.3
5.6
18.3
10.5
16.1
13.2
18.8
10.6
Pb
0.14
-
-
3.5
1.4
3.7
48
14.8
Cd
0.003
0.07
0.08
0.03
0.05
0.15
0.25
0.37
Ni
0.06
0.71
0.77
0.22
0.18
0.28
0.32
1.3
Sb
0.05
0.45
0.08
0.13
0.11
0.38
0.05
2.6
a Data collected by Puget Sound Air Pollution
1975.
Control Agency. August 14,
311
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All of the major pyrometallurgical processes release hot waste gases
that contain inorganic participates, 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 pro-
cesses within a smelter are frequently combined prior to control. It is not
unusual for the principal exhaust gas stream produced at a smelter to be
composed of gases from all the major processes. Other waste streams, con-
taining fewer particulates and less S02, are usually vented separately with
or without treatment. This practice of combining several 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 conventional
pyrometallurgical processes are batch operations, with gaseous emissions that
are not uniform. Unless the operations are carefully sequenced, 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; metallurgical operations therefore can result in
excessive emission rates. Smelter emission streams generally have a high
opacity due to the formation of sulfuric acid mist and the fine particulate
present.
The primary copper industry represents the most significant source of
atmospheric emissions, both S02 and particulate, in the primary nonferrous
industries. Most copper is produced in older plants with technologies that
have been in use for 50 years. Polluting process technologies such as
multiple-hearth roasters and reverberatory furnaces are only slowly being
replaced. In addition, the retrofit of controls is difficult and merely
shifts the problem of trace metals to problems of water pollution or solid
waste disposal. The primary lead industry is also a significant source of
trace metal particulate emissions, and S02 emissions are only about 50 percent
controlled. This industry faces similar retrofit problems as primary copper,
and adequate control would create additional problems with recovery and
disposal of solid residues. The primary zinc industry presents the fewest
atmospheric problems. Many older plants have closed in recent years, and
changes in process technology have allowed effective particulate control with
available equipment. An increase in the use of electrolytic processing has
reduced the potential for release of S02, and most sulfur is removed as a
concentrated stream in the roaster and converted to sulfuric acid.
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 temperatures of 1000°C or
312
-------
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 economi-
cally 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 usually have very tall
exhaust stacks to create a strong natural draft and thus allow a large quanti-
ty 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 dilution
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 larger
particulates. Most 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 typically contain 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 veloc-
ity enough for their collection, with the collected particles directly re-
cycled 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 minimize air
pollution but to prevent product loss. The third step involves removal of
fine particulates through the use of fabric filters and electrostatic pre-
cipitators, 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 temperatures
up to 370°C, collecting 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 both the final disposition of the gas and the
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
use of a hot ESP is more common.
Fabric filters 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 (approx-
imately 260°C). To prevent acid condensation, the stream is not cooled to
its dew point.
313
-------
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 efficiency rapidly for
particles less than 0.2 micron in diameter. Table 5-3 presents an analysis
of selected trace element emissions in participate matter from the ESP on a
copper smelter reverberatory furnace.
Trace metal compounds such as arsenic, antimony, and cadmium are vola-
tilized during high-temperature operations and enter the dust collection
system as fume. Because of inefficient control equipment and condensation of
the fume after the gas stream has passed through the collectors, stack gases
may contain significant quantities of trace metals. Recently field trests
have indicated that this condensation may be the single largest source of
trace metal emissions at smelters. High-temperature ESP's used at many
copper smelters cannot be expected to effectively control volatile metals.
The condensation problem is not as great at lead and zinc smelters, where
particulate collection systems are operated at lower temperatures.
Particulate collected by an ESP or baghouse is recycled or discarded
depending on its value. Many trace elements such as arsenic and cadmium tend
to concentrate in the fine particulate fraction. All zinc smelters process
particulate to recover cadmium. One copper smelter processes the particulate
to reclaim part of its arsenic content.
Atmospheric emissions from refinery operations are minor when compared
with those from pyrometallurgical process steps. However, many of the re-
fining operations are uncontrolled, and even where control devices are em-
ployed, their efficiency in collecting volatile metals has not been demonstrated.
S00 Control
—c.
The final step in treatment of some smelter waste gases is removal of
The U.S. copper industry has not achieved complete S02 removal, although
progress is being made. In many cases, smelter operators have relied on
dispersion of S02 into the air through tall stacks to minimize the ground-
l6?vel concentration of S02 in the vicinity of the smelter. Even the newest
smelters are being built with very tall stacks constructed of corrosion-
resistant materials for the purpose of ameliorating local air pollution by
dilution.
Air dispersion, however, is not an effective substitute for S02 control,
and for management of S02 in smelter gases, plant waste streams can be
divided into only two classes - those that are suitable for sulfur recovery,
usually in the form of 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 require-
ment, sulfuric acid manufacture via the double contact process represents the
best currently available control technology.
314
-------
TABLE 5-3. SELECTED TRACE ELEMENT EMISSIONS IN
PARTICULATE MATTER FROM THE ELECTROSTATIC PRECIPITATOR
ON A COPPER SMELTER REVERBERATORY FURNACE (1)
(kg/hr)
Element
Arsenic
Fluorine
Copper
Selenium
Antimony
Molybdenum
Nickel
Lead
Zinc
Cadmium
Test 1
34.5
4.3
0.8
0.29
0.01
0.1
0.01
0.2
0.1
0.04
Test 2
63.5
3.5
0.5
0.44
0.15
0.1
0.1
0.1
0.1
0.01
315
-------
Prior to acid manufacture, the gas stream must be conditioned to re-
move any residual particulate matter. This is accomplished in a series of
scrubbers and electrostatic precipitators. Particulate removed in these
steps forms a sludge which must be disposed of or recycled to the smelting
process. The remaining gas stream and the final acid plant tail gas con-
tain essentially no particulate matter. Acid plant emissions contain less
than 0.05 percent and 0.2 percent S02 for a double and single contact plant,
respectively. For even more effective control, any Japanese smelters clean
the tail gas from acid plants in scrubbers. Exit gas often contains 0.005
percent $62 or less, and concentrations of no more than 0.001 percent SOp
have been reported at some .facilities f2).
Sulfuric acid manufacture requires not only a high concentration of
SOp 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 sulfuric acid.
If the gas stream averages less than approximately 4 percent SO,
(or 2.5 percent as the absolute minimum, according to some advertisea
claims), control via wet scrubbing with a chemical solution could be imple-
mented. Two or three copper smelters have incorporated a DMA absorption
plant. This process can be used to produce a concentrated stream of SOp
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 S0£ suitable for acid manufacture. As of November 1977,
thirteen of the sixteen primary copper smelters were operating sulfuric
acid plants, the majority of which had been built since 1970. Two are
operated in conjunction with DMA units. Ground had been broken for con-
struction of an acid plant at one other copper smelter, the second smallest
in the industry, but work had stopped because of litigation. Four of
the six primary lead smelters had acid plants, the newest of which started
operation in October 1977. Of the two uncontrolled lead smelters, there
were plans for an acid plant at one (expected completion - 1979), and none
at the other, the smallest plant in the industry. All six primary zinc
smelters had acid plants.
Improvements in ore concentration have eliminated the need for most
multiple-hearth roasters in both the copper and zinc industries. Newer
processes have been adopted that use less fuel and yield more concentra-
ted exhaust streams. Oxygen enrichment has been utilized in a number of
smelter operations and may soon be used more commonly. This practice also
produces more concentrated exhaust streams.
316
-------
Less fundamental changes have been made at some smelters and could be used
at others to minimize the dilution of waste gas streams. These changes are
directed toward minimizing infiltration of outside air and substituting other
methods for cooling the gas stream. Copper converters formerly operated with
as much as a 30-centimeter gap between the converter and the hood. Substitu-
tion of water-cooled hoods allows a significant reduction of this gap. Re-
placement of some ducts with new ones made of high-temperature alloys has
minimized the need for infiltration air. Water-spray chambers have been in-
stalled to cool the gas stream. At some plants, holes and leaks in the duct-
work 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 country with flue gases
from coal-fired boilers. In one Japanese copper smelter, waste gases are
segregated into two streams, one with high S02 content that is sent to an
acid plant and one with low S02 content that is sent to a wet scrubber. The
scrubber treats all process gases before discharging them to the atmosphere,
including acid plant effluent, fire refining furnace flue gas, and ventila-
tion fan discharges that collect fugitive $62 emissions. More than 99 percent
of the sulfur that enters with the concentrate is contained. Other flue gas
desulfurization systems may also be employed. Use of organic acid deriva-
tives, such as salts or citric or lactic acid, has been investigated for use
as absorbents. A smelter in Sweden has installed a system using citric acid
absorption in combination with steam stripping for S02 recovery. The U.S.
Bureau of Mines, Salt Lake City, has developed the Citrate process. A
Canadian smelter uses a liquid ammonia solution as an absorbent to produce
either a concentrated S02 stream for further processing or ammonium sulfate
for fertilizer.
The Wellman-Lord S02 recovery process makes use of a sodium sulfite
scrubbing liquor. Contact with S02 converts the sodium sulfite to bisulfite
which is later heated for regeneration and concentrated S02 liberation. The
system has been applied to sulfuric acid plants and coal-fired boilers but
not a metallurgical process.
Lime/limestone and magnesium oxide absorbent systems are in operation at
Japanese smelters. These scrubbing systems are used at many coal-fired boilers
in the country. Many process variations are in operation in Japan utilizing
several types of additives to improve absorption efficiency. Aluminum sulfate
scrubbing is also practiced at two Japanese smelters.
317
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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, 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:
° 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 Rainfall and wind patterns
0 Feed composition
The effect of these emissions on ambient air quality may be appreciable since
they are emitted near ground level.
As described earlier, the S02 content of exhaust gas streams is maxi-
mized by preventing the infiltration of air into process exhaust ventila-
tion equipment. The elimination of dilution air may, however, cause in-
creased 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.
Fugitive emissions from pyrometallurgical processing equipment can be
greatly reduced through the use of hoods, enclosures, and exhaust systems for
charging and tapping operations. These devices can be vented through parti-
culate control devices and into the final discharge stack. Furnace configura-
tion and movement of the ladles to tap and charge various types of furnaces
present design difficulties and maintenance problems. Three U.S. copper
smelters now utilize fixed secondary hoods on their converters (3). Some
Japanese converters are equipped with swing-away secondary hoods, while at
other smelters in that country, the converter aisle is mechanically ventila-
ted with the exhaust from the entire building discharged through a baghouse.
This latter approach permits complete capture of converter emissions.
Although a potential problem with this type of system is providing adequate
ventilation for workers, no ventilation problems have been reported at this
smelter (2,3). One U.S. copper smelter has recently installed a building
evacuation system.
Particulate emissions from sources such as materials handling, roadways,
and slag piles may be reduced through use of covered conveyors, enclosures
and storage buildings, spray-wetting systems, frequent sweeping and good
plant housekeeping practices.
318
-------
Technology is available for all phases of fugitive emission control, 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 com-
plete unless provision is made for control of all air movements in the vicin-
ity.
References
1. Spaite, P.M., and M.J. Stasikowski. Overview of Environmental Impacts
Associated with Production of Nonferrous Metals. Prepared for Industrial
-Environmental Research Laboratory, U.S. Environmental Protection Agency.
Cincinnati, Ohio. December 1977.
2. Evaluation of the Status of Pollution Control and Process Technology -
Japanese Primary Nonferrous Metals Industry. Prepared by PEDCo Environ-
mental, Inc. for U.S. Environmental Protection Agency. Contract No.
68-02-1375, Task 36. July 1977.
3. PEDCo Environmental, Inc. Secondary Hooding for Peirce-Smith Converters.
U.S. Environmental Protection Agency. Contract No. 68-02-1321, Task 47.
Draft. December 1976.
319
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SECTION 6
WATER MANAGEMENT
Pollution of water by an industrial operation arises from three funda-
mental causes. Liquids may be intentionally discharged into natural waters;
these are the direct process wastes, easily quantified and recognized.
Liquids may also accidentally enter a watercourse in the form of fugitive
losses, difficult to quantify or predict. Finally, solids or gases may be
discharged that enter otherwise contaminated waters. These are secondary
sources, frequently the most difficult to locate and control.
The primary copper, lead, and zinc industries contribute water pollution
of 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 specific process.
Opportunities for fugitive liquid losses are limited in these industries.
Direct liquid wastes result from a number of process steps at the smelters.
As with gaseous emissions, individual liquid streams are typically combined
prior to treatment. Many impurities in the concentrates either result in
slags or furnace residues or are volatilized to be captured in dry air pollu-
tion control devices ore released to the atmosphere. However, where wet
scrubbers are used, as in the precleaning of gas streams prior to sulfur
recovery, metal compounds and organics accumulate in the scrubber water bleed
streams that must be discarded as they become contaminated. Because lead
smelters are located in areas of high precipitation, water containing flotation
agents is discharged from tailings ponds to surface waters during parts of the
year. Because of the relative effectiveness of air pollution control, waste-
water may constitute the principal environmental problem at primary zinc
smelters. As with any other industry involving liquid-handling, control at
smelters must be through the provision of facilities to capture and recycle
or treat process discharges or any accidental losses.
The major threat to water quality from these industries is secondary
pollution from the vast quantities of minerals that are extracted from the
earth and artificially-.exposed to the environment:. Drainage from nonferrous
mines, whether surface or underground, is often acidic in nature. It has
become increasingly evident that abandoned mines:'may constitute as great a
hazard as active mines. Wastewater from concentrating operations contains
metals, sulfur, and organic compounds and as with mining, the effectiveness
of treatment practices has not been established. Hydrometallurgical copper
32Q
-------
operations produce similar liquid wastes, both in the form of spent leach
liquor and runoff from stored residues. The remainder of this section is a
detailed examination of the character* and control of these problems, centering
on sulfide ores, recognized as a major contributor to water pollution.
Table 6-1 summarizes sources, quantities, and characteristics of liquid
effluents from mining, concentrating-, and smelting operations of these three
industries.
SOURCES OF SECONDARY WATER POLLUTION
Many elements appear in the wastewaters from the primary copper, lead,
and zinc industries. 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,
molybdenum and rhenium are frequently present in some copper ores and ger-
manium is found in Missouri lead and zinc ores. The platinum group of metals
is recovered from electrolytic slimes. Indium, gallium, thallium, cobalt, and
tin are found in some 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.
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, waste-
waters will be produced that must be pumped or drained from the workings.
Some of this water will be brought into the mine for equipment cooling, dust
control, and as life support for the miners. The remainder will be the
result of natural seepage, or in the case of a surface mine, from precipita-
tion.
2. Waste rock, or spoil, is produced during the mining process as
overburden in the construction of shafts and tunnels, or as low grade inclu-
sions 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 exposed to rainfall, seepage,
and surface run-off.
3. In mining the main ore body, 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. Stockpiles are also
exposed to the weather, and in some cases are purposely hosed down to minimize
blowing dust.
321
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TABLE 6-1. SOURCES OF LIQUID EFFLUENTS - MINING, CONCENTRATING, AND SMELTING OPERATIONS
Process
COPPER
Mining
Concentrating
Reverberatory smelting
Electric smelting
Sulfuric acid plant
DMA S02 absorption
Arsenic recovery
Fire refining and
anode casting
Electrolytic refining
Electrolytic purifi-
cation
Vat leaching
Waste
Seepage, runoff
Ore flotation water
Runoff from slag granu-
lation
Similar to above
Weak sulfuric acid
Water solution
Plant washdown
Anode cooling water
Electrolytic solution
"Black acid" and
electrolytic evaporate
Sluice water circulating
stream
Characteristics
More highly mineralized and acidic than
normal surface water; likely to contain
heavy elements found in ore body
Heavy elements plus thiosul fates, thio-
mates, organic and inorganic flotation
additives
Highly variable, heavy elements, sulfate
(see Table 2-20) open body of water
Similar to above
Sulfites, sul fates, suspended solids,
minerals, DMA
Ag, Cu, other metals, oil and grease
Ag, Cu, Zn
Sulfuric acid, other metals (see
Table 2-35)
Heavy elements, low pH
Comparable to concentrator plant
rich in iron
Quantities
Open pit: 0-0.3 m /metric
ton (7) of ore
Underground: 0.008 to 4
nr/metrlc ton
100 to 500 m3/ton of
concentrate (2)
-
4 to 8 m3/10 m3 gas
-
See Table 2-30
4000 to 13.000 ft/metric
ton metal produced.
see Table 2-33
-
-
350,000 to 1.000.000 i
spent liquor per metric
ton Cu produced
Control
Little control other than
site selection for mine
spoil dump
Tailings ponds, little
other control
To tailings pond if waste
produced
Similar to above
Neutralized, or other
(see Table 2-29)
Reuse in other processes
Tailing pond
Recycle, other processes
see Table 2-34
Reclamation, or pond
evaporation .
Tailings pond
Tailings pond
co
ro
PO
(continued)
-------
TABLE 6-1 (continued).
Process
Sulfide ore leaching
CLEAR regeneration
CLEAR oxidation
LEAD
Mining
Concentrating
Acid plant
Blast furnace
Slag fuming furnace
Final refining and
casting
ZINC
Mining
Concentrating
Waste
Sluice water circulating
stream
Sluice H20 .
Sluice water
Seepage, runoff, utility
use
Tailings slurry
Weak sulfuric add
Granulating H.O
Granulating H^O
Cooling water
Seepage, utility uses,
runoff
Flotation water
Characteristics
Comparable to concentrator plant
rich In iron
-
-
Dissolved and suspended solids
reflecting composition of ore
Small amounts of heavy metal In
solution high pH (8.5 to 11)
(see Table 3-6)
-
-
-
Participate matter (lead and lead
oxides)
Pb, Zn, associated minerals; metals
found in ore body
Tailings, unseparated minerals,
flotation reagents
Quantities
350,000 to 1.000,000 i spent
liquor per metric ton Cu
produced
-
"
Variable
4 m /metric ton ore
4 to 8 m /10 m gas
-
4 to 8 m /10 m gas
-
-
1000 to 16,000 m3/day
•
Control
Tailings pond
Settling pond
Settling pond
Liming and Impoundment
Tailings pond, biotlc con-
trol
Neutralized or other
Concrete settling pits
(see Table 3-25)
Concrete settling pits
(see Table 3-25)
Recycled, or tailings pond
or liming
Essentially same as copper
and lead
Lime precipitation,
settling ponds, sometimes
sulflde precipitation
00
ro
CO
(continued)
-------
TABLE 6-1 (continued).
CO
ro
Process
Vertical retorting
Electric retorting
Oxidizing furnace
Leaching
Melting and casting
Cadmium leaching
Cadmium precipitation
Waste
Gas wash water
Gas wash water
Gas wash water
Spent leaching solution
Cooling water
Probably very small
Filtration solution
Characteristics
Zn and other metal oxides, possibly
hydrocarbons, hydrolysis products
Zn and other metal oxides, possibly
hydrocarbons, hydrolysis products
Zn and other metal oxides, possibly
hydrocarbons, hydrolysis products
-
Suspended solids, oil, grease
-
Zn, cadmium chlorides and sul fates
Quantities
-
-
-
-
-
-
-
Control
Settling pits
Settling pits
Settling pits
-
-
-
-
-------
4. In one of the major processing operations, ores from the mine are
very finely ground and circulated through equipment as a slurry in water.
This process 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 decomposi-
tion products are common. Fugitive losses of dusts with similar characteris-
tics are also frequent. These materials are exposed to rainfall, and may
also be hosed down in attempts at housekeeping.
Unless controlled, water from all these sources will follow the natural
drainage pattern of the area and 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 more highly mineralized
than normal surface waters of the area, containing especially the sulfate
ion. They will be likely to contain measureable amounts of one or more of
the heavy elements found in the ore body, and will be more acidic than normal
surface waters of the area. The only exception is concentrator effluent,
which, as a result of treatment, is neutral or alkaline. This effluent has
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 requires
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.
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, organic matter, and chemicals added for processing all
accumulate. If their concentrations get too high, flotation is inefficient.
With the added incentive to minimize water discharges, most flotation
mills have approached this maximum limit of recycle. It is unlikely that
further reduction in volume of waste water from the concentrator will be
possible unless the industry adopts better waste segregation 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 wastes. At some milling operations,
high concentrations of partially oxidized sulfur compounds such as thiosalts
and thionates are produced. Some concentrators also discharge quantities of
cyanide salts and organic compounds. These materials must be biochemically
325
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oxidized if they are not to appear in the plant effluent. Some of these
materials oxidize only with difficulty and will not decompose in a simple
settling pond.
Various proposals have been advanced on how best to handle and treat
effluent from the concentrator, but at best these are theoretical exercises.
The fact is that in those industries it is the almost unanimous practice to
combine all waste streams into one. Water from mines, acid seepage, storm
drainage, plant floor drains, and even sewage from 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 considera-
tion 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 contamina-
tion of drainages with acid and heavy metals starts with iron pyrite (FeS2)
(1). This is the most common mineral in all sulfide ores. When this mineral
contacts 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
and are the natural weathering reactions of pyrite. If the outcrop is undis-
turbed, they take place slowly and the acid is quickly neutralized by reaction
with other minerals or with the natural alkalinity in surface water. The
iron hydroxide remains virtually in place, changing slowly into the mineral
1imonite.
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 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.
326
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When the pH drops below 6.5, conditions become favorable for the growth
of chemosynthetic 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 the pH of the water
continues to drop, action of the bacteria accelerates 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 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 minerals, 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 rain-
fall, and that they will be turned over often enough so conditions favorable
to the establishment of chemosynthetic bacteria will not develop. These
assumptions have been, tested primarily with coal stockpiles in reference to
acidity and iron content of the drainage. Tests specific to heavy metal
content of sulfide ore stockpiles have not been reported.
Run-off from smelter property has not been well quantified. Since many
of the minerals in the dusts on smelter property are partially oxidized by
smelter processing, theii rate of solution may be more rapid than from
unaltered sulfide minerals.
Spoil dumps are apparently quite variable in their effect on water
quality. Old dumps are not in many cases important contributors since they
may contain quantities of calcium- and aluminum-based minerals that themselves
decompose on weathering. These materials act to neutralize the acids pro-
duced and to re-precipitate heavy metals in new and often complex crystalline
structures. There is, over a period of time, a rearrangement of 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 percent calcium sulfate (2).
The problem with this analysis is that not all reactions take place at
the same rate; 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 greatest 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.
327
-------
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 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 Austra-
lian plants to the spoils of the Cloncurry district of that country (2). Few
plants other than metal-resistant strains can live in these soils (4).
Tailings are expected to be even more efficient in neutralizing any acid
that is formed, since 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 (6).
WASTE CHARACTERISTICS
Wastewaters from the mining and processing of copper, lead, and zinc
sulfide ores have one or more of the following characteristics:
1. High turbidity, containing both settleable and colloidal insoluble
inorganic material. Effluent from milling and ore concentration
processes is most representative of this characteristic, as is run-
off from smelter property.
2. Strong acidity, containing free sulfuric acid. Mine drainage and
seepage from sulfide containing tailings or spoil dumps is repre-
sentative.
3. Containing heavy metal ions, frequently toxic metals, in concentra-
tions higher than is allowable for discharge into public waters.
Most wastewater sources of these industries have this characteris-
tic.
4. Containing high concentrations of metallic and nonmetallic ions
that are not toxic in moderate concentrations. Most wastewaters
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 charac-
terized here as "pH adjustment".
328
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Correction of 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 demineralization. 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. This is
accomplished in these industries with the tailings pond, a basic waste treat-
ment system.
A 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 wastewaters in a sulfide mining
and ore processing operation is in principle very simple. Either the mineral
can be kept dry, or oxygen can be excluded to prevent the ore from weathering
(3). In practice, both are 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 also be possible to modify local drain-
age in such a way that ores do not stand in ponds of water or to design an
underground mine that will avoid seepage.
It is possible to provide cover for ore stockpiles and waterproof
storage for concentrates, and to seal spoil and tailings dumps with a water-
proof 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 (7), clay, tar (8), 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 feasibility 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 underground
mine sections, but thus far this has not been demonstrated to be practical.
The use of air-tight seals has been suggested for tunnels that are to be
329
-------
abandoned but that must continue to drain into operating sections (2). If
the air behind the seal were replaced with nitrogen, sulfide oxidation would
cease. Underground mines, however, contain numerous air passages, such as
cracks, joints, and fissures, which prevent effective air-tight seals. In
addition, atmospheric pressure changes allow the works to breathe (10).
Complete flooding of an underground mine section might also stop oxidation.
Neither method, however, would instantly stop pollution of the water due to
continued solution of the sulfite minerals already oxidized (2).
Sealing a spoil dump to exclude water, as described above, will also
exclude oxygen (2). Heavy compaction of some types of fine-grained waste
rocks may also prevent air from penetrating far below the surface. This
compaction may occur naturally in some deposits of tailings.
The addition of oxygen-scavenging chemicals such as calcium sulfite may
minimize the oxidation of the sulfides of heavy metals for a short period of
time. Sulfites may therefore assist in temporary stabilization until more
permanent treatment can be accomplished.
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).
The role of the chemosynthetic bacteria in increasing the rate of
formation of acids and soluble metals has been studied in recent years (11).
These bacteria are controlled in other industries by chlorination or ozoniza-
tion (12), but this procedure may be self-defeating when used with sulfide
ores. Antibiotics have been examined for control in mine waters, and 3 of 15
tested were found to be effective (13). 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
oxidizing bacteria (caulobactous) was purposely introduced into a mine (13).
One report states that there must be some undiscovered mechanism other
than 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. The literature cites serious pollution
incidents from this source. For example, one lead mine in Great Britain,
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 it con-
tained 140 milligrams per liter lead and 230 milligrams per liter zinc (14).
If exposed to air and water, there seems to be no limit to the degree of
v/eathering that can take place.
Wastewater Collection
If the assumption is made that any sulfide mine, spoil dump, tailings
bed, smelter property, or concentrator plant is a potential source of water
pollution, collection and treatment of all those waters is a major effort.
330
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In some areas, all operating units may be concentrated into a single reason-
ably small drainage area, but in other places, sources may be widely distri-
buted and discharge into several different watersheds. In an old, heavily
worked area, the identification 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 from the
practice of including sewage from operating facilities in process wastes. It
is recognized that the organic molecules thus introduced can seriously limit
reuse 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 simple treatment such as a septic tank or similar unattended
device, so the major cost in segregation of this waste is installation of
plumbing rather than the treatment plant itself.
If cooling towers or steam boilers are part of the mine, smelter, or
concentrator installation, blowdown from these sources also introduces chemi-
cals 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 better overall effluent quality might be obtained if dilution
were made downstream of the main treatment facility.
Highly acidic wastes, or wastes with high metals concentrations, can be
treated more effectively while they remain concentrated. If mine waters were
passed through a bed of coarse limestone and/or scrap iron prior to mixing
with other wastes, heavy metals concentration would probably be reduced and
less lime would be required for pH adjustment. Tests to confirm the effec-
tiveness of such simple treatments on specific streams are easy and inexpen-
sive to perform.
Separate treatment of cyanide wastes is virtually the only method for
preventing discharge of these poisonous ions, since cyanides are stable in
alkaline solutions. Cyanide decomposition is increased by exposure to ultra-
violet rays, aeration, and by certain types of aerobic bacteria. A pond
system would require minimum depth, aeration, and agitation to achieve these
conditions; in most instances this is not feasible (15). Mixing of cyanide
wastes includes a second hazard; 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
331
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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 kilometers before the water 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
(16). The large surface area speeds up oxidation processes and breaks down
many sulfides, sulfites, and organic materials. The retention time allows
most of the suspended solids to settle. In many cases, ponds 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 proceed 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.
Proper design of impoundment ponds is important to groundwater quality.
Most existing ponds leak effluent into the soil or rock upon which they are
constructed (17). A determination of the permeability of the underlying
material should be made prior to construction; if warranted, an impervious
lining should be installed. Table 6-2 lists 5 materials suitable for lining
impoundment ponds.
The embankment should be designed with consideration given to stability
and permeability. In some instances tailings slimes can be used to decrease
embankment seepage, if they are discharged from a spigoted line extending
around the embankment perimeter and if the ponded effluent is kept away from
the face of the embankment. This technique has been successfully used with
copper tailings slimes (17).
Runoff and precipitation may overload a pond system; the resulting
overtopping or breach could prove significant, especially in cases where toxic
effluent is being handled. The reservoir should be capable of accommodating
this additional input, or a diversion system needs to be utilized (17).
The apparent simplicity of the pond system has caused it to be widely
adopted. In an isolated mining area, it is frequently the least expensive
treatment system, requiring no utilities or energy and needing no highly
trained staff of people for its operation. In addition, the results obtained
by some tailings pond systems are impressive. One published report states
that a properly designed and constructed 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 (16):
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TABLE 6-2. COMPARISON OF TAILINGS POND RESERVOIR LININGS (17)
Lining type
Overall
rating9
Applications
Limitations
co
co
co
Imported Clayey
Soil
Tailings Slimes
Bentonite
Asphalt
Man-Made Sheeting
4
1
Most sites where clayey borrow
is abundant. Clayey over-
burden from pit excavation is
sometimes suitable.
Facilities producing high per-
centage of minus 400 mesh size
tailings.
Site with well-graded soil at
surface.
Sites with moderate temperatures
and relatively flat grades.
All sites.
Generally cannot be hauled economically
more than 1-2 miles. Chemistry of soil
may react adversely with effluents.
Compacted soil must have permeability
in 10-7 to 10~8 cm/sec range. Compac-
tion may be difficult if soil is very
dry or wet.
Sand fraction has to be removed. Need
relatively level site for deposition of
slimes. Need to compartmentalize
tailings disposal area.
Does not seal well in high or low pH
solutions.
Susceptible to cracking in cold weather
and on slopes.
High cost and requires smooth bedding.
Neglecting cost.
-------
Copper 30 yg/1 (extractable)
Zinc 150 yg/1 (extractable)
Lead 100 yg/1 (extractable)
Total iron 1 yg/1
These levels are lower than reported pilot plant studies using coagulation,
clarification, and filtration. Other reports state that best removal is with
greater alkalinity up to pH 11.0 (18).
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 wastewater or to
optimize the performance of the system. Any treatment will be trial and
error, and an error cannot be recognized until after several days the effluent
analysis indicates inadequate quality. Any change in treatment will then
require several additional days before there is an improvement in effluent
quality. Meanwhile, the pond is subject to continuous variation due to
changes in temperature, rainfall, and wind. The phenomenon of "thermal
skimming" may allow warmer wastewater to slide rapidly across colder pond
water. It is therefore difficult to assess the results of different treat-
ment 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 molecules is
aided by turbulence and works better at neutral or slightly acid pH. Removal
of heavy metals requires both precise control of pH at an alkaline level and
quiescent conditions for sedimentation (16), whereas in a pond, degree of tur-
bulence depends on the wind.
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 operator 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 any
more 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 treatment 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.
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Precipitation and Clarification - Secondary Treatment
The best known process for reducing the concentration of heavy metal
ions is chemical precipitation. The soluble ions are converted to insoluble
compounds by complex and variable chemical mechanisms. 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 precipitated:
copper, lead, zinc, cadmium, nickel, arsenic, chromium, manganese, and iron
(19). 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 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 the metals
concentration of the effluent (20). Practice has also shown that interactions
between the ions and with other substances in the water can greatly influence
the degree of removal. 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 adjustment,
flocculants are added to coagulate or agglomerate the precipitated metal
compounds and other suspended solids. Flocculants 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 flocculants.
These are usually proprietary mixtures, classifiable as polyelectrolytes or
polymeric materials, and are used either alone or in combination with the
metal salts. They are usually more effective than inorganics. Although more
expensive, organic flocculants 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 occasionally
practiced, usually either to add weight to the floe to increase 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 (21).
335
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The equipment used for this clarification process is most often a
commercially packaged clarifier. There must be three distinct 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 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 flocculation, and
in the treatment of a large flow of water, sizable amounts may be required.
The sludge concentration section operates on either of two principles.
The simplest is gravity settling, in which water is allowed to move slowly
through a large basin or pond while 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 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 par-
ticles of precipitated heavy metals and other suspended solids.
A recent pilot plant study used a polymer flocculant in the clarification
of sulfide mine drainage. The concentrations of metals in the effluent are
presented in Table 6-3.
TABLE 6-3. METAL CONCENTRATIONS IN EFFLUENT
FROM A MINE DRAINAGE CLARIFIER (22)
Copper
Zinc
Lead
Iron
mg/1, extractable
Average
0.05
0.37
0.25
0.28
Minimum
0.03
0.13
0.01
0.09
Maximum
0.11
0.90
0.62
0.60
These apparently high concentrations do not reflect the degree to which
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 milligram per liter, respectively. On the other hand, most of the
copper passed through the filter as a soluble ion. Its average concentration
was 0.04 milligram per liter after filtration.
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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 overcome (22). Returning
it to the tailings pond may upset the pond's complex chemistry, especially if
organic flocculants have been used. It could be discharged to its own pond
for storage in perpetuity, but any overflow would undoubtedly be highly
mineralized and would itself require treatment. It could be dewatered 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. Experiments 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 feasible, with
the resulting ash recycled to a pyrometallurgical process for recovery of
valuable metals.
Filtration
It is known that clarification alone cannot remove all suspended solids
from water. Some 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 clarifier effluent was present as
insoluble particles, capable of being removed by filtration (22).
If filtration is necessary, the equipment that is most suitable for
handling large volumes of water is the "rapid sand filter." 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 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 reconcentrated;
solids leave as part of the clarifier sludge.
Sand filters represent a large percentage of the capital cost of a water
treatment facility; they also account for much of the operating cost in terms
of labor.
It is not possible to use sand filtration as a substitute for floccula-
tion 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.
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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 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 filtration 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 application to a
major waste stream, however, do not recommend it.
Most other types of mechanical filters used throughout processing
industries are much too small for application to fine filtration 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, a point is finally reached where they cannot handle a reasonable
quantity of water. They are then allowed to dry, and the accumulated mate-
rials 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 treatment or in
sludge disposal.
Post-Treatment
Most effective removal of heavy metal ions will probably require precipi-
tation at a pH that is more alkaline than the allowable level for discharge
to a lake or stream. Reacidification 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. The
combustion gases, containing carbon dioxide, bubble up through the water,
338
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forming carbonic acid. This is a weak acid, easily controlled. There is no
danger of over-acidification since 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. 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 reacidification. Sulfuric
acid will undoubtedly be employed in the copper, lead, and zinc industries.
Although 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 pH (22). If this drop in
pH also applies to U.S. wastewaters, extended storage in a special pond prior
to release may be the best overall environmental solution.
Basic Treatment Effectiveness
Basic treatment methods outlined in the previous sections can never be
totally effective. Given enough time, many of the heavy metal compounds will
become soluble when exposed to the moist, oxygenated environment of the
earth's surface. Eventually, the metals artificially 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 down of 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 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 milligram per
liter, lead to less than 1.0 milligram per liter, and zinc to less than 10
milligrams per liter. As mentioned previously, pilot plant tests (23) 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 milligram per liter for lead and 0.15 milligram
per liter for zinc (19). Theoretically, copper could be reduced to 1 to 8
micrograms per liter, lead to 1 microgram per liter, and zinc to 10 to 60
micrograms per liter.
There is much less definitive information regarding the effectiveness 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 as a positive ion in a low energy compound. This would include nickel,
cadmium, bismuth, manganese, and chromium. In Canada, it has been reported
that 0.04 milligram per liter is believed to be an attainable level for
cadmium (18). This same report indicates that arsenic can be consistently
reduced to 0.05 milligram per liter. There is, however, little operating
data or theoretical evaluation of the effectiveness of basic treatment in
339
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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, clarification, and
filtration. Many methods proposed for this industry, or being used in other
industries, are not applicable to large volumes of water containing large
amounts of suspended solids, high concentrations of dissolved solids, and
much smaller concentrations of 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 cases such as reverse
osmosis and ion exchange, a clear water, free from suspended solids, is
necessary. In other cases, cost may be prohibitive 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 impurities. If
water is evaporated, minerals remain as a sludge or scale or as a more con-
centrated solution. If water is frozen, the ice that forms is relatively
pure and minerals are concentrated in the remaining liquid.
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 by
freezing (23). 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 (24). Capacities 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 suggested is
electrochemical. It is proposed that wastes be sent through electrowinning
cells to deposit heavy metals at the cathode. No data exists on this method
of waste treatment, but it is likely that 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 coagu-
lation and filtration for removal.
340
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Chemical absorption methods have been proposed. A few chelating chemi-
cals react quantitatively with selected metallic ions. The chelates, plus
unreacted chelating chemicals, are then extracted from the solution. The
chelates are broken, metals are salvaged, and chelating materials regenerated
to be recycled. This technique is being used to extract copper from solution
in some processing operations. Several developing hydrometallurgical pro-
cesses in the copper industry may use this extraction technique. For treat-
ment of wastes, however, chelating chemicals have the disadvantage of being
specific to only one or two ions, usually copper and iron. The published
literature indicates none that react with most other metals. This procedure
may become important if demand exists for development of the required chemical
materials.
Addition of sodium sulfide to wastewaters has been proposed as a method .
for the removal of heavy metal ions from a waste stream (19). Solubilities
of sulfides of the metals is much less than their carbonates or hydroxides.
A Swedish smelting facility is currently installing a system for treatment of
process and runoff water in which sodium sulfide will be used to precipitate
dissolved metals. Initially, sodium hydroxide is added to adjust the pH to
4.5 to 5, at which point sodium sulfide is admixed until the solution holds
a pH of 5.5 (25). Tests have indicated that addition of the sodium sulfide at
pH 4.5 practically eliminates the problem of hydrogen sulfide liberation
while achieving reasonable arsenic precipitation (25). To further insure
against hydrogen sulfide emissions, the treatment tanks will be kept under
pressure. Treatment in a Dorr thickener and dewatering by vacuum filtration
will follow precipitation. The filter cake will be recycled to the roaster for
metals recovery. Recycle of the precipitate necessitates effective treatment
of roaster off-gases due to their high arsenic content; where such control
is absent, alternate disposal strategies would be necessary. Effluent over-
flow from the Dorr thickener will be filtered and treated with calcium
hydroxide for precipitation of fluoride, calcium sulfide, and calcium sulfate..
These precipitates will be landfilled (25).
Two classes of chemical treatments are in use that remove metal ions from
water hy substituting other more desirable ions. One, usually called "cemen-
tation," uses ah 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. The overriding factor in determining the effectiveness of copper
cementation is the concentration of copper in the feed solution. A concentra-
tion of 300 milligrams per liter may yield 50 percent recovery given optimum
conditions including very low pH, agitation, and absence of oxygen or
hexavalent chromium. Under these conditions, it has been shown that iron shot
could reduce a copper solution to less than 1 milligram per liter in 10 minutes
or less (25). Even if conditions are not optimum, however, reduction of
metals concentration is significant, especially in acidic streams. The pro-
cess is in use with some mine drainage streams, and its application could be
increased if guidelines for its use were available. Given the requirements
for high copper concentration and low pH, the most efficient application of
this process could be achieved through effluent segregation. A special appli-
341
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cation of cementation has been studied in which zinc dust was used to remove
selenium from a smelter scrubber effluent (26). Similar applications may be
applicable to wastes from mining and concentrating. Copper recovered by cemen-
tation can be recycled for conventional treatment.
The second metal substitution process is ion exchange. This is considered
by some authors as the standard method for concentrating heavy metal ions (27),
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 potassium. These elements exist in wastewaters from mining
and milling at concentrations often thousands of times greater than the 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
heavy metal ions are removed.
Ion exchange has another disadvantage. When the resin is regenerated,
the heavy metals reappear as a solution mixed with large amounts of regener-
ating chemical, usually a sodium chloride brine. This presents 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 sulfates 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 membrane. This seems to occur quite rapidly, and if the strain of
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 western states, and a
group of plants has evolved on these soils that accumulates selenium in their
tissues. They are well known by ranchers of the West, who call them collec-
tively "poison vetch" because they cause selenium poisoning in livestock that
consume them. The group consists 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 milligrams per liter (29).
When they die, these plants may return so much selenium to the soil that
other plants growing on that spot can become toxic.
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There has apparently been no research to establish whether or not the
ability of Astralagus to concentrate selenium can be of benefit in waste
treatment. All of the plants are adapted to highly mineralized, poor quality
soils. It is not beyond possibility that they could thrive on seleniferous
tailings, or be irrigated with seleniferous wastewaters. 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. If the mechanism by which the element is accumulated by the
plants were studied, it might open a new route for articifial 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
(29). 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 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.
References
1. Smith, E.E., et al. Sulfide to Sulfate Reaction Mechanism. U.S. Depart-
ment of Interior, Federal Water Quality Administration. FPS 02^70.
2. . Glover, H.G. Acidic and Ferruginous Mine Drainages. The Ecology of
Resource Degradation and Renewal. Blackwell Scientific Publications.
London. 1975.
3. Mine and Mill Wastewater Treatment. Canadian Environmental Protection
Service. Report EPS 3-WP-75-5.
4. Tribe, I. The Plant Kingdom. Grossett & Dunlap. New York.
5. Wainwright, S.J. et al. Physiological Mechanisms of Heavy Metal
Tolerance in Plants. The Ecology of Resource Degradation and Renewal.
Blackwell Scientific Publications. London.
6. Berg, W.A., E.M. Barrus, and L.A. Rhodes. Plant Growth on Acid Mill
Tailings. Colorado State University. Fort Collins, Colorado.
7. Use of Latex as a Soil Sealant to Control Acid Mine Drainage. Environ-
mental Protection Agency. Water Pollution Control Resource Service.
EFK 06-72.
8. Hall, D.A. The Sealing of Coal Dumps with Road Tar. J. Institute of
Fuel 40, 474-6.
343
-------
9. Feasibility Study of Mining Coal in an Oxygen-Free Atmosphere. U.S.
Department of the Interior, Water Pollution Control Resource Service.
DZM 08-70.
10. Processes, Procedures and Methods to Control Pollution from Mining
Activities. EPA-430/9-73-011. U.S. Environmental Protection Agency.
Washington, D.C. October 1973. pp 221-257.
11. Lau, C.M., et al. The Role of Bacteria in Pyrite Oxidation Kinetics.
Proceedings of 3rd Symposium on Coal Mine Drainage Research. Pittsburgh,
Pennsylvania.
12. Beller, M., C. Waide, and M. Steinberg. Treatment of Acid Mine Drainage
by Ozone Oxidation. U.S. Environmental Protection Agency, Water Pollu-
tion Control Research Service. FMH 12-70. 1970.
13. Shearer, R.E., W.A. Everson, and J.W. Mausteller. 2nd and 3rd Symposium
on Coal Mine Drainage Research. Pittsburgh, Pennsylvania. 1968 and
1970.
14. Jones, A.N., and W.R. Haurall. The Partial Recovery of the River
Rheidol. The Ecology of Resource Degredation and Renewal. Blackwell
Scientific Publications. London. 1975.
15. Hyatt, D.E. The Chemical Basis of Techniques for the Decomposition and
Removal of Cyanides. Fall Meeting of Society of Mining Engineers of
AIME. Salt Lake City, Utah. September 10-12, 1975.
16. Bell. A.V. The Tailings Pond as a Waste Treatment System. Canadian
Min. and Metal 1. Bulletin. April 1974.
17. Toland, G.C., and R.E. Versaw. Design of Impoundment and Evaporation
Ponds and Embankments for Cyanide and Other Toxic Effluents. Fall
Meeting of Society of Mining Engineers of AIME. Salt Lake City, Utah.
September 10-12, 1975.
18. Base Metal Mine Waste Management in Northeastern New Brunswick. Canadian
Environmental Protection Service. Report No. EP S8-WP-73-1. 1973.
19. 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 Volume 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.
20. Unpublished operating data. Water Department. Batesville, Arkansas.
21. Ford, T., and J.F. Boyer. Treatment of Ferrous Acid Mine Drainage With
Activated Carbon. U.S. Environmental Protection Agency, Environmental
Protection Technology Service. EPA-R2-73-150. 1973.
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22. Bell, A.V., et al. Some Recent Experiences in the Treatment of Acidic
Metal-Bearing Mine Drainages. Canadian Min. & Metall. Bulletin. December
1975.
23. Purification of Mine Water by Freezing. Environmental Protection Agency,
Water Pollution Control Resource Service. DRZ 02-7k. 1971.
24. Acid Mine Waste Treatment Using Reverse Osmosis. Environmental Protection
Agency, Water Pollution Control Resource Service. DYG 08-71. 1971.
25. Lindquist, B., K. Lindegren, and C. Sund-Hagelberg. Water Pollution Work
at the Rb'nnskar Factories. Vatten 2-76. pp. 144-154.
26. Merchant, W.N., R.O. Dannenberg, and P.T. Brooks. Selenium Removal from
Acidic Waste Water Using Zinc Reduction and Lime Neutralization. U.S.
Department of Interior. Salt Lake City Metallurgical Research Center, 1975.
27. Lund, H.F. Industrial Pollution Control Handbook. McGraw-Hill. New York.
1971.
28. Case, O.P. Metallic Recovery from Wastewaters Using Cementation. Envi-
ronmental Protection Agency Office of Resource and Development. 1974.
29. Kingsbury, J.M. Deadly Harvest. Holt, Rinehart, and Winston. New York.
345
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SECTION 7
SOLID WASTE MANAGEMENT
The direct environmental impact of solid waste production in the primary
copper, lead, and zinc industries is difficult to assess; its secondary
effects are considerable, however. Exposing large quantities of sulfide
minerals to the atmosphere leads to oxidation and sulfuric acid formation; it
is the largest source of water pollution in the industry. Transporting and
stockpiling of ores and by-products are sources of dust as well. Solid
wastes produced at smelters include slags, fume and particulate matter cap-
tured in air pollution control equipment, sludges from electrolyte regenera-
tion and purification, electrolytic slimes, and water treatment sludges.
Many of these waste streams are recycled or processed further for recovery of
their metal values. Table 7-1 summarizes the sources, quantities, and char-
acteristics of the solid wastes from the primary copper, lead, and zinc
industries.
SOURCES AND CHARACTERISTICS
Solid waste produced in these industries consists of sulfide ores and
their components, processing reagents, or products of chemical reactions
among these constituents.
The mining of copper, lead, and zinc ores resulted in production of
412.7 million metric tons of solid waste in 1974 (1). Of this total, 368.2
million metric tons were waste rock and 44.5 million metric tons overburden
(1). About 2 tons of mine waste are generated per ton of raw nonferrous ores
produced. However, this figure can vary considerably depending upon the ore
body; because of the declining quality of ores, about 3 tons of waste result
from each ton of raw copper ore produced. Waste rock consists of low grade
ore, "country rock," or sulfide ores of metals other than the one being
produced. Overburden from surface mines is usually a mixture of soil and
rock with varying physical and chemical properties (2). Sulfide concen-
trations are generally low in the overburden.
Exposure of waste rock to the oxidizing environment at the earth's
surface can lead to oxidation of sulfides and production of sulfuric acid.
The acid solution is capable of leaching the metal content of the ore. This
process and its effects on water quality are discussed in Section 6, Water
Management.
Concentrator tailings amounted to 253 million metric tons in 1974; the
copper industry alone contributed 241 million metric tons (1). Copper mill-
ing wastes are primarily composed of rock-forming minerals, but approximately
346
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TABLE 7-1. SOURCES OF SOLID WASTE FROM MINING, CONCENTRATING, AND SMELTING OPERATIONS
Process
COPPER
Mining
Concentrating
Reverberatory smelting
Electric smelting
Slag treatment
Sulfuric acid plant
Scrubbers
Vat leaching
Sulfide ore leaching
Cementation
CLEAR regeneration
CLEAR oxidation
Waste
Waste rock, overburden
Tailings
Slag
Slag
Slag
Gypsum from acid
neutralization
Solid residue
Tailings
Tailings
Scrap iron
Filter cake
Solid residue
Characteristics
Sulfides, country rock, soil
Sulfides, heavy metals, milling
reagents, gangue
Iron silicates, heavy metals, CaO, HgO,
Al-0,, copper, sulfur (see Table 2- )
Similar to reverberatory slag, less
copper and sulfur
Similar to reverberatory slag
-
May contain volatiles, particulates, etc.
Comparable to concentrator tailings
Comparable to concentrator tailings
-
Iron sul fate/hydroxide, other
Colloidal sulfur, soluble chlorides,
cyanides
quantities
380.5 MMT (1974)(dry weight)
(1)
241 MMT (1974)(dry weight)
(1)
Up to 3000 kg/metric ton
copper produced
-
-
-
. •
'
-
2 to 4 weight percent of
spent electrolyte
-
Control
Spoil dump located to
control seepage
Tailings pond
Granulation
: Similar to rever-
beratory slag
Similar to rever-
beratory slag
Pond
Usually recycled
Concentrator tailings
pond
Concentrator tailings
pond
-
Mixed with CLEAR oxida-
tions wastes and
ponded
Settling pond
CO
-p>
-vl
(continued)
-------
TABLE 7-1 (continued).
Process
LEAD
Mining
Concentrating
Sintering
Blast furnace
Slag fuming furnace
Dressing
Kettle softening
Harris softening
Antimony recovery
Bismuth refining
ZINC
Mining
Concentrating
Multiple hearth roasting
Waste
Waste rock, overburden
Tailings
Slag
Slag
Slag
Slag
Slag
Leached slag
Excess arsenic trioxide
Slag
Waste rock, overburden
Tailings
Solid waste
Characteristics
Sulfide ores, country rock
Dolomite sulfides, heavy metals, milling
-
FeO, CaO, Zn, MgO, A10, Si02> metals
Compounds of Fe, Ca, Si, Al , Mg, etc.
-
Lead, sodium oxide salts of arsenic,
tin, antimony
•
-
Ca, Mg, Pb, lead chlorides
Country rock, sulfides
Dolomite, reagents
Quantities
0.13 ton/ton ore (1973)(3)
0.9-1.1 tons/ton ore milled
-
-
-
-
-
-
-
40 percent of dross feed
-
0.9 to 1 ton/ton ore milled
-
Control
Dumps, landfill,
backfill
-
Granulated and ponded
Concentrator tailings
pond
-
Dumped with blast fur-
nace or fuming furnace
slag
Dumped with blast fur-
nace or fuming furnace
slag
Sale
Settling pond with slag
from smelting
Spoil pile, pond,
backfill
Tailings pond, backfill
-
OO
4*
00
(continued)
-------
TABLE 7-1 (continued).
Process
Vertical retorting
Electric retorting
Oxidizing furnace
Leaching
Purifying
Electrolysis
Cadmium leaching
Cadmium purification and
casting
Waste
Residue
Residue
Residue
Filter cake
Copper cake residue
Filtration residue
Precipitate
Residue
Furnace residue
Filter cake
Characteristics
Lead, copper, silver, gold, nickel, ger-
manium, gallium, aluminum, magnesium,
manganese, blue powder
Similar to vertical retorting
Zinc, cadmium, chromium, copper, lead
slag (see Table 4-17), coke, ferro-
silicon
High metals content
Copper, cadmium, indium, thallium,
gallium, germanium
Lead, precious metals, iron
Copper, germanium, cadmium, nickel,
arsenic, antimony, cobalt, Iron
Lead, zinc, cadmium, indium, arsenic,
copper, silver, gold, silicates, quartz
Cadmium, zinc, lead
Iron, arsenic, indium, mercury, copper
Quantities
-
-
350 Kg/metric ton of zinc
oxide
360 Kg/metric ton of zinc
produced
_
-
-
1.8 Kg/metric ton zinc
produced
Control
Open slag dumps,
recycle
Similar to vertical
retorting
Dried, briquetted for
furnace feed
Sent to lead smelters,
leached
Sent to copper smelter
Sent to lead smelter
After lagoon storage,
sent to copper or
lead smelters
Recycle
Dump, pond with lime
treatment, or recycle
to sinter
GO
-P>
vo
-------
15 percent of the heavy metals originally in the ore are found in these
tailings. Milling reagents may also be included in the tailings. Because
tailings contain quantities of various heavy metals in concentrations above
background levels, the effects of sulfide oxidation and leaching of these
materials may have greater impact than for mine wastes. Newer copper concen-
tration practices are leaving a higher percentage of pyrite in the tailings.
This factor, plus the steadily decreasing metal content of available ores, is
causing the production of many more tons of tailings per ton of copper than
in years past. Environmental impact is primarily through secondary water
pollution (Section 6).
Additional solid wastes are generated throughout the industry, although
quantitative data are lacking for many processes. Pyrometallurgical opera-
tions produce dusts from gas cleaning, slags, and precipitates and dewatered
sludges from water treatment. During the electrolytic refining of copper,
metals that are insoluble in the electrolyte, such as gold, silver, selenium,
and tellurium, settle into the slime residue at the bottom of the cell. The
spent electrolyte from zinc electrolysis contains metals such as arsenic,
cadmium, cobalt, copper, nickel, and tin solution; these impurities are
precipitated and separated as the electrolyte is recycled.
WASTE TREATMENT
Treatment and handling of spoil from mining activities are most effective
when secondary water pollution resulting from runoff and seepage is pre-
vented. Segregation of surface mine overburden allows recovery of soil;
replacement of soil aids reclamation by providing a substrate for revegeta-
tion and by curtailing oxidation of sulfides through elimination of excess
contact with air (3). Proper grading of spoil is also important in both
these regards.
Spoil from underground mines is piled or used as backfill. Either
method may lead to secondary pollution of ground or surface water, and im-
proper piling may result in collapse. Consideration of topography and hydro-
geology is necessary when disposal sites are selected. A more detailed
discussion of spoil disposal is found in Section 6.
In the copper industry, an important control method is dump leaching,
which is now becoming economically attractive as a means of recovering
additional copper, disposing of excess smelter acid, and simultaneously
enclosing the spoil in controlled areas where slow sulfide oxidation is less
likely to cause long-term pollution of watercourses.
Concentrator tailings are slurried and sent to settling ponds where the
solids accumulate through the evaporation, recycle, or discharge of the
liquid fraction. Solid constituents have considerable impact on effluent
quality and residue piles constitute a dust or sediment source (3). The
small grain size of the milled material exacerbates the situation. The
location and construction of tailings ponds are critical in determining the
extent to which surface or groundwaters are impacted. Important design
considerations include permeability, enbankment freeboard, retention time,
and the discharge system. The ponds should be designed to allow for their
350
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eventual abandonment and revegetation. Oewatered tailings may be used for
backfill of abandoned stopes or as landfill. Section 6 contains further in-
formation on these problems.
Control of solid waste from smelting operations generally consists of
combining the material with concentrator tailings for ponding or, if there is
sufficient metal value remaining, reprocessing on-site or at other facili-
ties. In general, the dusts from air pollution control devices can be
recycled; however, the build-up of arsenic'and other trace metals may require
their disposal. In some cases, dusts not suitable for recycling are pro-
cessed further to recover by-products. Some slags may also cause disposal
problems; kettle softening slags in the primary lead industry contain soluble
salts of arsenic, tin, and antimony. Copper electrolytic slimes are pro-
cessed to recover metals of value such as gold and silver, leaving a residue
for disposal. The precipitates from the purification of zinc electrolyte are
also processed further for recovery of cadmium and other metals.
The combining of solid wastes adds to the difficulties of wastewater
treatment discussed in Section 6.
References
1. A Study of Waste Generation, Treatment and Disposal in the Metals
Mining Industry. U.S. Environmental Protection Agency, Office of Solid
Waste Management Programs. NTIS PB-261 052. October 1976.
2. Processes, Procedures, and Methods to Control Pollution from Mining
Activities. EPA-430/9-73-011. U.S. Environmental Protection Agency.
Washington, D.C. October 1973.
3. Williams, R. E. Waste Production and Disposal in Mining, Milling, and
Metallurgical Industries. Miller Freeman Publications, Inc. San Fran-
cisco. 1975.
351
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SECTION 8
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 econom-
ical operations that will result in fewer environmental problems. The
greatest amount of emerging technology is for the production of copper. To
some extent, the lead and zinc industries have preceded copper in adopting
new procedures or building new plants. Recent economic events have dis-
couraged 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 technol-
ogies that offer great promise in terms of improved economics and reduced
environmental impacts. There are several new hydrometallurgical processes at
the laboratory and demonstration plant scale that have permitted recovery of
metals without the use of high temperature reduction. Unique technologies
are being developed that for the first time permit separation of metals from
certain raw materials or that eliminate steps usually taken in conventional
processing. There are several possible methods of treating slags from con-
tinuous pyrometallurgical processes. This section of the report will outline
many of these developments. Coverage of the various processes varies con-
siderably; many of the new technologies are proprietary, and little informa-
tion has been released. In each case, a brief description of the process and
status of the technology is given. Where available, data on energy require-
ments, 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, partic-
ularly 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 (1) - In a process sometimes known as the
Betts process, many foreign lead smelters eliminate the pyrometallurgical
purification steps used in the U.S. These steps are thought to be major
sources of pollution by trace elements. Electrolytic refining permits direct
recovery of bismuth and other elements from electrolytic slimes. Impurities
are removed together, allowing for their later individual recovery and pro-
ducing refined lead in a single step.
352
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Process Description - Lead bullion is first prepared by conventional
pyrometallurgical steps. After dressing to remove copper and tin, the
bullion is cast into anodes. The anodes are placed in concrete electrolytic
cells, usually lined with asphalt or polyethylene. The electrolyte is a
solution of lead fluosilicate and fluosilicic acid (H?SiFfi), and is typically
analyzed as lead, 67 grams per liter; free H2SiFg, 95 grams per liter; and total
acid, 142 grams per liter. Electrolyte temperature is maintained at about
40°C.
Energy Requirements - Current density is 160 to 190
meter 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 of Lead from Galena (2) - Australian research has
indicated that it is possible to extract lead directly from galena concen-
trates by electrolysis. This development would replace the conventional 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 deposi-
tion, and it is reported that 75 percent of the lead in the concentrate is
extracted. Current efficiency is 85 percent, a considerably higher figure
than is found in conventional electrolysis.
Energy Requirements - Total energy requirements are about 2600 kilo-
calories per kilogram of lead produced.
Environmental Residuals - Although details of the residual slurries or
recycling operations are not available, it is likely that the perchlorate
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.
Purification of Zinc Electrolyte - During the electrolytic manufacture
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 electrolytic
cells. Treatment reagents such as zinc dust, organic chemicals, and occa-
sionally arsenic and copper sulfate are used at U.S. zinc smelters. Purifi-
cation is multi-stage, with heavier elements precipitated out first, followed
by the most soluble metals. There are three additional processes for purifi-
cation of the leach liquor in use at foreign smelters.
353
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Process Description - The formation of jarosite for iron precipitation
can be used to purify the zinc electrolyte (3,4). Jarosite is an insoluble
hydroxylated sodium iron sulfate (NaFe-JSOJpiOHjg). 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 (3,4). Goethite,
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" process
(3,4) 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. Substitution of antimony for arsenic as a
treatment reagent may be of environmental 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 environmen-
tal advantages or disadvantages, use of the PRC process increases the pro-
duction rate for an electrolytic copper refinery. Using this technique, each
cell can process about 15 percent more copper than a conventional 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 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 amperes per square
meter can be used.
Environmental Residuals - PRC waste streams are no different from
conventional refineries in proportion to the quantity of copper processed.
Status of Technology - At least two copper refineries in Sweden and
Japan are operating with the PRC process. There may also be other installa-
tions using this technology.
354
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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 mechanical detail and fur-
nace configuration. Although all the designs are based on sound theoretical
and practical calculations, continuous smelting 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.
WORCRA (6,7,8) - 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 claimed to be one of the prin-
cipal 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 resulting 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 con-
tent of the slag is continuously reduced as it moves through the slag-clean-
ing zone. The matte entering the converting zone is reduced to blister
copper which flows continuously from the furnace. The blister copper con-
tains 97.0 to 98.5 percent copper and 0.6 to 0.9 percent sulfur, a low-grade
product requiring considerable fire-refining before it can be electrolyti-
cally refined.
Energy Requirements - Fuel consumption is expected to be 50 to 60
percent of that needed for a reverberatory-converter process of comparable
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.
355
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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 commercial installa-
tions.
Mitsubishi (6,7,9,10,11,12) - 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 require-
ments 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 enrichment 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 cleaning, 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 that blow air or oxygen-enriched air and lime-
stone 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 is analyzed
as 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 re-
verberatory 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 increased far
more than in conventional furnaces, which have separate reaction zones for
oxidation and settling. Conventional furnaces operate at a slower rate
because oxidation is promoted by good mixing while settling requires quies-
cent 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 oercent,
and can be as high as 50 percent when coupled with oxygen enrichment. Energy
consumption per metric ton of product has been reported as follows:
356
-------
Oxygen 0.3 metric tons
Fuel oil 2.4 million kilocalories
Natural gas 360,000 kilocaiories
Electricity 400 kilowatt-hours
Total energy consumption is 3.3 million kilocalories per metric ton of copper,
as opposed to 6.4 to 7.2 million kilocalories per metric ton for conventional
reverberatory smelting.
Environmental Residuals - The mixed exhaust gas leaving the furnaces
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.
Flue dusts, consisting of copper oxides and minor elements, are gener-
ated 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 buildup 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 slag
granulation, 50,000 liters; acid plant blowdown, 14,000 liters; and contact
cooling, 7,800 liters.
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.
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 operat-
ing at Naoshima and a 118,000 metric ton per year plant is being designed
for the proposed Kidd Creek smelter of Texasgulf.
357
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Kivcet (7,13,14) - The Kivcet process is being developed in the Soviet
Union. It is said that any combination of copper, lead, and zinc concen-
trates may be used as feed, simplifying concentrating procedures and permit-
ting greater by-product recovery.
Process Description - The Kivcet smelting process contains three steps.
In the first, dried concentrates are reduced in a cyclone furnace 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 concentrate
feed because of their high temperature. Zinc metal is recovered directly
through condensation of off-gases from the cyclone furnace.
Status of Technology - A 45 metric-tori-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 (15,16) - Dr. M. Paul E. Quenaw and Dr. Reinhardt
Schulmann, Jr., of the United States have developed a continuous smelting
process that uses pure oxygen for the autogenous 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, overall
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.
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 particulate
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.
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Momoda Furnace (7) - One replacement for the conventional copper reverb-
eratory furnace is the Momoda furnace, a new small-size 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 reverberatory 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 concentrate. 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(furnace is very low in particulates. The exhaust
Mgas is produced continuously and has an SOg 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.
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 classifications of hydro-
metallurgy:
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 pyrometallurgical
steps.
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Simple Leaching - Simple heap or vat leaching (Section 2, Process Nos.
31 and 32) as practiced in the western states in an accelerated form of natural
weathering. 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 (17). 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 sulfate, has
been tested in Michigan and Nevada (18). However, if uncontrolled, these
processes could cause much wider pollution of ground water than surface
operations.
Thin Layer Leaching (19) - A thin layer (TL) leaching technique has been
developed by Holmes and Narver, Inc., of Anaheim, California. The method is
claimed to be more efficient that other types of leaching for many oxide and
mixed ores.
Process Description - Ore is crushed to less than about 1 centimeter and
is mixed with a small amount of acid. A unique curing step follows in which
the ore is placed in a curing pile for up to 24 hours. During this period,
the acid solubilizes the metal values and generates heat which accelerates
the leach reaction. The cured ore is than spread over large concrete pads in
a layer 1 meter thick or less. A leaching solution is percolated through
this layer to remove metal values. The initial leach effluent has high metal
values and low acid content. This liquor is treated by solvent extraction;
the highly acidic solvent extraction raffinate is recirculated as a leaching
solution.
Environmental Residuals - Exhausted beds should be nearly dry when
dumped in landfills; this should result in less groundwater pollution than
other leaching processes.
Status of Technology - A commercial facility is being constructed in
Chile which will handle 2300 metric tons per day of copper ore.
Hybrid Systems - In hybrid hydrometallurgical processes, roasting of
the ore concentrate is generally followed by leaching and electrolysis. The
"sulfation roasting" used at one U.S. cooper plant (Section 2, Process No.
36) is an example of a hybrid process, as is the electrolytic process in
use at four 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.
Treadwell (20) - 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.
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Process Description - Copper concentrate is leached in a kiln with
sulfuric acid, forming soluble copper sulfate which is separated from in-
solubles 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.
Minemet (21,22) - A process developed by Minemet Recherche of France is
said to be nonpolluting, economically competitive with pyrometallurgical
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 fa-
vorable 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 extraction conditions. The cupric chloride leach-
ing agent is regenerated in both of these processes. The solvent is then
stripped with spent electrolyte, producing copper sulfate. Copper is re-
covered by conventional electrolysis.
Energy Requirements - Energy use is said to be low. Only electricity
and air are used, and the extraction and electrolysis stages are conducted at
about 50°C.
Environmental Residuals - If it does not receive further processing, the
leach residue could prove to be an unstable solid waste. However, the ele-
mental sulfur it contains could be easily recovered by flotation or other
means for internal use or sale. After copper removal, the electrolyte solu-
tion 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 - Minemet has been studying this process on a
laboratory scale. Construction of a pilot plant is planned for 1977.
Fused-Salt Electrolysis (23,24,25,26) - A 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 kilograms per square
centimeter 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 solution by evaporation.
To form a stable low-melting electrolyte, potassium and lithium chlorides are
mixed with the lead and zinc chlorides, which are then electrolyzed indi-
vidually in sealed cells with graphite electrodes. Electrolysis is conducted
at 500°C, and the electrodes operate at a current density of 0.8 ampere per
square centimeter. 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 replacement. 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 kilocal-
ories per kilogram for lead production and 455 kilocalories per kilogram for
zinc. Current efficiency is 95 percent.
Environmental Residuals - The disposition of impurity metals in these
fused-salt electrolytic processes has not been described. However, they must
be discarded as wastes since the purity of the lead or zinc product is ex-
pected 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.
Battelle (27) - A hybrid hydrometallurgical process using a different
approach has been developed by Battelle-Pacific Northwest Laboratories.
Process Description - Following a neutral roast of copper concentrate,
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 concentrations of copper. The
residue is processed to copper by conventional pyrometallurgical methods.
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Environmental Residuals - S0£ from the pyrometallurgical operations can
be reacted with the hydrogen sulfide from leaching to form elemental sulfur.
Careful process control could reduce 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 competitive
with conventional technology.
Status of Technology - Laboratory-scale tests are being conducted.
Lime-Concentrate-Pen et Roast (28) - Another modification of sulfation
roasting which has not been as thoroughly studied as the Minemet is lime-
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 environ-
mental impacts to those of conventional sulfation roasting.
Status of Technology - Research has been limited to laboratory tests.
Nitrogen Roast (29) - Another hybrid hydrometallurgical process being
investigated by the U.S. Bureau of Mines involves roasting in an atmosphere
of nitrogen.
Process Description - Copper concentrate is roasted in a nitrogen
atmosphere, removing about 20 percent of the sulfur present. A subsequent
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
Environmental Residuals - Environmental hazards cannot be determined
based on available data.
Status of Technology - Laboratory testing only has been conducted.
Distillation with Acetonitrile (30) - A process for removing copper from
sulfati on-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 widespread
use might cause serious pollution.
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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 concen-
trates directly with chemicals, using no pyrometallurgical steps. Current
research appears to be in progress primarily in the United States and Canada.
Cymet (6,31,32,33,34) - An advanced hydrometallurgical process is being
developed with the trade name of Cymet.
Process Description - The Cymet process is expected to be based on a
leach using a solution 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 silver, 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. It is believed that following the ferric
chloride leach cuprous chloride is to be separated from the leach liquors by
crystallization. The designer of the Cymet process states that it is a
closed-loop system. One or more regeneration steps will therefore be re-
quired to recycle the hydrochloric acid and to oxidize the ferrous iron 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 reactions. Details of these procedures have not been announced.
Theoretically, 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.
Energy Requirements - Steam is probably used to maintain leach tempera-
ture and electricity for materials handling. Fluidized bed processes require
sizable amounts of electricity to circulate high-velocity gas streams.
Details of utilities have not been published.
Environmental Residuals - Possible wastes include hydrochloric acid
fumes from leaching, a liquid effluent from crystallization, and solids from
the regeneration process.
Costs - Information is not available at the present.
Status of Technology - A pilot plant was built in 1974 but was never
started. In 1975, substantial changes were made, but no information was
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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.
Anaconda Arbiter (6,11,31,35,36) - The Anaconda Company is developing a
hydrometallurgical process which has made use of both chemical and electro-
winning techniques for treatment of leach liquor.
Process Description - Copper ores are leached with ammonia and oxygen in
two parallel reactors, forming soluble copper ammonium sulfate, ferric oxide,
and hydroxide. Froth flotation mechanisms in the reactors aerate the pulp
and increase the dissolution rate. In an earlier process variation, the
pregnant liquor was next filtered, boiled to drive off excess ammonia, and
treated with SO? gas. An insoluble copper ammonium sulfite formed which
precipitated and was filtered from solution. Autoclaving in the presence of
sulfuric acid converted the crystals to metallic copper.
The more recent Arbiter process separates the leach liquor and residue
in a series of countercurrent decantation thickeners. The pregnant solution
from the first thickener is pressure filtered and contacted with an organic
reagent in a solvent extraction step. The extracted copper is washed to
remove ammonia and stripped in spent electrolyte. This copper-bearing
electrolyte then enters a conventional electrowinning stage for cathode
production.
An ammonia recovery system is included in the process stream.
. Energy Requirements - This process, especially the later version, is
energy intensive. Electrical energy is required for electrowinning, and
steam generation is necessary as well. Quantities are unreported.
Environmental Residuals - Quantities of leach residue are considerable,
1.3 to 1.8 tons per ton of copper produced. This residue contains iron
oxide, silica, pyrites, bismuth, sulfides, lead, arsenic, zinc, and other
materials. Calcium sulfate sludge is discharged from ammonia recovery at a
rate of 4.5 dry tons per ton of copper produced. Atmospheric emissions
include fugitive H2SOa and ammonia from the more recent process. Liquid
effluent^from the earlier process include an ammonium sulfate solution from
the precipitation and autoclave steps. This solution contains copper and
free sulfuric or sulfurous acid.
Further details are not available.
Status of Technology - Both variations have operated in large process
demonstration units.
Sherritt-Gordon/Cominco (10,20,31,37) - A 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.
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Environmental Residuals - As with other chloride-based hydrometallurgi-
cal 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-Kennecott (31,38) - 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 filtration, and the copper
is extracted by electrolysis. The process is pressurized, operating at
temperatures up to 200°C.
Status of Technology - A pilot plant is now under construction, and this
has the appearance of a serious and well-organized development effort.
Lurgi-Mittenburg (37) - 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 kilograms per square centimeter and temperatures of 115°C
promote direct oxidation of the minerals with oxygen gas. Copper is re-
covered by electrolysis.
Status of Technology - A demonstration plant is now operating.
Treatment of Zinc Silicate Ores (39) - An auxiliary application of
hydrometallurgy is being tested in connection with zinc production from
silicate ores by the Electrolytic Zinc Company of Australia. The problem
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 industrial 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.
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Miscellaneous Technologies
There are several metallurgical technologies that have been proposed or
that are in use outside the U.S. that fit no generalized classification.
Torco (7,40) - Copper silicates cannot be extracted by flotation method:;
from gangue rock. Torco overcomes this problem. It is a pyrometallurgical
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.
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 fluidized-bed reactors. The low-grade
copper product is mixed with the feed to a pyrometallurgical smelter.
AMAX Base Metals Research and Development Corporation has experimented
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 flotation 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 Mauri-
tania. The AMAX modification has only been demonstrated in laboratory tests.,
TBRC (6,7,10,41,42) - The TBRC (top blown rotary converter) was origi-
nally designed for nickel production. However, tests by its developer, the
International Nickel Company of Canada (INCO), have shown that it is appli-
cable to other metals. They claim that the TBRC, using U.S. concentrates,
can produce copper that can be directly cast as anodes, thus eliminating
fire refining. The TBRC is said to provide increased operational flexibility
through close control of both turbulence and temperature 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 air, oxygen, 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
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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 temperatures resulting eliminate more impurities and cause better
separation of slag from the copper than with other converters.
Energy Requirements - Because of its high thermal efficiency and elim-
ination of fire refining, the TBRC uses less fuel than other converters.
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 smelter'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, was scheduled for completion in late 1977.
Oxygen-Enriched Smelting (10,11) - 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 ca-
pacity. 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 conventional
reverberatory smelting, increasing fuel efficiency. The increased 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 improves separation of the matte and
reduces metal losses. In addition to increasing S0£ concentrations, en-
richment also increases the converter's scrap melting capability.
The specific capacity of furnaces such as flash smelters can be in-
creased 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 manufacture is
taken into account. Fuel efficiency is increased particularly 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
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no published data in this area, the higher operating temperatures may result
in an increased circulating dust load. The higher temperatures may also
result in more NOX formation, although these gases are probably absorbed in
the acid after scrubbing and treatment in an acid 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.
Sea Nodule Processing (43) - A highly specialized hybrid process in
Canada is recovering copper, nickel, and cobalt from sea nodules.
Process Description - This process features reduction in a pyrometal-
lurgical kiln followed by smelting in an electric furnace and hydrometallur-
gical 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. Its use by U.S. technologies is currently
limited by legal difficulties regarding international use of oceanic materials,
Slag Treatment
Slags from most continuous copper pyrometallurgical processes contain
sufficient quantities of the product metal to justify recovery before dis-
posal. 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. This pro-
cess has been previously discussed (Section 2, Process No. 12).
There are two other basic recovery techniques available for slag treat-
ment. At some Japanese and Canadian smelters, flotation techniques are used
with slags that have a high sulfur content (7,44,45). Flotation slag treat-
ment will also be used at the new Noranda continuous smelter in Utah,
although few details are available. This method is described in Process No.
13, Primary Copper Production. Another possibility in use outside the U.S.
is hydrometallurgical processing of slags, consisting of grinding and leach-
ing with acids to form copper solutions, from which metal is recovered by
conventional concentrating techniques or electrolysis (46). With both these
methods, however, the slags are discarded as finely ground tailings. Without:
adequate stabilization or proection, they are less stable in weathering
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and would generate,more fugitive dusts than electric furnace slags, which are
masses of good crystalline composition. In addition, crystals in hydro-
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 signifi-
cantly higher in either case. Hydrometallurgical treatment is probably the
most expensive, since it requires both grinding and electrolysis for complete
recovery of the metal.
References
1. Encyclopedia of .Chemical Technology. Interscience Publishers, a division
of John Wiley and Sons, Inc. New York. 1967.
2. Shewes, H.R. Electrowinning of Lead Directly from Galena. Proc.
Australasian Inst. Min. Met. December 1972. (244). pp. 35-41.
3. Kawakita, T., T. Kitamure, Y. Sakoh, and K. Sasaki. Design, Construc-
tion, 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. Claessen, P.L. Electrowinning of Zinc under Periodic Reverse and
Pulsating Current Conditions. Presented at the 104th annual meeting of
the AIME. New York, New York. February 16-20, 1975.
5. 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.
6. Background Information for New Source Performance Standards: Primary
Copper, Zinc, and Lead Smelters. Volume I, Proposed Standards.
EPA-450/2-74-002a. Environmental Protection Agency. Research Triangle
Park, North Carolina. 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 T. Suzuki. 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, H., et al. Environmental Considerations for Emerging Nonferrous
Metal-Winning Processes (Field Review Copy). U.S. Bureau of Mines.
Salt Lake City Metallurgy Research Center.
370
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11. Arthur D. Little, Inc. Environmental Considerations of Selected Energy
Conservation Manufacturing Process Options. Vol. 14, Primary Copper
Industry. EPA Contract No. 68-03-2198. Draft. Cambridge, Mass.
12. PEDCo Environmental, Inc. Evaluation of the Status of Pollution Control
and Process Technology - Japanese Primary Nonferrous Metals Industry.
Contract No. 68-02-1375, Task 36. U.S. Environmental Protection Agency.
July 1977.
13. Kivcet. Pamphlet by V/0 Licensintorg Mostva-SSSR. Distributed by
Southwire Company. Carroll ton, Georgia.
14. Kivcet Process for Complex Ores. World Mining. June 1974. pp. 26-27.
15. Queneaw, P.E., and R. Schuhmann, Jr. The Q-S Oxygen Process. Journal
of Metals. New York, New York. August 1974. pp. 14-16.
16. Matyas, A.G., and P.J. Mackey. Metallurgy of the Direct Smelting of
Lead. Metallurgical Society of AIME. New York, New York. TMS Paper
No. A75-80. February 1975.
17. Michalek, Z. and A. Malik. The Influence of the Mineral Composition of
Copper Ores on the Effect of Biological Leaching. Archives Hutn. 1975.
20, (3) pp. 421-432.
18. Hockings, W.A. and W.L. Freyberger. 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.
19. Teter, E.K. Personal communication on TL Process. Holmes and Narver,
Inc. Anaheim, California. 1977.
20. Chemical Route to Copper? Chemical Engineering. New York, New York.
April 20, 1970. pp. 64-66.
21. Demarthe, J.M., L. Gandon, and A. Georgeaw. A New Hydrometallurgical
Process for Copper. Presented at the 105th annual meeting of the AIME.
Las Vegas, Nevada. February 22-26, 1976.
22. 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.
23. Haver, F.P., C.H. Elges, D.L. Bixby, and M.M. Wong. Recovery of Lead
from Lead Chloride by Fused-Salt Electrolysis. U.S. Department of the
Interior. Bureau of Mines Report of Investigations 8166. Washington,
D.C. 1976.
24. Haver, F.P., and M.M. Wong. M.M. Ferrix Chloride - Brine Leaching of
Galena Concentrate. U.S. Department of the Interior. Bureau of Mines
Report of Investigations 8105. Washington, D.C. 1976.
371
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25. Haver, P.P., D.C. Shanks, D.L. Boxby, and M.M. Wong. Recovery of Zinc
from Zinc Chloride by Fused-Salt Electrolysis. U.S. Department of the
Interior. Washington, D.C. Bureau of Mines Report of Investigations
8133. 1976.
26. Bureau of Mines Research 1975. U.S. Bureau of Mines. Washington, D.C.
27. Rohrmann, C.A., and H.T. Fullam. Control of Sulfur Dioxide Emissions
from Copper Smelters - Volume II, Hydrogen Sulfide Production from
Copper Concentrates. Environmental Protection Agency, Office of Research
and Development. Research Triangle Park, North Carolina. PB 237 928.
September 1974.
28. Bartlett, R.W., and H.H. Haung. The Lime-Concentrate-Pellet Roast
Process for Treating Copper Sulfide Concentrates. Journal of Metals.
December 1973. pp. 28-34.
29. Gabler, R.C., B.W. Dunning, R.E. Brown, and W.J. Campbell. Processing
Chalcopyrite Concentrates by a Nitrogen Roast-Hydrometallurgical Tech-
nique. U.S. Department of the Interior. Bureau of Mines Report of
Investigations No. 867. Washington, D.C. 1975.
30. Propose Copper Leaching Process. Industrial Research. August 1976.
pp. 38.
31. Rosenzweig, M.D. Copper Makers Look to Sulfide Hydrometallurgy.
Chemical Engineering. Volume 83, No. 1. January 5, 1976. pp. 79-81.
32. 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.
33. Druesi, P.R. Cymet Copper Reduction Process.
34. Copper Hydrometallurgy: The Third Generation Plants. Engineering and
Mining Journal. June 1975.
35. Arbiter, N. and D. Milligan. 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.
36. Arbiter, N., D. Milligan, and R. McClincy. Metal Production from
Copper Ammine Solution with Sulfur Dioxide. Presented at IChemE Inter-
national Symposium on Hydrometallurgy. Manchester, England. 1975.
37. Habashi, F. Pressure Hydrometallurty: Key to Better and Nonpolluting
Processes. Engineering and Mining Journal. New York, New York. May
1971. pp. 88-94.
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38. Brenneck, H.M. Recovery of Metal Values from Ore Concentrates. U.S.
Patent No. 3,888,748. June 10, 1975.
39. Matthew, I.G., and D. Eisner. The Hydrometallurgical Treatment of Zinc
Silicate Ores. Metallurgical Society of AIME. TMS Paper No. A75-74.
New York, New York. February 1975.
40. Opie, W.R., L.D. Coffin, and D.C. Casanelli. 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.
41. Daniele, R.A., and L.H. Jaquary. TBRC A New Smelting Technique. Dravo
Corporation, Pittsburgh, Pennsylvania. Also presented at 1972 annual
meeting of the AIME. February 20-24, 1972.
42. Mason, J.L., and S.F. Siscoe. Afton Mines Smelter to Use Top Blown
Rotary Converter Process. The Northern Miner. April 22, 1976. pp.
B12-B13.
43. Sridhar, R., and W.E. Jones. 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.
44. Hallett, G.D. Continuous Copper-Smelting Process Uses Single Vessel.
Chemical Engineering. New York, New York. April 26, 1976. pp. 62-64.
45. Mills, L.A., G.D. Hallett, and C.J. Newman. Design and Operation of
the Noranda Process Continuous Smelter. Presented at the International
Symposium on Copper Extraction and Refining. 105th annual meeting of
AIME. Las Vegas, Nevada. February 22-26, 1976.
46. Kmetova, D. Hydrometallurgical Method of Copper Recovery from Slags.
Zbornik V.S.T. Kosiciach. 1971.
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SECTION 9
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 consists of
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 enter 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 con-
taining 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.
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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 many 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 S0£ 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 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 $63 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 materials determined
in (a).
(c) Demonstrate disposal technique selected.
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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 effec-
tiveness 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.
Determination of slag cleaning furnace exhaust stream characteristics
and emission control methods.
Need:
Electric slag cleaning furnaces are being installed in conjunction with new
pyrometallurgical 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
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the development of emission control technology should be determined.
Project Scope:
(a) Determine exhaust gas flows and pollutant exhaust concentrations,
including suspected heavy metals, carbon monoxide, and metal
carbonyls.
(b) Determine types of pollution control systems that are effective to
control the emissions that were characterized in (a).
(c) Establish emissions performance standards for slag cleaning fur-
naces.
Further evaluation of the use of sulfides to precipitate heavy metals from
wastewaters.
Need:
Sulfides will be used in Sweden 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 necessary to
remove metallic species from wastewaters.
(b) Determine heavy metal removal effectiveness and secondary resolution
of precipitated metals.
(c) Evaluate potential of H^S formation from treated waters.
(d) Determine disposal techniques for metal sulfide sludges that
result from the treatment process.
Study the problems of arsine evolution from tank houses.
Need:
Arsine 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.
Project Scope:
(a) Measure arsine concentrations at various suspect points in the
refinery operation.
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(b) Develop means of extracting arsenic before the electrolytic refine-
ment stage.
(c) Investigate transport of arsine in the ecosystem.
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 $03 control requirements tend to
support an expansion of hydrometallurgical methods for the recovery of
copper. The disposal of increased quantities of resulting spent leach solu-
tions 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.
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 at the smelter will be considerably decreased.
Project Scope:
(a) Investigate new leaching procedures for preleaching of the con-
centrate.
(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.
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Assessment of the environmental effects from tailings containing elemental
sulfur, and determination of methods for their control.
Need:
Most of the developing hydrometallurgical processes for copper production
will produce large quantities of elemental sulfur, which will be discarded
along with waste rock as 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 could make hydrometallurgy more economically
attractive and reduce the environmental problems of sludge disposal.
Project Scope:
(a) Phase-identify residual solids from conventional hydrometallurgical
processes.
(b) Investigate possible conversion of leach residue into concentrate
for iron production.
(c) Consider alternative solvents for recovery of copper and iron in a
desirable form.
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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 condition.
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 effects associated with discharges
determined in (a).
(c) Determine control system requirements and costs to reduce effluent
concentrations to acceptable 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 alkalinity of
surface waters in the Missouri area is reported to inhibit dissolution of
heavy metals, but possible long-term effects 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.
Investigation of economics for hydrometallurgical production of lead.
Need:
Hydrometallurgical production of lead has been suggested for environmental
reasons, but economics are still not sound. Considerable development will be
required to make the process competitive.
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Project Scope:
(a) Conduct research to determine more favorable kinetics for lead
chlorination hydrometallurgy.
(b) Investigate the possibility of bacterial leaching; some geochemists
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 sul-
furic acid plant. The weak gas $62 concentration is too low for acid manu-
facture, 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 SO? con-
centrations, 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.
Lead blast furnace SO,, recovery.
Need:
S02 emissions from a lead blast furnace account for only about 5 to 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.
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Metal and trace material emission control efficiency - baghouses, electro-
static precipitators, 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. Emissions of zinc, arsenic, and
cadmium, as well as other substances, should be determined.
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.
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 operating
lead smelters.
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(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.
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) Investigate 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 flotation separation. Other chemicals
may offer an alternative once this operation is better understood.
Project Scope:
(a) Conduct research on the surface kinetics and flotation chemistry
required for better understanding of the phenomenon.
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(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 properties.
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 wastewater
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 treatment.
(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 requirement for deposition of
these wastes.
Project Scope:
(a) Establish conceptually a method for mechanically separating mill
tailings from flotation water.
(b) Develop detailed conceptual drawings for the device.
(c) Define the manufacturing and operating costs.
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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 effec-
tively 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 dis-
charge to a public watercourse. Reacidification may redissolye some of the
precipitated heavy metals, especially if mineral acids are being used. A
study of the methods and necessary degree of reacidification would be valu-
able.
Project Scope:
(a) Perform laboratory tests to establish optimum pH for removal of
heavy metals from typical base metal wastewater streams, by treat-
ment 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.
Investigation of chelating 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
385
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solvent, thus concentrating the heavy metal. Chelating chemicals are in use
that are specific to copper and are used in solvent extraction techniques for
copper recovery. It is possible that other less-specific chemicals or chemi-
cal mixtures could be effective in removing many 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 com-
pounds 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 heavy
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 heavy metals, the water must be completely
softened to ensure complete removal of these elements.
It has been postulated that a resin based on magnesium would be insensitive
to the common heavy 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 exchanging
substance specific to heavy metals in solution.
(b) Synthesize these materials and define their operating suitability.
Investigation of selenium-accumulating plants as a method of controlling
selenium in base-metal production wastes.
Need:
There is no proven process for extracting selenium from mine spoil, concen-
trator 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.
386
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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 artificially duplicating this mechanism.
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