EPA 670/2-75-018
APriii975 Environmental Protection Technology Series
RECLAMATION OF
METAL VALUES FROM
METAL-FINISHING
WASTE TREATMENT SLUDGES
National Environmental Research Center
Office of Research and Development
U.S. Environmental Protection Agency
Cincinnati, Ohio 45268
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EPA-670/2-75-018
April 1975
RECLAMATION OF METAL VALUES FROM
METAL-FINISHIHG WASTE TREATMENT SLUDGES
By
Arch B. Tripler, Jr.
R. H. Cherry, Jr.
G. Ray Smithson, Jr.
Battelle-Columbus Laboratories
Columbus, Ohio 1+3201
Program Element No. 1BB036
Project Officer
Eugene F. Harris
Mining Pollution Control Branch (Cincinnati, Ohio)
Industrial Waste Treatment Research Laboratory
Edison, New Jersey 08817
NATIONAL ENVIRONMENTAL RESEARCH CENTER
OFFICE OF RESEARCH AND DEVELOPMENT
U.S. ENVIRONMENTAL PROTECTION AGENCY
CINCINNATI, OHIO U5268
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REVIEW NOTICE
The National Environmental Research Center - Cincinnati has reviewed
this report and approved its publication. Approval does not signify
that the contents necessarily reflect the views and policies of the
U. S. Environmental Protection Agency, nor does mention of trade
names or commercial products constitute endorsement or recommen-
dation for use.
11
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FOREWORD
Man and his environment must be protected from the adverse effects
of pesticides, radiation, noise and other forms of pollution, and
the unwise management of solid waste. Efforts to protect the en-
vironment require a focus that recognizes the interplay "between
the components of our physical environment—air, water, and land.
The National Environmental Research Centers provide this multi-
disciplinary focus through programs engaged in
• studies on the effects of environmental contaminants
on man and the biosphere, and
t a search for ways to prevent contamination and to
recycle valuable resources.
As part of these activities, efforts of this study have included
the determination of the worth of recovering metal values from
metal finishers1 wastewater treatment sludges, the definition
and research of processes for such recovery, and the selection,
design, and costing of a recommended process.
A. W. Breidenbach, Ph.D.
Director
National Environmental
Research Center, Cincinnati
iii
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ABSTRACT
The efforts of this program have included the determination of the
worth of recovering metal values from metal finisher's wastewater
treatment sludges, the definition and research of processes for such
recovery, and the selection, design, and costing of a recommended
process.
The study included a survey of the literature to determine the state
of the art regarding the generation, disposal, and recovery treatment
practices relevant, to metal-finishers' sludges, and to identify metal
recovery processes possibly applicable to those sludges. This infor-
mation was supplemented with a survey by questionnaire to determine
the current status of relevant practices and conditions.
Field investigations provided detailed examples of plant practices,
sludge storage conditions, and sludge characteristics such as water
content and chemical composition. Further, a detailed study was
performed to determine, for one lagoon, the degree of leaching of
metals from the sludge into the underlying soil structure. Samples
of sludge were obtained from operating plants for use in the experi-
mental program.
The extraction of metal values from waste sludges by various leaching
agents, and the recovery of metal values by techniques including
electrowinning, cementation, and liquid-liquid ion exchange were
studied on a bench scale during the program.
On the basis of the information developed in the prior portions of
the program, a portable pilot process for the treatment of waste
sludges and recovery of metal values was selected and equipment and
operating costs developed.
This report was prepared by Battelie for the Metal Finishers' Foundation
and was sponsored by Grant No. 12010 FXD with the U. S. Environmental
Protection Agency.
IV
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CONTENTS
Page
Abstract iv
List of Figures vi
List of Tables vii
Sections
I Conclusions 1
II Recommendations 4
III Introduction 6
IV Background Information 8
V Objectives of the Research Program 10
VI Acquisition of Existing Information 11
VII Field Studies 19
VIII Laboratory Studies 28
IX Recovery Experiments 31
X Economic Studies 4°
XI Appendixes ^2
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FIGURES
Mo. Page
1 Schematic Drawing of 3-Compartment Cell 33
2 Current Efficiency as a Function of Time 35
3 Schematic Drawing for Appartus Used 37
for Constant Cathode Potential Electrowinning
C-l Schematic Drawing of Location of Bore-Holes, 57
Results of Gross Examination of Cores, and
Percentage Recovery of Cores
D-l Flow Diagram for Proposed Method for 66
Separating Copper and Nickel Using
Naphthenic Acid
D-2 Generalized Flowchart of Liquid-Liquid 69
Extraction and Recovery Process
D-3 Run 7 Flow Chart 71
E-l Flowsheet of Proposed Sludge Treatment 80
Pilot Process
VI
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TABLES
1 Tabulation of Questionnaire Returns 14
2 Reported Metal Contents of Sludges 15
3 Associations of Plating Metals in Sludge 17
Reported
4 Treatment and Disposal Practices 20
5 Sludge Characteristics 21
6 Semi-Quantitative Spectroscopic Analysis 22
of Dried Sludge
7 Semi-Quantitative Spectroscopic Analysis 23
of Dried Sludge
8 Semi-Quantitative Spectrographic Analyses of 24
Three Samples of Dried Sludge
9 Semi-Quantitative Spectrographic Analyses of 25
Dried Sludge Samples from Company G
10 Selected Leaching Experiments on Dried 29
Sludges
C-l Semi-Quantitative Spectrographic Analysis 58
of Selected Sludge and Core-Drilling Samples
D-l Selected Leaching Experiments 60-6l
D-2 Materials Used in Liquid-Liquid Extraction 68
Runs 7 through 14
D-3 Stripping Liquor Analyses 72
D-4 Rough Quantitative Comparison of Enriched 7^-76
Stripping Liquors
E-l Equipment Cost Summary 82
E-2 Operating Cost Summary 83
vn
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SECTION I
CONCLUSIONS
On the basis of information gathered in the literature survey, it is
concluded that the metal-finishing industry in the United States
rarely treats sludges to recover plating metals, either for economic
gain or for further detoxification of the sludge. This finding is
borne out by the results of the plant survey by questionnaire.
The quantity, distribution, and physical and chemical properties of
metal-finishing sludges are believed to account for this lack of
activity in metal recovery and further detoxification.
There are undoubtedly some specific instances, none of which came to
light in this study, where sludges could be profitably worked for
metal recovery and detoxification. Such instances would have to be
uncovered and evaluated by pointed studies. It must be emphasized
that, in general, there is no discrete "sludge treatment" process
available.
Two general approaches to the recovery of metals and the detoxification
of sludges, which were investigated in a preliminary manner in this
study, are believed to be the most promising routes to follow in the
design and development of a given metal-finishing sludge treatment
process. There are:
(a) Sulfuric acid dissolution to dissolve metal
hydroxides and precipitate calcium sulfate,
followed by filtration to obtain a clear
pregnant liquor.
(b) Solution in ammonia-ammonium carbonate
solution followed by filtration to produce
a pregnant liquor.
The details of recovery processes based on these two approaches could
be settled upon only after full information was available on the
quantity and composition of a sludge under consideration.
On the basis of data developed during this study, it would be uneconomic
to apply the sulfuric acid leaching approach to mixed sludges (con-
taining nickel, copper, and chromium) of typical composition (3 to 57,
Ni, 3 to 5% Cu, 5 to 15% Cr, on a dry solids basis) unless the process
could be carried out on a large, and preferably continuous scale.
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Preliminary cost estimates on a small batch operation capable of
handling only 100 Ib per day of dried sludge indicated a daily
operating cost of about $85.00 with but a few dollars realizable from
recovered metals. It is estimated that the minimal feed rate would
have to be in the range of 2,000 to 5,000 Ib of sludge (dry basis)
solids per day even with a thoroughly developed process.
Although no economic studies were made on the ammonium carbonate
leaching approach, it is judged that this type of process would be
applicable only on an even larger scale, owing to the cost of reagent
and the necessity for incorporating ammonia recovery system in the
plant.
In the design of any process based on either of these two approaches,
electrolysis or electrowinning of such metals as copper, nickel, and
cadmium should receive prime consideration, because its tools and
technology are shared with metal-finishing. The laboratory work
demonstrated the general feasibility of this approach.
Chromium and zinc in mixed metal-finishing sludges are dubious assets,
if not liabilities, because of their low unit price and the compli-
cations they can introduce to chemical processing. The technical and
economic feasibility of treating segregated sludges with chromium and
zinc as single base-metal components would have to be judged on a
case-by-case basis.
Although no work was done to examine the feasibility of recovering
tin and precious metal reported in some of these slugdes, their
extraction and recovery would require additional and possibly exotic
processes.
Liquid ion exchange as a process for separating metals in mixed sludges
was examined at length in this study. It does not appear to offer much
hope as an economic approach to the problem.
In the single case investigated in this study, it was determined by a
core-drilling and analytical program that a sludge lagoon in service
about 12 years had not contaminated ground water underlying the lagoon
bed.
It was proposed originally that fluidized-bed drying and/or calcination
might be an attractive route for dewatering the sludges so that they
could be shipped a reasonable distance to a centralized recovery plant.
Following the survey of plating plants, it was concluded that this
approach generally would not be feasible. The data indicated that the
quantities of sludge produced at most individual plant sites were too
small to warrant the expense of installing and operating a fluidized-bed
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system. For those areas in which a relatively heavy concentration
of plating plants exists, the installation of a centrally-located wet
processing facility to process filtered sludge is believed to be
more practical.
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SECTION II
RECOMMENDATIONS
In general, the recovery of metals by treating mixed sludges that
currently are being produced by metal-finishing plants or that have
been stored by chemical processes is not economically attractive.
This is because most such sludges are too low in grade in metal values,
are complex mixtures that require difficult and expensive separations,
and are variable in quantity, composition, and consistency. However,
increasing environmental restrictions may preclude the use of the
landfill option in some areas and make other techniques, e.g., segre-
gation of waste streams and metal recovery, necessary.
Currently, the only practical way to treat most mixed metal-finishing
sludges now being produced is to landfill them, preferably after
dewatering and/or solidification.
There may be a few instances in the metal-finishing industry where the
quantity and grade of the sludges being produced or those stored may
be great enough to warrant complex chemical processing at a centralized
treatment facility. A separate study aimed at uncovering such
instances should be made. The specific aims of this study should
parallel those of the mining prospector and developer: (1) to locate
likely sites (such as large sludge deposits at a single facility or
clusters of sludge deposits from a number of facilities within which a
centrally located treatment plant might be established, (2) to deter-
mine the quantity and grade of sludge available for treatment in the
given area, (3) to devise a process that could be applied to the
treatment of such sludges, and (4) to work out preliminary economic
estimates on the processes.
Metal finishers should be encouraged to keep sludges from various
plating lines separate to enhance the possibility that simple, straight-
forward, and hence, inexpensive, recovery processes might be used.
Advantageous metal associations would be copper alone, nickel alone,
cadmium alone, tin alone, precious metals alone, copper plus nickel,
copper plus nickel plus precious metals. Chromium should, where
possible, be kept out of mixed sludges because in the form in which it
is readily recoverable, it has little value because it complicates and
adds expense to separation processes, and because when eventually
separated, its ultimate destination would very probably still be a
landfill. Similarly, zinc, where possible, should be excluded from
mixed sludges, even cadmium sludges, because it, too, is a low value
material that complicates chemical separations.
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If segregation of metal values in sludges can be achieved, the oppor-
tunities for metal recovery probably would increase enormously, even
for metals like chromium and zinc. In such cases, simple filtering/
drying could conceivably produce a product that the plater could store
in a low volume and dispose of to refineries or centrally located
disposal plants. It is recommended that the problem of sludge disposal
and detoxification be reexamined from the specific viewpoint of what
could be done if proper segregation of plating sludges were practiced.
A study of this nature would involve an assessment of foreign practice
(England, Japan, etc.) where some form of segregation and recovery is
practiced. It also would include a survey of sludge production in the
U.S. to determine the major areas in which sludges are produced and
the nature of these materials. The overall objective would be to
determine the feasibility of centralized processing facilities for
treating segregated sludges.
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SECTION III
INTRODUCTION
Many of the potential hazards accompanying the unrestricted disposal of
wastes from biological and industrial processes have been recognized
for several hundred years. Only in recent years, however, have the
complexity and intensity of the problems involved in waste management
and disposal reached the point where the quality of the air-land-water
environment is critically impaired in many geographical areas. Increas-
ingly restrictive legislation has been passed as efforts to curb
pollution have been intensified.
Industrial wastes have contributed heavily to the degradation of the
quality of man's environment and many industries have invested consid-
erable sums of money in efforts to abate pollution. The nature of the
waterborne wastes produced in metal-finishing operations is such that
complex and costly methods have been needed to prevent the contamin-
ation of water supplies with the compounds used by this segment of the
industrial world.
The specific steps involved in the treatment of these wastes vary
considerably from installation to installation. However, the general
treatment consists of: (1) separation of grease and oil; (2) reduction
of chrornates; (3) oxidative destruction of cyanides; (4) neutralization
and liquid-solid separation; and (5) disposal of the sludge. The metals
initially present in the wastewaters are precipitated as hydroxides
during treatment operation number (4).
The metal-bearing sludges are difficult to handle and their final
disposal often is the most troublesome and costly aspect of the treat-
ment of metal-finishing wastes. Increasing congestion of urban areas
in which many of the larger plating facilities are located together with
the accompanying increases in the cost of land add to the problems
involved in ultimate disposal. The finely divided precipitates contained
in these sludges are a potential source for pollution of surface and
ground waters.
In addition to the problem of handling and disposal of these sludges,
they contain large quantities of relatively valuable metals--copper,
nickel, chromium, cadmium, etc.—which often are in short supply. The
value of the nickel and chromium lost in plating wastes in the United
States has been estimated at 25 to 30 million dollars per year.
Sludges selected from various steps in the waste treatment procedure
may contain 10 to 15 percent of nickel, chromium, etc., on a dry basis--
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concentrations greater than those found in most ores containing these
metals. The extraction and recovery of the more valuable metals from
the sludges should ideally involve relatively simple techniques. The
economic feasibility of recovery would be largely dependent on the
size of the plating facility, the cost of the primary raw materials
used in the plating operations, and the cost of alternative methods of
sludge disposal, e.g., lagoons, landfills, etc.
Because of the importance of eliminating potential water pollution
problems, conserving natural resources, and recovering valuable
materials, an investigation was undertaken of the possible recovery of
metals from sludges produced during the treatment of metal-finishing
wastes. Because this is a general problem in the metal-finishing
industry, the investigation was supported by a grant from the U. S.
Environmental Protection Agency to the Metal Finishers' Foundation.
This allowed Battelle to work closely with an important segment of the
industry in seeking a solution to this problem. The National Asso-
ciation of Metal Finishers has approximately 100 members who have waste
treatment facilities in which sludges are produced.
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SECTION IV
BACKGROUND INFORMATION
There are between 15 and 20 thousand metal-finishing facilities in the
United States. These include several thousand small shops, a number of
large independent shops and many captive metal finishing facilities
associated with industries such as automobiles, aircraft, and home
appliances. Although the composition and volume of the wastewaters
from these shops vary considerably from plant to plant, they include
in general the following:
(1) Rinse waters from plating, cleaning, and other
surface finishing operations.
(2) Concentrated plating and finishing baths that
are intentionally or accidentally discharged.
(3) Wastes from plant or equipment cleanup.
(4) Sludges, filter cakes, etc., produced by
naturally occurring deposition in operating
baths or by intentional precipitation in the
purification of operating baths, chemical
rinsing circuits, etc., when flushed down sewers.
(5) Regenerants from ion exchange units.
(6) Vent scrubber waters.
The volumes as well as the compositions of the sludges produced in the
treatment of metal-finishing waste effluents are widely varied. Among
the variables which influence the volume and composition are: (1) the
characteristics of the finishing effluent treated, and (2) the treatment
method employed.
Although some companies have considered the recovery of metals from
metal-finishing waste sludges, the additional processing required
generally has not been considered seriously because of the relative
simplicity of current disposal methods and the lack of economic
incentive. Increasing restrictions governing the disposal of solid
wastes together with the increasing cost of land in urban areas for
disposal purposes are among the factors which make a study of additional
processing of sludges of this nature desirable. In addition to advan-
tages which may accrue from effecting reduced volumes of sludges by
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further processing, the potential recovery of valuable metals and the
reduced cost of transporting the sludges to disposal sites are further
incentives for the consideration of such treatment. Some of the larger
metal-finishing installations each year discard sludges containing
metals valued at $150,000 to $200,000 and often pay in excess of
$50,000 for hauling these sludges to disposal sites. If a simple,
technically feasible technique can be developed to recover the more
valuable metals and to effect a substantial reduction in the volume
of sludge to be transported, the value of the metals recovered and the
anticipated reduction in hauling costs may be more than sufficient to
offset the additional processing costs. This would provide an economic
incentive for such processing which, in turn, would help to alleviate
the problem of sludge disposal and pollution of valuable land areas
with solid wastes.
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SECTION V
OBJECTIVES OF THE RESEARCH PROGRAM
Battelle's preliminary surveys had indicated that the removal of metals
from sludges may simplify their disposal and diminish their potential
for pollution by leaching into ground and surface waters. These surveys
also suggested that the reclamation of valuable metals (nickel, copper,
cadmium, etc.) from sludges produced in waste-treatment plants might
significantly reduce the net cost of sludge disposal.
This study was made in some depth to evaluate these possibilities.
Specific objectives of the study were:
(1) To acquire further information on the quantity
and chemical and physical characteristics of
metal finishing sludges being produced in the
United States by plants of various sizes and
types;
(2) To investigate and evaluate techniques that
might be applicable to the treatment of metal
finishing sludges and to the recovery of metal
values from them, basing the investigative and
evaluative work on information gathered from
the literature plant surveys;
(3) To conduct bench-scale laboratory work on
actual sludges using specific approaches
suggested by the literature studies, or other
considerations, to have the most promise;
(4) To make preliminary economic estimates of the
capital and operating costs that might be
encountered in such treatment and/or recovery
processes;
(5) To investigate on a necessarily limited scale,
the effects of weathering on stored sludge
beds, to provide some data as to the permeability
of leachates from these beds into ground water;
(6) To formulate specific recommendations based on
the findings of this study, as to the direction
that future work on the sludge disposal and
treatment problem should take.
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SECTION VI
ACQUISITION OF EXISTING INFORMATION
Literature Review
The literature search performed as part of this program served to
gather information relative to all phases of the work, including both
industrial practices and trends, the state of the art in electroplater's
sludge generation, disposal, and recovery processes, and technical
details of the unit processes selected for experimental use. An
extensive literature review is given in Appendix A of this report,
providing more detailed descriptions of the content of various pub-
lished references and guidance for any subsequent follow up to this
wbrk.
In summary, the literature survey established the following bases or
trends relevant to this study:
(1) That, currently, in the Utiited States, only a
few instances of metal recovery from metal
finishing wastes are being applied on a routine,
commercial-scale basis.
(2) That the practices of recovery of metal values
from such wastes is slightly mor^ advanced in
England than in the United States, and that,
there, the economics of recovery processes are
dictating a trend toward segregation of waste
streams to produce single-metal treatment sludges.
(3) That access to lagoon storage and landfill
disposal is decreasing due to factors of decreased
availability of space or more stringent anti-
pollution regulations.
On the basis of the literature survey, observations relating to metal-
recovery processes included the following:
(1) That current metal reclamation processes are
tailored to specific plants and generally reclaim
only one metal.
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(2) That the few instances of the practice of
sludge dewatering (by vacuum-filtration,
natural evaporation, or incineration) are
applied principally to achieve decreased
sludge volumes, i.e., to extend storage
capacity. Concurrent oxidation during
incineration destroys or stabilizes cyanides
but solubilizes chromium.
(3) That the unit processes currently applied or
studied for recovery of metal values include:
(a) cementation (to recover copper)
(b) a portable copper electrowinning
unit
(c) electrowinning of zinc (with
variable success)
(d) electrolytic regeneration of chromic
acid
(e) constant cathode potential electrolytic
recovery of copper or nickel
(f) precipitation of iron or nickel as
hydroxides.
The balance of the literature reviewed in Appendix A contributed details
of these major points or were consulted for specifics of experimental
process conditions.
Survey of Plants
Six hundred and sixty-six questionnaires (see Appendix B, "Details of
Plant Survey - Questionnaire") were mailed to members of the NAMF,
including a few foreign members. One hundred and fifty replies were
received for a percentage of 22.5. This is considered an average
response. In some cases the respondents did not return a filled-in
questionnaire but, instead, wrote letters explaining their special
position in the metal-finishing field. Each of these letters was
counted as a response even though in most of these cases not all of the
desired information was given.
As shown in Table 1, slightly more than one-half the respondents are
collecting various amounts of waste slude (Item 1). Slightly less than
half of those companies accumulating sludge (about 25 percent of the
total number responding) employ a special treatment operation to
12
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precipitate the sludge (Item 2). The others who accumulate sludge do
so as a result of what might be termsed "accidental" neutralization
where alkaline and acid wastewaters are mixed (Item 3). Fifty of the
respondents (33 percent) who do not now have treatment facilities are
planning installations in the near future (Item 4). Probably others
will be following because 103 respondents (69 percent) reported activity
on the part of various governmental authorities (Item 5).
Sizable amounts of sludge appear to be available for the reclamation
of metal values. An extrapolation of the information obtained suggests
that the 666 members of MFF may be accumulated an estimated 16,000 tons
of sludge annually. Reported sludge volumes ranged from 20 to 5,000,000
Ib per year on a wet basis. Only 5 respondents gave data which per-
mitted calculation of weight on a dry basis. Going on the assumption
that the average composition of the sludges from these 46 companies
was 10 percent solids, it may be calculated that these 46 companies,
in aggregate, produced 1,130 tons of dry sludge. If this average were
preserved throughout the entire field of 20,000 metal finishing
companies, annual production of sludge on a dry basis would range
between 400,000 and 500,000 tons.
Significantly, twelve percent of the respondents have apparently worked
out methods of selling their sludge or reclaiming metal values to
achieve an economic advantage from the problem of sludge disposal.
These companies, reporting that they practiced metal recovery, were
further queried by letter to determine methods used. Response was
poor with only one U.S. and one foreign company answering. The U.S.
firm used ion exchange to recover precious metals from rinse waters
and cementation on iron to recover copper from spent circuit board
etching solution. The foreign firm indicated it segregated nickel and
cadmium rinse waters. Nickel waste was precipitated, thickened, and
filtered by vacuum. The company reported it produced about 440 Ib
of filter cake (presumably dry) containing about 40 percent Ni. This
was sold to a refiner. Cadmium recovery at this foreign firm, now
being installed, will consist of electrowinning cadmium from a dragout
solution onto steel cathodes which will be recycled as anodes to the
cadmium circuit. The company claims the economics for electrowinning
and recycling cadmium in this manner are favorable and that cadmium
pollution is significantly reduced. The foreign company has been
unable to interest metal refiners in mixed sludges, however, claiming
that refiners object to sludges containing zinc and iron.
The data given in Table 2 indicate the metals found in sludges and, to
some extent, the frequency at which they may be expected to be encoun
tered, i.e., chromium, copper, and nickel most frequently, zinc and
cadmium less frequently, and the other metals more or less rarely.
13
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Table 1. TABULATION OF QUESTIONNAIRE RETURNS
Number of
Questions Companies
(1)
(2)
(3)
(4)
(5)
(6)
(7)
(8)
(9)
(10)
Accumulate waste sludge
Controlled neutralization
Uncontrolled neutralization
Planning to install a
treatment plant soon
Activity by local, state,
or Federal agencies
Companies reporting annual
sludge volume
Method of sludge disposal
Public landfill or dump
Private landfill or dump
Sell for reclamation
Pay to haul away
Reclaim metal values
(See Table 2)
Cooperate with MFF
77
37
35
50
103
46
88
32
45
8
3
11
106
Percent of
Total (150)
Responses
51
25
23
33
69
31
58
21
30
5
2
7
71
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Table 2. REPORTED METAL CONTENTS OF SLUDES
Kind of Metal
Potentially Valuable
Base Metals
Copper
Nickel
Chromium
Zinc
Cadmium
Tin
Precious Metals
Silver
Gold
Rhodium
Unspecified
Other
Iron
Aluminum
Lead
Unspecified
Number of
Companies
78
80
88
58
63
36
16
10
1
2
8
10
2
Percent of
Total (150)
Responses
52
53
59
39
42
24
11
7
1
1
5
7
4
1
15
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The potentially valuable base metals of greatest interest to this
study are listed in the first division of the table. The recovery of
precious metals (second division of the table) from sludges constitutes
a special problem and one that is probably being adequately handled by
precious metal platers through collection and processing of sludges,
either in-house or by custom refiners.
The recovery of iron and aluminum from metal finishing sludges is, of
course, unthinkable. Lead recovery would be justified only under a
very special set of conditions.
Of the potentially valuable metals it appears that the greatest poten-
tial for economic recovery is with copper, nickel, cadmium, and tin.
Only under very special circumstances would chromium and zinc recovery
be economically worthwhile.
Table 3 lists the data obtained in the questionnaire broken down by
combinations of metals in the sludges.
These data indicate that most sludges produced up to now by metal-
finishing companies are complex mixtures resulting from the combination
of wastes from a number of plating lines. By percentage the distri-
bution is as shown in the following tabulation:
Number of "Plating" Metals Percent of Total
in Sludges Reported Reporting Cumulative
1 11.5 11.5
2 16.8 28.3
3 20.4 48.7
4 18.6 67.3
5 8.0 75.3
6 15.0 90.3
7 9.7 100.0
Of the metals cited, only nickel, copper, cadmium, tin, and the precious
metals are of real economic significance, because of their price as
well as the marketability and possible direct utilization of products
from which they might be recovered. On a similar basis, zinc, chromium,
and lead would have no economic attractiveness and very probably repre-
sent, in the majority of cases, economic liabilities in that they
complicate chemical processing to recover the valuable metals. There
16
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Table 3. ASSOCIATIONS OF PLATING METALS IN SLUDGE REPORTED
Single Metal
Ni
Cr
Cd
Pm
Two
Ni;
Cu;
Cu;
Cr;
Cf;
Zn;
Cd;
Metals
Cr
Cr
Zn
Cn
Cd
Cd
Pb
3
7
1
2
Total 13
4
3
1
3
2
5
1
Total 19
Four Metals
Ni;
Ni;
Ni;
Ni;
Ni;
Ni;
Cu;
Cu;
Cr;
Five
Ni;
Ni;
Ni;
Cu;
Cu;
Cu;
Cu;
Cu;
Cr;
Cr;
Zn;
Zn;
Cr
Cr
Cr
Cr
Cd
Zn
Sn
Cd
Cd
9
;
9
;
9
J
>
9
;
Zn
Cd
Sn
Pm
Zn
Cd
Pm
Sn
Su
5
5
2
2
1
2
1
1
2
Total 21
Metals
Cu;
Cu;
Cu;
Cr
Cr
Zn
9
9
9
Zn;
Zn;
Cd;
Three Metals
Ni;
Ni;
Ni;
Ni;
Cu;
Cr;
Cu; Cr
Cu; Cd
Cu; Pm
Cr; Cd
Cr; Cd
Zn; Cd
14
1
2
2
1
3
Total 23
Six
Ni;
Ni;
Ni;
Ni;
Cd 5
Pm 1
Sn 3
Total 9
Metals
Cu;
Cu;
Cu;
Cr;
Cr
Cr
Cr
Zn
9
9
;
9
Zn;
Zn;
Zn;
Cd;
Cd; Sn 13
Cd; Pm 2
Cd; Pb 1
Sn; Pm 1
Total 17
Seven Metals
Ni; Cu; Cr; Zn; Cd;
Sn; Pb
Ni; Cu; Cr; Zn; Cd;
Sn; Pm
10
Total 11
17
-------
may, however, be special cases which were not brought to light in
this study—where chromium, lead, and zinc might be profitably won
from sludges, as, for example, if they occurred singly in a segregated
sludge and in such quantities to support adequate processing.
A review of Table 3 indicates that for most of the sludges reported,
more or less complex processes involving chemical separations would
be required.
18
-------
SECTION VII
FIELD STUDIES
Field studies were made for three purposes:
(3.) To augment the information gathered in the literature
review, and in the survey by questionnaire, by personal
conferences with company officials, inspection of waste
treatment facilities, and the sampling and analysis of
sludges produced.
(2) To obtain working samples of sludges for subsequent
laboratory studies.
(3) To evaluate the weathering characteristics of a lagoon-
stored sludge and to determine the extent of percolation
of leachate from the sludge bed into underlying ground.
Treatment and Disposal Practices
This program included visits to plants, inspections of sludge lagoons,
and the acquisition of actual sludges for use in the laboratory
studies. Information relating to the wastewater treatment which
results in sludge formation is summarized in Tables 4 and 5. Details
of each plant are given in Appendix C (Plant Visitations and Sludge
Characteristics).
As stated previously, one purpose of the field studies was to obtain
a first-hand view of the problem, so as to augment the information
gathered in the literature review and survey by questionnaire. Of
the seven plants visited, six could be classified as multiple-operation
metal-finishing plants employing electroplating (Tables 4, 5, 6, 7, 8,
9). One, an aluminum anodizing plant yielded little information helpful
to this study, its sludge being largely aluminum hydroxide and diatoma-
ceous earth used as filter-aid.
Characteristically, the lagoon-stored sludges from the metal-finishing
plants doing plating or pickling (Plants A, B, C, E, and G) were
found to be quite thin, containing only 2 to 4 percent solids.
Laboratory filtration of some of these sludges produced filter cakes
containing 10 to 16 percent solids. (Plant A, E, G). Several sludges
(B and C) were oily and were not filtered but dried directly.
19
-------
Table 4. TREATMENT AND DISPOSAL PRACTICES
Company
A
Cu
X
Ni
X
Cr
X
Zn Cd Sn
X
Pd Other
X
Treatment of
Wastewaters
Cr reduced to Cr with bisulfite.
Remarks
Mixed sludge to concrete
and
X X
X X
X X
,(1)
Phosphating
Ag plated
Al etched
Phosphat ing
Al anodizing
Metals precipitated as hydroxides with
CaO and/or NaOH, then to clarifier.
Acid and alkaline wastewaters mixed.
Sludge collected in sump. I'nprecipi-
tated metals plus unneutralized waste-
water to sewer.
Same as Company B.
Sludge from caustic etch vacuum fil-
tered.
Cr reduced to Cr with SO
fite. CM oxidized b\
9 or bisul-
Cl, first to CNO-
finally to Cu + N . Metals precipi-
tated as hydroxides, then to settling
t anks.
Cr reduced to Cr . Metals precipi-
tated as hydroxides, then to settling
tanks.
Cr+6reduced to Cr+3with bisulfite
CN completely destroyed by Cl oxida-
tion. Metals precipitated as hydrox-
ides, then to settling tanks.
block-lined lagoon. Water
95 + %. Sludge periodically
hauled to dump. Two lagoons
used alternately.
Sludge hauled to dump.
Same as Company H.
Sludge hauled to dump.
Mixed sludge to earthen lagoon.
Four lagoons available. Two
in service at a given time.
Efficient dewatering. Sludge
finally hauled to storage
lagoon. Water 607.
Two lagoons. When settling
occurs in :-"l, then supernatant
liquid runs into lagoon -"-"2.
Further settling here. Water
finally discharged to stream.
Water in sludge 95 + "'.
(1) These metals are found in the sludge as a result of the pickling of brass and copper.
-------
Table 5. SLUDGE CHARACTERISTICS
Si-nrpH fi1uHs»f> Fllt-pred Sludee
Company
A
B
C
D
E
F
G
% Solids
2 to 4
4
4
UK(3)
3
Ca 40(1)
3-4
Other % Solids
10-13
Organic
Material
Grease and Oil?
Organic Material
Grease Oil
Filter Aid 32(3)
Yellow-Green 11-13
--
Segregated 16
Variable Color
Dried Sludge
( ' Composition^ '
Cu Ni Cr pd Sn Ag Zn Cd Fe Al
3-5 3-5 5-15 0.1(5) 0.7
0.3 0.3 0.7 -- 3 — 3-5 0.2 3-6 —
3-5 1 1 0.2 -- 0.5 3-6 --
5-15
1 2-4 5-15 -- 0.04 -- 5-10 0.3 -- --
2-6 0.1-0.2 5-15 -- 0.1-0.3 -- 0.5-1.0
0.1-0.2 — 5-15 — 0.1 — 3-5 3-5 -- -- "
(1> Dried in lagoon, presumably by seepage and/or evaporation.
(2) Samples taken from lagoon filtered through Whitman #2 paper on Buchner funnel by vacuum for three minutes after
visible filtrate had stopped.
(3) Filtered through commercial rotary vacuum filter, using diatomaceous earth, at the company's plant.
(4) Semi-quantitative spectrographic analyses. Values shown in columns are for those metals used in plant.
-------
ro
(V)
Table 6. SEMI-QUANTITATIVE SPECTROSCOPIC ANALYSIS OF DRIED
SLUDGE
Weight Percent
Company
A
B
C
D
Cu
3-5*{"
0.3*
3-5*
0.03
Ni
> 3-5*
0.3*
1.*
0.01
Cr
5-15*
0.7*
1.*
0.1
Sn
0.7*
3.*
0.05
<0.01
Pd Zn
0.
<0.
<0.
<0.
1* <0.1
1 3-5*
1 0.1
1 <0.1
Cd
<0.1
0.2*
.0.5*
<0.1
AR
<0.005
<0.005
0.2*
<0.005
Al
0.03
2.
3-5*
5-15*
Mn
0.01
0.05
0.1
0.02
Fe
0.2
3-6*
3-6*
1
Ba
0.02
0.4
0.04
0.01
Weight Percent
A
B
C
D
B
0.05
<0.01
<0.01
-------
Table 7. SEMI-QUANTITATIVE SPECTROSCOPIC ANALYSIS OF DRIED
SLUDGE
Weight Percent
Sample Ca
E-l l.*(a)
(near inlet)
E-2 1.*
Ni Cr
2. -4* 5. -15.*
2. -4.* 5. -15.*
Zn
3. -5.
5. -10
Cd
* 0.2*
.* 0.3*
Sn
0.04
0.04
Fe
2. -4.
2. -4.
Mn Pd Ag
0.03 <0.1 <0.005
0.05 <0.1 <0.005
Al
0.2
0.3
Ba
0.01
0.01
Weight Percent
B
E-l 0.03
(near inlet)
E-2 0.02
Si Mg
2. 1.
2. 1.
Pb
0.04
0.03
Bi
<0.01
<0.01
Mo
<0.01
<0.01
Ca
3.-6.
3. -6.
Na Ti Zr
0.5 <0.01 <0.01
0.5 <0.01 <0.01
Co
<0.01
<0.01
Sr
0.04
0.02
(a) Elements whose values appear with an asterisk(*) beside them are those elements which one would ex-
pect to be present from the metal finishing operations.
-------
Table 8. SEMI-QUANTITATIVE SPECTROGRAPHIC ANALYSES OF THREE SAMPLES OF DRIED SLL'DGE
Sa-nnle
Number
F— 1
F-2
F-3
F-l
F-2
F-3
Location , ..
and DcpthU;
II Center
1-6 inches
II Center
12-18 inches
II N.W. Corner
Center
1-6 inches
Center
12-18 inches
K.W. Corner
12 inches
Cu
4.-6.*
-------
Table 9. SEMI-QUANTITATIVE SPECTROSCOPIC ANALYSIS OF DRIED
SLUDGE SAMPLES FROM COMPANY G
Cu Ni Cr Cd Zn Sn Pd Ag Al Mn Fe Ba B Si Mg Pb
G-l Southeast 0.1 0.03 5.-15. 3.-5. 3.-5. 0.1 <0.1 <0.005 0.07 0.03 2. <0.1 <0.01 2.-4. 0.2 0.04
G-3 South Center 0.2 0.05 5.-15. 3.-5 3.-5. 0.1 <0.1 <0.005 0.07 0.04 4. <0.1 0.01 2.-4. 0.3 0.05
Bi Mo Ca Na Ti Zr Co Be
G-l Southeast <0.01<0.01 5.-.10 0.1 0.2 0.01 <0.01 <0.01
G-3 South Center <0.01<0.01 10.-20. 0.1 0.1 <0.01 <0.01 <0.01
-------
Copper composition of the metal-finishing sludges, on a dry basis,
from plants in which copper was either plated or leached to pickle
(A, B, C, and F), ranged from 0.3 percent to 6 percent as determined
by semi-quantitative spectrographic analyses, with most of the values
ranging between 2 and 5 percent. The low value shown for copper in
the sludge from plant B (0.3 percent) may reflect a relatively minor
copper circuit within the plant or the loss of copper to the sewer.
Variations in the nickel analyses between 0.3 and 5 percent of those
plants plating nickel (Plant A, B, C, and E) is probably due to similar
causes. The same applies for those plants plating chromium. The
quantity of palladium apparently present in the dried sludge from
Plant A, which plated a thin tin-palladium layer on plastic in one
of its unit operations, is surprisingly high and may be the result of
analytical error. If indeed the sludge has 0.1 percent of palladium
in it, or 2 Ib. per ton, avoirdupois, this is the equivalent of about
30 troy oz per ton of palladium. The technology and economics of
recovering palladium, if it does exist in significant quantity in such
sludges would require a specific study. The concentration of tin in
the sludge from Plant B, about 60 Ib per ton, could also be attractive
economically. The silver in the sludge from Plant C amounting to about
60 troy oz per ton is also worth further consideration. The recovery
of palladium, tin, and silver, if they did exist in significant con-
centrations in sludges, would best be studied specifically.
There is considerable doubt the quantities of Zn and Cr in the various
sludges examined could be profitably exploited. Aluminum and iron are,
of course, valueless in such sludges.
Sludge Weathering and Leaching
A major objective of this work was to make on-site studies which would
determine the extent to which sludge in a given storage lagoon would
yield its toxic metal values to solution in rain, runoff, or intruding
ground water and so contaminate aquifiers.
The literature search had uncovered a report of the U.S. Geological
Survey that discussed an example of ground water contamination by
plating wastes held in storage basins or lagoons in Nassau County,
New York.
The uppermost aquifer (water bearing bed or stratum) in the South
Farmindale-Massapequa area consists chiefly of permeable deposits of
sand and gravel having a contaminated region ranging from 60 to 140
feet thick. The water table was about 1 to 2 feet below the bottom of
the lagoons in December. 1962. a seasonably low period, but in sorine
the water table may rise above the bottoms of the lagoons. Plating
26
-------
wastes were found to have seeped down from the lagoons into the aquifer
intermittently since 1941. This seepage resulted in a plume of con-
taminated water some 4,300 feet long, up to 1000 feet wide, and as
much as 70 feet deep, extending downgrade to the headwaters of
Massapequa Creek. The concentration of polluting metals in this plume
were often as high as 40 mg/1.
Originally this plating wastewater was untreated. Maximum treatment
began in 1958 apparently resulting in the formation of sludge which
effectively sealed the lagoon sides and bottom. Since the start of
chromium treatment, concentrations reportedly decreased to less than
5 mg/1 in most of the plume.
Work to determine whether such weathering and leaching had occurred
was dqne at the Company F lagoons. This site seemed particularly worth
investigating because the company had been utilizing these lagoons for
over 10 years.
In carrying out this investigation, shale structures underlying Company
F sludge lagoons were core-drilled to approximately 20 feet below the
bottom of the lagoon. The cores were analyzed for heavy metal compon-
ents known to be present in the stored sludge.
Details of the core-drilling procedure and analytical data are shown
in Appendix C, "Field Studies - Sludge Weathering and Leaching Studies".
No evidence of the leaching of metal ions into ground water was found
in this instance.
This is believed due to the fact that the sludge was, in the first
place, adequately treated by reduction and neutralization when stored,
and secondly, that the stratum underlying the lagoons was relatively
impermeable.
27
-------
SECTION VIII
LABOKATORY STUDIES
Work in the laboratory comprised a study of the leachability of sludge
samples obtained from the plants in the field studies; and detailed
investigations of three routes to the recovery of metals suggested by
the literature study to have some promise. These were electrowinning,
cementation, and liquid ion exchange.
Leaching
The leaching experiments were made on dried samples of sludge from each
of the seven plants that had been sampled. Tables 6 through 9 show the
complete semi-quantitative spectroscopic analysis of these sludges was
not made.
These dried sludges were leached under a variety of conditions in
mineral acids, solutions of ammonium salts, ammonium carbonate, and
sodium hydroxide. Sodium hydroxide was used in only a few runs for
the purpose of removing zinc.
The results, in general, indicated that copper, nickel, and chromium
in all the sludges were readily and almost completely dissolved in
mineral acids and ammonium carbonate solutions, as might be expected.
Residue weights after leaching varied from sludge to sludge and were
dependent on the type of leaching medium used.
The results of the leaching tests on the dried sludges are summarized
in Table 10 and are presented in more detail in Appendix D, "Selected
Leaching Experiments".
A few leaching tests were made on fully incinerated sludges, using
mineral acids and ammonium carbonate solutions. Details of these are
also included in Appendix D, "Selected Leaching Tests".
The results of these leaching experiments indicated that sulfuric acid
would yield acceptable extraction of copper, nickel, and chromium when
applied to undried or dried sludge. There is the possibility, however,
that was not explored, that calcined sludge might not respond as well
owing to the possible formation of silicates or ferrites during the
calcination. Although fairly drastic conditions with respect to time,
acid concentration, and temperature were used in most of the sulfuric
acid leaching tests, it is believed that in practice, the precipitated
28
-------
Table 10. SELECTED LEACHING EXPERIMENTS
ON DRIED SLUDGES
Leaching Conditions
Run
No.
1
2
3
5
7
8
9
10
11
12
13
17
19
23
24
25
26
27
28
29a
30
31
32
33
34
35
36
37
Sludge
Designation
A
It
II
11
II
II
II
11
11
II
11
B
C
E-l
E-2
E-l
E-2
E-2
Residue from 27
E-2
F-l
F-2
F-2
G-l
G-3
G-l
G-l
G-l
Wt.
OTIS.
25.0
25.0
25.0
100.0
10.0
10.0
10.0
10.0
10.0
10.0
10.0
15.0
15.0
10.0
10.0
10.0
10.0
10.0
25.0
10.0
10.0
10.0
10.0
10.0
10.0
10.0
10.0
Media
1*1 H SO
it
n
20% H2S04
20% (NH ) CO
20% (NH4)CL
20% (NH ) SO
20% (NH4)2C03
H
7% (NH4)2C03
14% NH4OH
35% (NH. ),CO,
4 L j
20% (NH ) CO
42 3
1:1 H2S04
1:1 H S04
35% (NH4)2C03
"
25% NaOH
1:1 H2S04
25% NaOH
1« 1 14 QM
:i H bu.
1:1 H2S04
it
23% (NH4> C03
1:1 H2S04
H
9 n°/ H en
20/. H2S04
20% H.SO.
2 4
23% (NH4)2O>3
Vol.
ml.
100
150
150
1000
200
200
200
200
200
200
200
200
200
100
100
200
200
100
100
250
40
150
150
125
150
150
100
100
125
Temp . ,
F
170
11
11
145
175
11
11
70
70
200
70
150
150
70-150
70-150
70
70
150
150
150
140
150
150
150
150
150
75
175
75
Time,
hrs.
16
11
11
17
6
6
6
168
24
23
71
6
6
8
8
69
69
2
2
1
1
1.5
1.5
4
1.5
1.5
24
24
2
Leaching Results
7. Weight
Lost
67.6
60.4
67.0
69.8
21.6
25.4
2.8
22.3
22.3
23.3
34.3
(a)
17. 51" ;
7.9
-------
hydroxides present In typical metal-finishing plant wastes will be
quickly and readily solubilized under much milder conditions.
In the leaching runs made with ammonium carbonate solution, copper
and nickel extraction were generally good while trivalent chromium
remained with the residue. This selectivity for copper and nickel
over chromium may be an advantage of the ammonium carbonate system.
As in the sulfuric acid leaching experiments, drastic conditions were
used in the ammonium carbonate leaching experiments. It is believed
that temperatures not much above ambient and relatively short leaching
periods should suffice to extract the bulk of nickel and copper present
as hydroxide in a conventional metal-finishing sludge. A disadvantage
of the ammonium carbonate system is that it will readily dissolve zinc
hydroxide from such sludges. Experience has shown that zinc is likely
to be more of an economic liability than an asset in the recovery of
metals from mixed wastes. Another possible disadvantage, one that was
not explored, is that ammonium carbonate can react with calcium sulfate
to form ammonium sulfate and calcium carbonate, with the result that
ammonium sulfate might be expected to build up in the leach liquor.
The leaching experiments also served to evaluate the various sludges
from the different plants for applicability to the recovery work.
Sludges from Plants B and C which were mixtures of miscellaneous plant
wastes containing oil, grease, and sulfuric materials, and which were
not the result of controlled treatment, were considered too atypical
for further work. The sludge from Plant D, largely an aluminum hydrox-
ide-dialomaceous earth mixture originating from the treatment of waste
caustic used to etch aluminum was also discarded from consideration.
Although it was noted that some of the sludges contained possibly
economic quantities of Sn, Pd, and Ag, no effort was made in these
leaching studies (or in subsequent recovery studies) to study their
behavior. The recovery of such metals properly requires a pointed
investigation.
The sludge from plants A, E, and G were selected for starting materials
in the recovery experiments.
30
-------
SECTION IX
RECOVERY EXPERIMENTS
Three approaches to the recovery of copper and nickel were investigated.
These were electrowinning, cementation, and liquid ion exchange. The
recovery of chromium was not considered.
Electrowinning
For reclamation of metal values from metal-finishing (large electro-
plating) wastes, electrowinning appears to be a readily feasible approach.
Electroplating plants have DC electrical power available; the process of
electrowinning is a process closely allied to electroplating; if the
leach liquor requires minimal preparation, electrowinning can be less
costly and more direct than many other methods of separation.
Preliminary Experiments
Copper, being fairly noble, can be electrodeposited from a variety of
solutions. The first solution from which it was successfully separated
in the work was the acidic solution which resulted from 1:1 sulfuric
acid leaching of sludge A. After filtering off the calcium sulfate, the
copper-containing filtrate was used without any modification and electro-
winning took place at constant current and room temperature. The depo-
site on a rotating stainless steel cathode was powdery. Agitation of
the solution and raising the temperature from 75 F to 122 F resulted
in an improved copper deposit. At a current density of 2 amps/sq ft,
an efficiency of 85 percent was achieved. Very high current densities
are not required but a ten-fold increase in this initial current density
was used in later experiments. Additional experiments on the electro-
winning of copper from sulfuric acid leach solution resulted in a better
deposit by use of small gelatin additions. A current density of 18 amps/
sq ft and a temperature of 126 F were used in these experiments. It
was not deemed necessary to produce a deposit comparable to a good grade
electroplate. If a powdery deposit forms, it can be removed from the
rotating cathode by a "doctor" blade which in effect scrapes the powder
from the cathode. With a horizontal rotating cathode which is only
partly immersed in the solution the copper could be collected in a
trough which runs along the side of the blade away from the cathode.
Where the deposit is adherent and fairly smooth, a thin copper starting
sheet can be bent around the rotating cathode. After a certain thickness
31
-------
of copper has been deposited, the starting sheet with the extra copper
can be removed and sold as copper scrap.
With all the copper removed from the leach liquor by electrowinning,
the next step was to attempt to electrowin nickel. Nickel could not be
electro-deposited from the sulfuric acid solution under any of the
conditions applied in the preliminary work. The failure was attributed
to the presence of interfering elements, possibly chromium.
I-t was found, however, that both copper and nickel could be deposited
from ammonium carbonate leach solution with minimal modification of the
solution. With constant current electrolysis, the cathode potential
changed slowly with time to more negative values as the solution became
depleted in copper. The result of this was that although copper alone
deposited at first, gradually the material depositing contained more
and more nickel, and finally consisted of nickel alone. Such a con-
stantly varying deposit of copper and nickel could probably be disposed
of by sale to makers of cupro-nickel.
In view of these promising preliminary results, the electrowinning work
was focussed on electrolysis from ammonium carbonate solution.
Electrolysis of (NH,) CO- Leach Solutions
Three-Compartment Cell Runs
three-
Following preliminary experiments with (NH.KCO-j leach liquors, a thr
compartment electrolysis cell was constructed from Lucite (Figure 1).
The general method for operating this cell is to place the (NH,)2CO-
leach liquor, including the insoluble portion, in the center compartment.
The slurry is agitated by the magnetic stirring bar, and the solution
containing copper and nickel passes through the diaphragms to the anode
and cathode compartments. For the first experiments, Number 2 Whatman
filter paper served as diaphragms. The advantage of the three-compart-
ment cell was that filtering and electrowinning took place simultan-
eously.
The first experiments in which the three-compartment cell was used were
run at constant current. As noted previously, copper and nickel code-
posited from ammonium carbonate solution, the composition of the deposit
depending on the elapsed time. The rate of diffusion through the filter
paper membrane initially selected was too slow to keep the copper concen-
tration constant in the cathode compartment. At the selected current
32
-------
Slurry
Rotating
Cathode
Graphite Anode
Magnetic Stirrer
Figure 1. Schematic drawing of
3-compartment cell
33
-------
density, the deposition rate exceeded the rate of diffusion of copper
into the cell compartment. At 20 amps/sq ft, the efficiency of 43 per-
cent for the copper-rich stage of electrowinning was judged to be too
low. The efficiency of the nickel-rich deposition averaged somewhat
higher, being 56 percent.
The low current efficiency was fairly constant throughout the period
of electrodeposition. This was found to be related to the rate at
which the surface of the rotating cathode moved through the solution.
In the experiments just cited, cathode rotation was at a rate of
130 linear inches per minute. This expression of movement is used
rather than rpm because several diameters of cathode were used.
Runs in Three-Compartment Cell
The three-compartment cell was then modified, so that after electro-
winning had been conducted for 1 hour, the catholyte could be pumped
continuously to the center compartment. This amounted to recycling of
the partially stripped ammonium carbonate solution from the cathode
compartment to the center compartment so that additional sludge (now
maintained as a slurry in the center compartment) could be dissolved,
and thus flow through the diaphragm as enriched leach liquor to the
catholyte. Fresh sludge was added to the center compartment period-
ically. Typical drop in efficiency with time for this type of cell is
shown in Figure 2.
Nylon diaphragms were substituted for the paper diaphragms in the
three-compartment cell. The rate of diffusion was more rapid through
the nylon, but some of the fine particles of sludge were not retained,
resulting in a dilute of light slurry in the cathode compartment. At
21 amps/sq ft and 115 F, with a rotating cathode linear speed of 259
linear inches/minute, the cathode efficiency started off at 32 percent
and became lower with time. By increasing the cathode linear speed
periodically the efficiency was restored to something near the original
value.
Electrolysis from a Slurry in a
Single Compartment Cell
The foregoing experience suggested that it might be of value to study
the electrowinning of copper and nickel from a relatively heavy slurry
in a single compartment cell. The slurry was agitated by a magnetic
stirrer so the degree of agitation was greater than indicated by the
cathode surface speed. Results were favorable in that nickel and
-------
100
u>
vn
60
120
180 240
Time, minutes
300
360
420
Figure 2. Current efficiency aa a function of time
-------
copper were effectively deposited without serious contamination from
suspended solids. This approach, however, requires additional work.
Electrolysis of Nickel and Copper by
Cathode Potential Control
Copper and nickel can be separated by means of controlled cathode
potential electrowinning. This requires an electronic instrument called
a potentiostat which will maintain a constant potential between the
cathode and some reference electrode such as the saturated calomel
electrode (SCE). By holding the cathode potential at about 0.3v
relative to the SCE, copper alone will deposit. When all the copper
has been removed from the solution, then the potentiostat is adjusted
to give the cathode a more negative potential and nickel will deposit.
A schematic drawing of the apparatus used is shown in Figure 3.
Controlled cathode potential electrowinning was used in one run with the
ammonium carbonate slurry of sludge A. The maximum cathode surface
speed (325 inches/minute) was used throughout the experiments. The
cathode potential was held constant at -0.75 volt versus SCE for the
electrowinning of copper. This is higher than the -0.3 volt given
above as the standard value, but the higher voltage was used to achieve
a higher current density. The solution consisted of 1 liter of slurry
containing 300 g/1 of (NH.) CO., and 50 g of sludge A. The slurry was
stirred for 1 to 2 hours prior to start of electrowinning. The initial
current density was 30 amps/sq ft and the average current density for
the 7.41 hours run was 11.5 amps/sq ft. Temperature was constant 111 F,
and gelatin solution was added periodically to maintain the concentration
at 1.0 g/1. Again, a magnetic stirrer was used to keep the sludge in
suspension. Periodic additions of sludge were made to replace the copper
that was deposited on the rotating cathode. Taking the copper content
of the dried sludge to be 5 percent, the recovery of copper was calcu-
lated at a value slightly over 100 percent.
The cathode potential was then raised to -1.25 volts versus SCE and
the electrowinning was continued. The initial current density was
25 amps/sq ft and the average for the 5.91 hour run was 13.8 amps/sq ft.
Taking the nickel content of the dried sludge to be 5 percent, 93 per-
cent of the nickel theoretically available was recovered. Probably a
small amount of nickel codeposited with the copper toward the end of
the first run, thus accounting for the recovery of over 100 percent.
The total weight of metal recovered by electrowinning was calculated as
100.5 percent of the copper and nickel present. Codeposition of nickel
with copper could be prevented by using a lower cathode potential,
possibly 0.5 volt.
36
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Potent iostat
+
O
Ammeter
Voltmeter
Carbon
Anode
Porous
Alundum
Cup
Probe
o
-Electrometer
Saturated
calomel
electrode
Rotating
'Cathode
Salt
Bridge
Magnetic
Stirrer
Figure 3. Schematic drawing for apparatus used for
constant cathode potential electrowinning
37
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In both cases, the deposits were acceptable for the purpose of resale.
Some copper flaked off the stainless steel cathode but this is easily
recovered by filtration.
A few exploratory electrowinning experiments were also made on
stripping liquors produced in a parallel liquid ion exchange study but
these were marked by extremely low efficiency due, it is believed, to
residual organics (naphthenic acid) carried into the stripping solution.
Further work on this phase of the study was deferred.
General conclusions from the electrowinning studies were that while the
approach held promise, considerable amounts of additional development
work would have to be done to establish the lines of a process that
would be technically and economically acceptable.
Cementation
In extractive metallurgy, copper is often separated by adding metallic
iron to the copper-containing acidic solution. Copper being more noble
precipitates in the metallic state and iron goes into solution according
to the following equation
In the single experiment performed to investigate the concentration of
copper, 100 g of dried sludge A were dissolved in 1 liter of 20 percent
H_SO, solution. Calcium sulfate was removed by filtering the mixture.
Tne iron was shredded pieces of 5 mil-thick low carbon steel. The
solution was heated to 130 F and was stirred slowly. The excess
undissolved iron was removed by magnetic means and the solution filtered
to recover the copper. The recovered copper weighed 4.15 grams. It
was in the form of a "mud" and could be washed so that it was fairly
pure.
The use of cementation to separate copper can be questioned because one
metal in solution is being replaced by another and this can be viewed as
not lessening the potential for pollution. However, iron is much less
toxic than copper. In addition, many sludges already have iron present
to the extent of several percent. The big advantage of cementation is
its relatively low cost. In specific instances, however, cementation
would be applicable. A limited amount of work was also done to inves-
tigate the cementation of cadmium. Dried sludge G contains 3 to 5
percent cadmium. The sludge was dissolved in 20 percent sulfuric acid
solution. After filtering off the CaSO, the filtrate was electrolyzed
at a controlled cathode potential of -0.3 volt to separate the small
38
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amount of copper. Following this the cadmium was cemented by the
addition of powdered zinc. The cadmium metal was then filtered off.
Recovery appeared to be good, the reclaimed cadmium being 4.04 percent
of the original quantity of dried sludge. In the case of cadmium
cementation, cadmium is replaced in solution by zinc. However, zinc is
already present in the solution and at a scrap price of about five cents
a pound it is hardly worth a special effort at recovery.
Liquid Ion Exchange
The possibility of employing liquid ion exchange for the recovery of
potentially valuable metals in metal-finishing wastes was investigated
in some detail in the laboratory. Both sulfuric acid and ammonium
carbonate leach liquors were employed as feed stock.
The laboratory work, described in detail in Appendix D, "Liquid Ion
Exchange", brought to light a host of problems including formation of
precipitates, emulsions, the necessity for precise pH admustment to
obtain selectivity, the carryover of extraction into stripping
solutions, in some cases the necessity for large numbers of extraction
and stripping stages, incomplete separations, etc.
On the basis of what has been revealed by the laboratory studies, it is
highly doubtful that liquid ion exchange could be considered a gener-
alized process for treating metal-finishing sludges.
39
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SECTION X
ECONOMIC STUDIES
Preliminary economic studies were made to determine the cost of small-
scale processing sludges of the compositional range encountered in the
sludges obtained from the various companies. Details are presented in
Appendix E.
The results of this study indicated that a batch experimental operation
treating 100 Ib per day of dry solids (equivalent to 3,330 Ib/day of
wet sludge would cost about $85 per day including amortization. Plant
cost for this small unit was estimated at about $15,000.
Metal recovery per day was assumed to be 5 Ib of copper and 0.3 Ib of
nickel. Obviously at this scale, type of operation, and with a sludge
of the composition assumed, there are no economics to consider.
If the scale of operation were increased ten-fold to about 1,000 Ib per
day of dry solids, of material containing 5 percent copper and 3 percent
nickel instead of 0.3 percent, and some credit be taken for chromium,
economics might be extrapolated as follows:
Capital Cost of Plant $90,000.00
Operating costs exclusive of labor 177.00
Labor (3 times that for 100 Ib/day plant) 180.00
Total daily operating cost $ 357.00
Revenue from Cu at 50 cent/lb (50 Ib) $ 25.00
Revenue from Ni at $1.00/lb (30 Ib) 30.00
Revenue from Cr at 10 cent/lb (100 Ib) 10.00
Total revenue $ 65.00
Further extrapolation to 10,000 Ib/day of dry solids yields:
Capital cost of plant $540,000
Operating costs exclusive of labor 1,200
-------
Labor (3 times that for 1,000 Ib plant) $ 540
Total daily operating cost $1,740.00
Revenue from copper, nickel, chromium 650.00
The inescapable conclusion is that a process of the type studies
operating on the sludges considered would not even approach good
economics.
Ul
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SECTION XI
APPENDICES
Page
A. Literature Review ^3
B. Details of Plant Survey Questionnaire 50
C. Details of Field Studies: Plant Visitations 52
and Sludge Characteristics; Sludge Weathering
and Leaching
D. Details of Laboratory Studies, Leaching 59
Experiments, Liquid Ion Exchange
E. Economic Studies 79
F. Bibliography 8^
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APPENDIX A
LITERATURE REVIEW
The literature search began with a perusal of the large number of
punched cards which had accrued /fs^a result of a previous search on
metal-finishing waste treatment . This previous search, although
primarily seeking knowledge of methods for treating metal-finishing
liquid wastes (rinses, etc.) was nevertheless quite comprehensive, and
some pertinent information was obtained.
The search continued throughout the course of the project as new tech-
nical requirements appeared.
A complete set of "Selected Water Resources Abstracts'^ which began
publication in 1968 was available, and these -have been studied through
to the current issue. "Pollution Abstracts"^ ' which began publication
in 1970 also has been consulted. "Chemical Abstracts" was used in
searching for specific topics and was not used for a comprehensive
search because such a search had been made in preparing reference (1).
The British "Water Pollution Abstracts" W was checked from 1947
(Volume 19) through May, 1971 (Volume 44). Also of value were the
yearly bibliographies or reviews of literature relating to plating
wastes as given in ''Plating11^ ' and the "Journal of the Water Pollution
Control Federation"^ .
For specific topics, such as, for example, the chemistry of a givai
element or the ..solubility of a compound, a series of general references
was consulted. For specific operations, such as electrodeposition,
filtration, etc., a number of references were available. '
Very little was found on the subject of reclamation of metal values from
metal-finishing waste treatment sludges. These sludges, resulting from
the precipitation of soluble metals as hydrated oxides, are usually
hauled away to a public or private landfill. In some locales, where
stringent pollution laws are not in effect, much of the sludge finds
its way into the sewer system or is run directly into a nearby stream.
(25)
Pepperl describes the incineration of electroplating sludges
containing copper, nickel, zinc, chromium, iron, and cyanides. It was
Numerical figures in parentheses relate to references in the biblio-
graphy (Appendix F).
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first necessary to filter the sludge to reduce the water content to 55
to 60 percent. This was done in a filter press or vacuum filter. The
incineration took place in a specially designed fusion cone incinerator
which used combustibles such as saw dust, wood shavings, varnished
papers, oil sludges, and varnish sludges, all of which were waste
products in the factory producing the electroplating sludges. Suitable
precautions were taken to minimize air pollution from the incineration.
The sludges containing about 60% water were incinerated at 900 to
1100 C. The volume of the materials was substantially reduced, the
final volume being about the seventh of the initial volume. The
material thus required less land for disposal and was decontaminated,
that is, the cyanides were completely destroyed and the metal compounds
were converted to oxides which, because of their low solubility, pose
less threat to the ground water. The author did mention in passing that
metal reclamation was a possibility. He did not, however, give any
indication of this having been done. The sludges were comparable in
composition to the sludges used in the experimental work on this
project.
Rehm and Nietz described the low- temperature incineration of
sludges. This work was done at temperatures of 200 C or below. They
found that, if the sludge is heated above 200 C, some of the trivalent
chromium is oxidized to hexavalent chromium in the presence of alkali.
The sludge was slightly alkaline since it had been treated with NaOH
or CaO to precipitate the hydroxides. When incinerated sludge, con-
taining highly soluble chromates, is dumped in a landfill the chromium
can find its way into the ground water. However, the low- temperature
incineration (if it can really be termed this) resulted in a large
reduction in volume, the final volume being reported as 1/13 of the
initial volume. Thus, a piece of land which would accommodate a 10-year
accumulation of sludge could not contain a 130-year accumulation.
(27)
Weiner also pointed to a reduction in volume and cautioned about
the danger of solubilizing zinc and chromium by heating sludge above
a certain temperature.
( 28)
Niemetz in a long discussion about sludge handling discussed
dewatering and incineration briefly. He mentioned the high cost of
dewatering, but gave no definite figures.
(29)
J. J. McGrath dealt with treatment of sludges at a brass mill.
These resulted from pickling of brass, and the sludge contained iron,
copper, chromium (from the chromate solution used in surface treatment
of brass), zinc, and calcium. This sludge was partially dewatered by
a rotary vacuum filter to a cake containing about 40 percent solids.
Some 28 tons of filter cake were produced per month. Although the
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author did not mention it, it was supposed that the sludge was dumped
at some landfill or other appropriate site. The author proposes
developing a use or market for the filtered sludge.
At this point it is appropriate to look to those references which
discuss the reclamation of metal values from metal-finishing wastes
before they are converted to sludge. The reference just cited (29)
was one of these. McGrath recovered copper by cementation on iron.
Iron in the divalent state also serves to reduce hexavalent chromium.
About 5 tons net weight of copper were recovered per month.
Several papers which describe reclamation, of. metal from waste solutions
were found. In a paper published in 1917 Jones described the
electrolytic recovery of copper (and sulfuric acid) from copper mill
pickling solutions. In one run of 60 days (10 hours per day) for an
expenditure of 1600 kilowatt days of electricity 3,755 pounds of
copper were recovered. This is equivalent to slightly less than
1 pound per kilowatt hour. The efficiency was 72 percent from a solution
containing 14 percent sulfuric acid and 3.5 percent copper. The
electrow inning was done on a continuous basis at 6.2 amps/sq ft and
2 volts. Other examples of recovering copper from pickling solutions
are given in a. publication of the British Non-Ferrous Metals Research
tion^ ;;, and those which have been consulted are listed here
Association ;;, and those which have been consute are ste
(32, 33, 34, 35, 3b)^ A11 of these discuss the electrowinning of
copper from waste solutions, and relevant information in these publi-
cations was used in choosing initial experimental conditions for this
program.
Sierp ' preferred a copper concentration of 2 percent with the
sulfuric acid at 5 percent. A good copper deposit was obtained at
5 amps/sq ft and 2.3 volts in the presence of up to 10 percent zinc.
Recovery of copper took place at almost 100 percent of theoretical
efficiency. The solution was agitated by pumping. The anodes were
8 percent antimonial lead. Sierp calculated that the energy requirement
was about 1.9 kwh per pound of copper deposited but he includes
conversion losses (a-c to d-c), distribution losses, etc.
Other references ' give data which are in approximate
agreement with the findings cited in the above two papers.
Muller also discussed waste recovery in the copper and brass
industries. He described a transportable electrolytic copper reclama-
tion unit for use in large plants where pickling units are scattered.
He found recovery of zinc by electrowinning from acid baths was some-
what difficult and since the scrap value of zinc is low, he concluded
that the effort might not be economically feasible.
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/OQ\
Wise and Wise and Dodge studied the reclamation of metal values
in the brass and copper industry. The first of these two papers
describes how reclaimed electrolytically, copper was deposited on brass
chips by cementation (in this process copper is displaced from solution
by the zinc in the brass), and, trivalent chromium was oxidized to the
hexavalent form by roasting . The zinc was electrodeposited from
acid sulfate solutions and_no great difficulty was mentioned, contrary
to the remarks of Muller . The subsequent paper by Wise and Dodge
includes a description of how trivalent chromium was oxidized electro-
lytically to the hexavalent form .
The electrolytic regeneration of chromic acid was discussed in some
detail by McKee and Leo . The cell in which the oxidation took
place utilized a diaphragm to isolate the anolyte in order to keep the
cathode from reducing the hexavalent chromium. Best results were
obtained with the temperature at 70 F. Lead electrodes were used and
the starting electrolyte was essentially a sulfuric acid solution of
trivalent chromium. The current efficiency was 59 percent and the
authors judge the process to be economically feasible.
(41)
Mitter and Dighe reported a process for recovering dichromate and
sulfuric acid from the waste pickle liquor used to pickle coinage
metal. Again, a diaphragm cell was used. Initially the current
efficiency was 90 percent but after 8 hours had dropped to 40 percent.
Descriptions of the commercial methods for producing chromium compounds
were reviewed for relevant processing information . In one method,
sodium chromate is prepared by first roasting chrome ore with sodium
carbonate or sodium carbonate and lime. The chromate is then leached
from the mix with water. Chrome ore contains chromium in the form of
Cr20o and the waste sludges under investigation contain hydrated
Cr?0, so that the similarity is apparent. Chromium trioxide (CrCL),
orchromic anhydride as it is also called, is prepared by reacting
dichromate with sulfuric acid (66 Be or 96 w/o) . The temperature is
gradually raised to just above the melting point of CrO . The mixture
separates into an upper layer of predominantly sodium bisulfate and a
lower layer of molten CrO.,.
The Bureau of Mines has been active in waste reclamation for some years.
Several publications_relating to plating wastes have appeared recently.
Three of these ' ' are concerned with electroplating wastes.
The initial work was carried out on HNO_ rack-stripping solution. The
solution was neutralized with lime to pH 2.3-2.7, causing precipitation
of ferric hydroxide which contained small amounts of copper and nickel.
After precipitation of iron, the copper was reclaimed by controlled
cathode potential electrolysis. The nickel was recovered by precipi-
tation as nickel hydroxide, or treated with sulfuric acid and heated to
100 C to expel the HNO~ and precipitate CaSO, after which it was
recycled as NiSO, to the nickel plating bath.
1*6
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An organic-cyanide stripping solution also was used and collected as
waste. It was found technically economically feasible to accomplish
the neutralization by adding nitric acid waste to the organic-cyanide
waste. The metals precipitated at pH 4.5 as cyanides. This method is
called the "waste-plus-waste" method. The precipitate was heated in
air at 250 C to convert the cyanides to oxides which were leached with
10 percent sulfuric acid. The iron was then precipitated with lime,
and the procedure used above was then followed to reclaim copper and
nickel values.
The Bureau of Mines.have also studied reclamation of aluminum-bearing
electronic scrap . Two methods were described. In the first the
aluminum was separated by a sodium hydroxide leach with the copper
eventually being recovered electrolytically. -A precious metal sludge
formed during the electrowinning of copper and the precious metals were
recovered by conventional means. In a second method, the aluminum was
separated by fused salt electrolysis and the copper and precious metals
recovered as in the first method.
Two publications' ' ' by Bureau of Mines personnel describe tech-
niques in the chemical reclaiming of superalloy scrap. Of interest to
this project was the method for separating chromium and nickel. Prior
to the chromium-nickel separation, liquid-liquid ion exchange techniques
were used to extract cobalt, and the chromium and nickel were contained
in the acidic raffinate from this separation. Chromium basic sulfate
was selectively precipitated using sodium sulfate, sodium carbonate,
and urea. The chromium precipitate was contaminated with sodium sulfate
so the precipitate was redissolved in sulfuric acid. The chromium was
reprecipitated with soda ash, yielding a basic sulfate considered
suitable as a leather tanning agent.
Finally, the nickel was precipitated from the chromium-free leach
liquor as basic carbonate by addition of soda ash to pH 8. The
precipitate was filtered, thoroughly washed, dried, and roasted to
give a steelmaking grade of NiO containing 77 percent nickel (the
stoichiometric quantity is 78.6 percent). Another method which combines
liquid-liquid extraction and electrowinning to reclaim copper and nickel
is described in reference (47a).
References (18, (20), (21), and (22) provided data on electrowinning.
Reference (18) contained detailed data on the methods used in electro-
winning of copper, nickel, and chromium. Copper, because of its noble
character, can be deposited from most solutions; nickel is more difficult
to reduce electrolytically; and the deposition of chromium from
solutions of trivalent chromium is the most difficult of the three.
Although electrowinning of chromium has been successful, a high degree
of control is required and, in general, results can be erratic. Infor-
mation also was obtained from the above four references on the other
three base metals which are commonly electrodeposited.
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Anodic Oxidation
In reference (40) McKee and Leo described an electrolytic process for
regenerating reduced solutions of chromic acid. By using a diaphragm
cell Cr can be oxidized to Cr using a lead anode and a current
density of 12 amps/sq ft in a sulfuric acid solution at 70 F. McKee
and Leo reported an anode efficiency of 59 percent. In experiments
for this project similar efficiencies were obtained. About three
kilowatt hours are required to produce a pound of
When zinc is present and the chromium is anodically oxidated in acid
solution, zinc chromate, an insoluble compound, is formed. Zinc
dichromate, on the other hand, is soluble.
One of the easiest and cheapest methods for separating copper from an
acid leach solution is cementation. In this process copper is dis-
placed from the solution by metallic iron according to the following
equation
Attesting to the low cost of such an operation is the fact that at
Inspiration Consolidated Copper Company as much as 1,000,000 pounds of
cement copper have been produced per month witb.gne cranetnan and three
laborers per shift, working two shifts per day^ . The process
removes 97 to 98 percent of the copper.
Habashi in his chapter on precipitation by metals also discusses
the conditions for cementing of copper on iron. In addition, he
refers to several works on precipitation of cadmium or nickel by zinc.
Electrowinning or other means for reclaiming the metal values from
mixed sludges cannot be used until the sludge is solubilized. In the
minerals industries the leaching of metal values from ores is an exten-
sively used process. Information on various leaching agents and the
chemistry of leaching was also reviewed (48, 49, 50). The ideal
leaching agents would be those which were highly selective to individual
metals and, therefore, would make for easy separation of the compounds
in sludge. Such leaching agents were not found in the literature.
Sulfuric or hydrochloric acids are low in cost and would dissolve any
of the hydrated oxides. Hydrochloric acid has the disadvantage of
being very corrosive which makes it difficult to find suitable materials
of construction that are not overly expensive. Neither of the acids
is selective. Ammonium carbonate and other ammonium salts have good
leaching qualities. The use of additional ammonia with these salts
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may or may not be desirable. In some cases in extractive metallurgy
more than one leach is necessary to get all the metal into solution.
Agitation, temperature, time, and the particle size of the solid
material are of importance.
In recent years liquid-liquid extraction, or liquid ion exchange as it
is also called, has been studied for use in extractive metallurgy. The
technique has been successful for the recovery of uranium and thorium
for nuclear applications. Refractory metals such as zirconium and
throium also have been recovered by liquid ion exchange techniques.
The literature on this technique is very extensive and the research
has included just about all metals. Not all of these, however, have
found commercial application.
Several excellent review papers were found (51, 52, 53, 54). Many
different organic compounds have been used for liquid extraction. The
papers of Fletcher and Associates, using naphthenic acid seemed
directly applicable to this work. Their numerous publications (55, 56,
57, 58, 59, 60) provided information used to guide the experimental
work in this program.
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APPENDIX B
DETAILS OF PLANT SURVEY QUESTIONNAIRE
(1) Do you accumulate waste sludge from your metal finishing operation?
( ) Yes ( ) No
(2) Is this sludge the result of controlled neutralization, in a
special treatment setup, of waste rinse waters and/or neutrali-
zation of spent plating solutions?
( ) Yes ( ) No
(3) Is this sludge the result of uncontrolled neutralization of
waste rinse waters and/or neutralization of spent plating
solutions? (Example-allowing cleaner rinse to neutralize acidic
rinse water from plating baths)
( ) Yes ( ) No
(4) If you do not have a treatment plant for neutralizing waste water
and handling sludge, are you planning to install one in the near
future ?
( ) Yes ( ) No
(5) Has there been any activity in your locality on the part of local,
state, or federal agencies to require waste-water treatment?
( ) Yes ( ) No
(6) What is the approximate volume or estimated weight of sludge that
you accumulate in a year?
(7) How do you now dispose of your sludge?
( ) Public landfill or dump
( ) Private landfill or dump
( ) Sell to private firm for reclamation
(8) Do you make or have you ever made an attempt to reclaim metal
values from the sludge?
( ) Yes ( ) NO
50
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(9) What metals might be found in your sludge?
( ) Copper ( ) Zinc ( ) Precious metals
( ) Nickel ( ) Cadmium (specify)
( ) Chromium ( ) Tin ( ) Others (specify)
(10) Would you be willing to cooperate in the MFF--sponsored project
by allowing Battelle personnel to take sludge samples from your
plant?
( ) Yes ( ) No
(11) If the answer to question 10 is yes, please give name of person
to be contacted.
(12) Please give the name, address, and telephone number of your
company.
Upon completion of this questionnaire, please return to:
Arch B. Tripler, Jr.
Waste Control and Process Technology Division
Battelle Memorial Institute
Columbus Laboratories
505 King Avenue
Columbus, Ohio 43201
51
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APPENDIX C
DETAILS OF FIELD STUDIES; PLANT VISITATIONS
AMD SLUDGE CHARACTERISTICS; SLUDGE WEATHERING AMD LEACHING
Company A. This company has more than one plant. The one from which
the sludge samples were obtained plates copper, nickel, and chromium
on plastics. No cyanide solutions are used. The facility in which
wastewaters are treated is automatically controlled. The hexavalent
chromium is reduced with bisulfite and then trivalent chromium, copper,
and nickel are precipitated as hydroxides by raising the pH to about
8 with sodium hydroxide. In the past lime had been used in neutral-
ization to promote flocculation and some calcium is still found in the
sludge lagoons. The hydroxide slurry is pumped to a clarifier, and
finally to one of two cement- and block-lined lagoons. The settled
slugde is removed from the lagoon and hauled to a municipal landfill.
The water going into the sewer contains no dissolved solids in excess
of local standards. Some tin is found in the sludge because in making
the plastic surface conductive, a very thin tin-palladium layer is
deposited, with the tin being subsequently stripped. The sludge in
the lagoon was very "clean", that is, it looked uniform and free of
extraneous materials.
Company B. At this facility sludge is collected in a sump and sludge
formation depends on chance neutralization as acid and alkaline waste-
waters are brought together. Precipitation probably is not complete,
allowing some of the dissolved metals to pass to the sewer. It is
suspected that the sump is not efficient in retaining all the precipi-
tate and some of the sludge also passes to the sewer. The resultant
wastewater passing to the sewer may be alkaline or acid depending on
which of the wastewaters was in excess.
Copper, nickel, chromium, zinc, tin, and cadmium are plated at Company
B. Phosphating is also done. The sludge had oil, grease, and other
extraneous matter mixed with it and, judging from the variations in
color, the sludge was segregated within the sump. The sump is period-
ically pumped out and the sludge hauled to a dump for disposal.
Company C. The sludge at Company C also is collected in a sump. Just
as at Company B, the acid and alkaline wastes are mixed, with neutral-
ization occurring and the resultant wastewater being acid or alkaline
depending on which was in excess at the time of mixing. Copper, nickel,
chromium, cadmium, and silver are plated. Some metal surfaces are phos-
phated and aluminum is etched. The sludge appears to be even more
52
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highly segregated than that from Company B. The sludge contains extran-
eous material such as cloth, wood, paper, and oil or grease. When the
sump is filled, the sludge is pumped out and hauled to a dump.
Company D. This company anodizes aluminum and the sludge that is
collected precipitates from the alkaline solution used to etch the
aluminum prior to anodizing. This sludge is separated from the
wastewater by a rotary vacuum filter. Presumably the sludge is
largely sodium aluminate, AL(OH),..
Company E. Copper, nickel, chromium, zinc, and cadmium are plated at
Company E. This plant treats its wastewater by reducing hexavalent
chromium to the trivalent state using SO or bisulfite, and destroying
cyanide completely in two stages with the alkaline chlorine treatment.
The metals are precipitated as hydroxides with NaOH or CaO. The sludge
goes to a settling tank and then to the earthen lagoon. The sludge
appeared segregated and samples were taken from several places.
Collection of sludge at this plant was started only a short time prior
to our sampling. It is estimated that about 25,000 pounds will be
collected per year. No arrangement had been made for its disposal.
Company F. This company has been accumulating sludge for 10 to 11
years in a storage lagoon. Because of this relatively long storage
period, the lagoon was of interest for the study of "weathering" effects
on solubility and leaching.
The metals which are found in the wastewaters get there largely as a
result of the pickling of brass, copper, and bronze. Hence, one would
expect copper, zinc, and tin in addition to chromium. Hexavalent
chromium plays a large part in this process and part of the waste
treatment consists of reducing this hexavalent chromium to the trivalent
state. The metals are precipitated as hydroxides by addition of lime.
The stream passes through a series of settling tanks and finally arrives
at the holding lagoons where final settling occurs, and also, some
natural dewatering. When the sludge reaches a certain stage of dryness,
it is transferred to the storage lagoon.
Company G. Company G does copper, chromium, cadmium, and zinc plating.
Wastewaters are treated to reduce chromium and to destroy cyanide
completely. After some settling the sludge flows by gravity to a lagoon
where further settling takes place. The supernatant liquid is then run
53
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by gravity to a second lagoon where any suspended matter is permitted
to settle. Finally, the supernatant liquid from the second lagoon goes
to a nearby stream. The sludge was taken from the first lagoon and
appeared segregated so samples were taken at several places. Dewatering
is not nearly so efficient as with Company F sludge.
Sludge Characteristics
Determination of the amount of water in a given sludge is difficult
because the sludge concentration varies with depth due to settling;
the storage lagoons vary in depth, and the dense layer of settled sludge
varies in thickness, being thickest at the point where sludge enters
the lagoon. The samples for use in this work were taken at several spots
in a given lagoon if possible. A stainless steel conical, 1/2-gailon
sampler was plunged into the sludge layer which was usually 2 to 3
feet below the top surface of the liquid, held a few seconds and
withdrawn. The samples were vacuum filtered using a Whatman No. 2
paper, and after all the solids were in the Buchner funnel, the vacuum
was maintained until no visible water passed through the funnel for a
period of three minutes. This was arbitrary but it did provide a
uniform basis for filtration. The sludges filtered fairly easily.
Dried samples were prepared by heating the filter cake to 200 F for
several hours. Semi-quantitative spectrographic analyses were made
of these dried samples.
Company A. The first batch of sludge obtained from this company was
rather "thin"; that is, the percent solids was low, being about 1.6
percent of the slurry as sampled. The second batch obtained was much
thicker and contained at least 4 percent solids. However, these
figures are semi-quantitative because uniform sampling is not possible.
The filter cake from the first batch contained 87 percent water, and
from the second batch about 80 percent water. Even with 87 percent
water the filter cake can be added to the leaching agent and little
dilution occurs.
Companies B and C. The sludges from these companies were very much
alike in that they contained organic material. This organic material
appeared to be partly grease and oil and would be expected to coat
sludge particles and prevent contact with leaching agents. Dried
samples of each were incinerated, and it was observed that Sludge B
burned autogeneously once it had started to burn. Sludge C did not
burn autogeneously. The gross heating values of dried samples of
Sludges B and C were measured calorimetrically as 2155 Btu/lb, and
1170 Btu/lb, respectively.
-------
The solids contents of these samples were about the same as for the
second sample of Sludge A, that is, about 4 percent.
Company D. The "as received" sludge from Company D was moist filter
cake from the vacuum filtration. It contained 32 percent solids. The
sludge contained about 10 percent aluminum and 10 percent silicon.
Petrographic examination of an incinerated sample showed a fairly
large amount of diatomaceous filter aid material present. Filter aid
was used on the rotary vacuum filter.
Little further work was done with this sludge because it appeared
unlikely that the aluminum could be profitably recovered.
Company E. The solids content of the sludge samples "as taken"
averaged 2.7 percent. The filter cake obtained by the standard method
described above from two samples contained 13 percent and 11.4 percent
solids, respectively. The filter cake had a yellowish green color.
Company F. The sludge from Company F was quite dry in the "as taken"
condition. In taking the samples one could walk on the surface of the
sludge in the storage lagoon. For three samples the solids contents
were 41.7 percent, 36.5 percent, and 43.5 percent, respectively. The
solids contents here were higher than those obtained from filter cakes
of previously described "wet" sludges. This is an indication of the
efficiency of "air and solar" dewatering obtained by the sludge handling
system at Company F.
Company G. Two samples of "as sampled" sludge from Company G contained
3 percent and 3.9 percent solids. The moist filter cake contained 16
percent solids. The overall color of one sample was brown with spots
of bluish-green and yellow, while a second sample appeared blue-green
with spots of yellow-brown. The sludge appeared highly segregated.
The first sample was taken near the entry pipe and the second was taken
farther away.
Sludge Weathering and Leaching
A schematic diagram of the locations of the three core drillings and
the results of a gross examination of the structures just beneath the
sludge bed are shown in Figure C-l. At the indicated locations, casings
were driven down through the sludge into the shale formation beneath
the sludge bed. The sludge cores were taken from within these casings.
55
-------
For all three sludge cores, two sludge samples were taken from within
the 1-foot segment of the sludge cores. An extra sludge sample was
taken from the 4 to 5 foot segment of core B-l, making a total of three
samples for this segment of this core.
After all the sludge had been removed from within the casings, the
underlying shale structure was cored. The core drill was dropped
through the now-empty casings to ensure that the shale cores were not
contaminated by the sludge.
The sludge samples were placed in glass jars, the shale cores in
special wooden cases; then the entire lot was shipped to Battelie's
Columbus Laboratories.
Figure C-l indicates the degree of recovery of the shale cores. Less
than complete recovery was obtained where the shale formation was
partly or completely decomposed, or where siltstone was encountered.
Two shale core samples from each of the three bore holes were selected
for analysis. The samples represented the portion of the core from
immediately below the sludge bed and the portion at the bottom of the
bore hole. Analyses were also made on sludge samples taken from as
near the bottom of the sludge bed as possible. There was not a clear
demarcation where the sludge stopped and the shale began. This is
probably due to the fact that there was loose shale left at the bottom
of the lagoon prior to adding the first load of sludge. Therefore,
those sludge samples were selected which, upon visual inspection, had
no shale present.
The six core samples were first crushed in a jaw-crusher, then in a
roll crusher, and finally in a pulverizer. The powdered samples were
rolled in a roll mixer for several hours and small samples were taken
for analysis. The three sludge samples were dried, ground, thoroughly
mixed, and sampled.
Table C-l shows the results of the semi-quantitative spectrographic
analyses.
56
-------
B-l
OB-2
B-3
35ft
70ft
7.0'.
100% Rec.
10.0'.
100 % Rec.
12.5'.
100 % Rec.
14.8'.
100 % Rec. —^ 156'
63% Rec.
I7.51.
80% Rec. ,
I9.0'_
75% Rec.
21.5-
i
\
/•{
\
y>
a
2
2
A.
••Sludge
7.0'-
25%,
Rec.
Red shale Rec.
and
siltatone
100%,
Rec.
64%x
Rec.
6.0'-
10.0'-
1 1 .5'-
' I2.0'~
i
> 5.
17.0'.,
,. 24.0'
r
I
V*
V"
•y>
v
/
Y
y
y1
^
^
'/
f
/
to solid shale r
44%-^
Rec.
y I4.0_
Red shale
t. badly broken
and fractured
33%<
Rec.
(
19.0.
. 24.0'.
^
/
•y
V
'/
y
y
y
/
/
^
/
^
^
^
<
/
/
/
/
^
/
s.
1
/
/,
• some clay and
fine to course
gravel
Decomp. to
, portly decomp.
shale
Solid to
, partly decomp.
shale
Figure C-l. Schematic drawing of location of
bore-holes, results of gross examination
of cores, and percentage recovery of cores
57
-------
Table C-l. SEMI-QUANTITATIVE SPECTROGRAPHIC ANALYSIS OF SELECTED
SLUDGE AND CORE-DRILLING SAMPLES
VJl
00
SAMPLE1
NUMBER
Bore
Hole
Approx-
imate
Depth
Weight
Percent
Cu
Ni
Cr
Zn
Fe
Ca
Sn
Ba
S
Si
Mn
Mg
Al
Mo
Na
V
Ti
Zr
Pb
Ag
sludge
No. 13
B-l
5'8"-6'0"
(a)
2. -4.
0.1
5. -15.
0.2
0.5
5. -10.
0.2
<0.01
0.01
2. -4.
<0.01
0.3
2.
<0.01
0.1
--
0.05
<0.01
0.05
<0,005
Core A-l
B-l
7'0"-7'9"
.005
.005
<0.01
0.1
3. -6.
2.
<0.01
0.03
0.02
10. -20.
0.05
2.
5. -15.
<0.01
2.
0.01
0.3
0.01
--
--
Core t-2
B-l
19I9"-21'6"
.005
0.01
0.02
<0.1
3. -6.
1.
<0.01
0.03
0.02
10. -20=
0.03
2.
5. -15.
<0.01
2.
0.01
0.3
0.01
--
--
No. 13
B-2
6'6"-7'0"
2. -4.
0.3
5. -10.
0.3
0.7
5. -15.
0.2
<0.01
0.01
2. -4.
<0.01
0.4
2.
<0.01
oa
--
0.1
<0.01
0.04
<0.005
Core A-l
B-2
8'0"-12'0"
.005
0.005
<0.01
<0.01
3. -6.
0.5
<0.01
0.03
0.02
10. -20.
0.05
2.
5. -15.
<0.01
2.
0.01
0.3
0.01
—
- -
Core D-3
B-2
23'1"-24I0"
.005
<0.005
<0.01
<0.1
3. -6.
5.
<0.01
0.03
0.01
10. -20.
0.1
2.
5. -15.
<0.01
3.
0.01
0.3
0.01
--
- -
Sludge
No. 13
B-3
3I6"-4'0"
3. -6.
0.8
5. -15.
2.
0.7
5. -10.
0.4
<0.01
0.02
3. -5.
<0.01
0.6
3,
<0.01
0.1
--
0.1
<0.01
0.08
<0.005
Core A-l
B-3
U'O'^U'C"
0.01
<0.005
<0.01
<0.1
3. -6.
0.5
<0.01
0.03
0.01
10. -20.
oa
1.
5. -15.
<0.01
3.
0.01
0.3
0.01
--
"* *™
Core B-2
B-3
19'0"-24'0"
0.005
<0.005
<0.01
-------
APPENDIX D
DETAILS OF LABORATORY STUDIES, LEACHING
EXPERIMENTS, LIQUID ION EXCHANGE
Leaching Experiments
For the most part, leaching experiments were made on dried sludge.
(Doing this made possible quantitative experiments, because with filter
cake there were variations in moisture content. However, in a
commercial process filter cake would be used.) Some selectivity was
shown by ammonium carbonate solutions where chromium compounds
remained largely undissolved while nickel and copper compounds were
solubilized. Sulfuric acid solutions left calcium sulfate undissolved
but did not differentiate between the heavy compounds.
Because the sludges showed individual leaching characteristics, they
are discussed below on an individual basis.
The conditions and results of the leaching tests are given in Table D-l.
Company A Sludge
Sludge from Company A was, for a time during the early stages of this
project, the only suitable sludge available. It contained only three
heavy metals in significant quantities, and it was very clean, having
neither grease nor oil nor other gross impurities.
The undissolved material consisted mostly of CaSO,. The difference in
the percentage dissolved between experiments 2 and 3 is attributed to
difference in agitation.
When Sludge A was incinerated at 1200 F it lost 25 percent of its
weight. This is mostly due to water loss. Drying sludge at 200 F
does not rid the hydrated oxides of all the water. Incinerating also
oxidizes part of the insoluble trivalent chromium to soluble hexavalent
chromium. If incineration is used to reduce volume prior to dumping,
then the hexavalent chromium can leach out and find its way into the
ground water. This was discussed in reference (26). The concentration
of reclaimable metals is higher in the incinerated sludge and this is
an advantage because where leaching follows incineration less material
handling is required. The higher concentration of metal values is
reflected in the smaller amount dissolved in experiment 4.
59
-------
Table D-l, SELECTED LEACHING EXPERIMENTS
Run
No.
1
2
3
4
5
15
15a
16
16a
23
24
30
31
33
34
5
5
5
7
7a
7b
7d
10
13
13a
13b
13c
14
17
18
19
19
25.
25a
25b
11
lib
12
32
12a
26
26a
26b
8
8u
8b
8c
9
9a
9b
9c
21
22
27
28
29a
29b
Sludge
A
A
A
A
A
B
B
C
C
E-l
E-2
F-l
F-2
G-i
G-3
A
G-l
G-l
A
A
A
A
A
A
A
A
A
A
B
E
C
C
E-l
E-l
E-l
A
A
A
F-l
(Above residue)
E-2
E-2
E-2
A
A
A
A
A
A
A
A
D
D
E-2
E-2
E-2
E-2
Condition
Dried
Dried
Dried
Incinerated
Dried
Incinerated
(Residue from above)
Incinerated
(Residue from above)
Dried
Dried
Dried
Dried
Dried
Dried
Dried
Dried
Dried
Dried
(Residue from above)
(Residue from above)
(Residue from above)
Dried
Dried
(Residue from above)
(Residue from nbove)
(Residue from above)
Incinerated
Dried
Incinerated
Dried
Incinerated
Dried
(Residue from above)
(Residue from above)
Dried
(Residue from above)
Dried
137, NH.OH Dried
137 NH,OH Dried
4
Dried
(Residue from above)
(Residue from above)
Dried
(Residue from above)
(Residue from above)
(Residue from above)
Dried
(Residue from above)
(Residue from above)
(Residue from above)
Incinerated
Incinerated
Dried
(Residue from above)
Dried
(Residue from above)
Wt.
gm.
25.0
25.0
25,0
10,0
100.0
20.0
12.9
20.0
17.8
10.0
10.0
10.0
10.0
10.0
10.0
100.0
10.0
10.0
10.0
7.8
7.3
ND
10.0
10 0
6.6
5,6
50
5.0
15.0
15.0
15.0
15.0
10.0
7,9
6.8
10.0
7.7
10.0
10.0
8.6
10.0
7.4
6.7
10.0
7.5
6.6
ND
10.0
9.7
7.9
ND
5.0
15.0
10.0
6.8
25
ND
Lea chant
1:1 H SO
1:1 H,S07
1:1 H,S07
1:1 H,S04
307, H,S07
1:1 H,SO*
1:1 1CSO''
1:1 H,S07
1:1 H,SO?
1:1 H,SO^
1:1 HjSO*
1:1 H^SO*
1:1 H^SO;)
1:1 H,S07
1:1 H^SO*
307. H SO
207 ICSO*
207. HjSO^
207, (NH, ) CO
207, (NH*) Co..
207, (NH7),CO^
1:1 H SO2
207 (NH, ) CO.,
357 (NH7),CO^
357 (NIC), OK
357, (NH7),CO,
1:1 H.SO,2 3
207 (fiH J CO
207 (KH7),co,
207, (wCKco^
207, (NHp^Cof
207. (NH,),CO,
H L J
357, (NH4)2C03
357, (NH, )«CO.,
357, (NH^KCO3
207 (NH ) CO
1:1 H2S04Z
147, (NH ) CO
237.(KH,J.CO,:».'H,
237. (NH, ) CO^+KH"*
423 4
207 (NH ) CO
207. (NH7),CO::
207, (Nn7)2CO:;
207, N'H^Cl
207. NH.C1
207 KTCCl
1:1 H.SO,
207 (NH J SO
207, (NHp,S07
207 (NHp^SO^
1:1 H_SO -
I 4
1:1 HC1
1:1 HC1
257. NaOH
1:1 H SO
257, NaOH
1:1 H-SO,
Vol.
ml.
100
150
150
150
1000
75
75
75
75
100
100
150
150
150
150
1000
100
100
200
200
200
100
200
200
200
200
175
100
200
200
200
200
200
200
200
200
175
200
OH
OH
200
200
200
200
200
200
100
200
200
200
100
100
300
100
40
250
150
Temp.
F.
170
170
170
150
145
180
195
180
190
70
150
150
150
150
150
145
75
75
175
175
175
175
70
70
70
70
140
150
150
150
150
150
70
70
70
70
70
200
150
150
70
70
70
175
175
175
175
175
175
175
175
150
150
150
140
150
140
Time
hrs.
16
11-1/4
11-1/4
1-1/2
16-3/4
14-1/2
21
14-1/2
21
4-1/2
3-1/4
1-1/2
1-1/2
1-1/2
1-1/2
16-3/4
24
24
5-2/3
9
6-1/2
7-1/2
168
71
22
24
17-1/2
1-1/2
6-1/4
6-1/4
6-1/4
6-1/4
69
24
64
24
24
23
4
4
69
24
6.7
5-2/3
9
6-1/2
7-1/2
5-2/3
9
6-1/2
7-1/2
3
3
2
1
1
1
37
G-l
Dried
10 g 237
125
75
60
-------
Table D-l. SELECTED LEACHING EXPERIMENTS (Continued)
Leach Solution
Cu Cr Ni
B/l B/l B/l
1:1 H SO. Leaching
(NH4)2C03 Leaching
Insert in (NH,)7CO,
Modified (NH^) CO
Insert in (NH,)-CO-
Residue Properties Extraction
Wt. '/. "1. 7. 'I. 7. •/.
gms. Cu Cr Ni Cu Cr Ni
8.1
9.9
8.3
5.0
30.2
12.9
12.0
17.8
17-4
5.0
7.0
4.8
5 3
4.0
6.0
Dilute N,SO,
30.2^ •*
9.9
6.0
7.8
7.3
ND
3.1
7 7
6 6
5.6
5,0
3.0
4.0
12.75
13.0
13.8
150
7.9
6.8
7.3
in proper sequence
7 7
3.0
Leaching
8.6
8.0
6.8
Leaching in proper sequence
7 4
6.7
6 4
7.
Wt . Remarks
67.6
60.4
67,0
50.0
69.8
36.4
40.0
11.0
13.0
50.4
30.2
51,6
47.1
61.0
40.0
69.8
1
40
21,6
26,9
ND
69.0
22.3
34.3
43 8
50.0
70.3
20.0
17.5
13.2
7.9
0
21.3
32.5
36.3
22.3
70.3
23,3
20.1
31.8
25.8
32.5
36.3
(Ammoniun) Salt Leaching)
NCI Leaching
KaOH Leaching
7.5
6.6
ND
0,2
9.7
7.9
ND
0.2
5.8
6.3
6.8
1 5
ND
1.1
25.4
36.5
ND
97.5
2.8
21.4
ND
97.5
42.0
36.6
32.0
85.0
ND
86.4
Put in proper place in (KH) CO, Leaching
9.4 J 6.3
61
-------
Although very concentrated sulfuric acid solutions were used initially,
hopefully to speed up leaching, it is advisable for economic reasons to
use a more dilute acid solution. In experiment 5, 20 w/o (weight per-
cent) sulfuric acid solution was used with good results. Ten or twenty
percent solutions would do for commercial processing.
Hydrochloric acid is very efficient, dissolving 98 percent of sludge A.
The calcium goes into solution, accounting for the increase compared to
sulfuric acid leaching. Hydrochloric acid is very corrosive and gives
rise to toxic fumes and in addition to its lack of selectivity, these
disadvantages have given rise to some hesitancy in using it as a
teachant,
Ammonium carbonate has been used as a leachant for many years in
extractive metallurgy with good results. In experiments 7 through 9,
ammonium carbonate, ammonium chloride, and ammonium sulfate were used
at 175 G. Three successive leaches were used and finally the undis-
solved residue was treated with sulfuric acid. Considering the first
two leaches, the chloride is more efficient, with the carbonate being
next. The chloride and the sulfate will tend to act similarly to
HC1 and H SO, . Calcium should remain as an insoluble carbonate in
ammonium carbonate solution but some calcium may dissolve as the bicar-
bonate. Ammonium carbonate does not take the chromium into solution.
The chromium from Sludge A settled out of the carbonate solution as a
violet precipitate. An experiment (28575-55) in which an attempt was
made to dissolve hydrated chromium trioxide (present alone) in ammonium
carbonate failed. This selectivity makes possible the separation of
one metal during leaching.
Ammonium carbonate solutions at elevated temperatures tend to lose
some ammonia. Two experiments (10 and 11) at room temperature demon-
strated that with an increase in the time leaching efficiency was about
the same. The time need not be as long as in experiment 10 (7 days).
Supplementary additions of ammonia are often used in ammonium carbonate
solution and this was tried in experiment 12. A slight increase in the
amount dissolved was effected but this could have been due to the
longer time. An increase in the ammonium carbonate concentration from
200 g/1 to 350 g/1 did result in an increase in the amount of sludge
dissolved (experiment 13). Based on experience with dried sludge,
incinerated Sludge A dissolved to about the extent expected in
ammonium carbonate solution.
The major principles demonstrated by these experiments are that the
advantage of sulfuric acid as a leaching agent is that dissolution of
copper, nickel, and chromium compounds is completed in a single treat-
ment, and that ammonium carbonate leaching has the advantage that
chromium is not dissolved, i.e., ammonium carbonate is somewhat
selective.
62
-------
Company B and Company C Sludges
It will be recalled that Companies B and C did not treat their waste-
water to obtain sludge. All waste streams ran into a common sump where
neutralization or precipitation was uncontrolled. In addition, both
sludges contained considerable organic material and some sulfides. Upon
treatment with acid, H S was given off from dried sludge samples. No
H S evolved from the incinerated sludge on treatment with acid,, In
general, the leaching with sulfuric acid or ammonium carbonate resulted
in less dissolution than with Sludge A, Data on experiments 15 through
19, Table D-l, give information on the leaching behavior of these
sludges. These sludges were not used for reclamation experiments
because more suitable sludges were available,,
Company D Sludge
The sludge from Company D originated from caustic etching of aluminum
and aluminum alloys. About 40 percent of the incinerated sludge was
soluble in 1-1 hydrochloric acid. The resulting solution was yellow
and tests for trivalent iron were positive. Recovery of aluminum was
judged to be uneconomical. The incinerated sludge could be used for
landfill, or for some application in the ceramic industry.
Company E Sludge
Two samples of sludge were taken from the lagoon at Company E, one
close to the inlet (El) and the second away from the inlet (E2). Dried
portions of these samples were leached with sulfuric acid, ammonium
carbonate, and sodium hydroxide. Neither sludge dissolved as well as
sludge A did in sulfuric acid or ammonium carbonate (experiments 23-27,
Table D-l. The reason for this is not known. The only difference in
the treatment at the plants was the complete oxidation of cyanide at
Company E. Company A used no cyanide baths.
Because Company E sludge contained zinc, an initial leach with 25 w/o
NaOH (experiments 27 and 29) was used with the intent of separating the
zinc as zincate. However, it was found that chromium dissolved as a
chromite and that some copper also dissolved. It was reported in the
literature that both chromium and copper will precipitate on boiling
the alkaline solution. Experiments showed that the precipitation of
chromium and copper was not complete. The caustic leach was eventually
63
-------
abandoned because of these drawbacks. The residue recovered was then
leached with sulfuric acid (experiments 29a and 29b, Table D-l.)
Company F Sludge
Several samples were taken at various places in the sludge storage
lagoon. Leaching results with two samples are given here. Again the
amount dissolved in sulfuric acid was less than with Sludge A. The
behavior in ammonium carbonate was similar to that of Sludge A.
Company G Sludge
This sludge dissolves fairly well in sulfuric acid at elevated temper-
atures but very slowly in ammonium carbonate (experiments 33-37).
Liquid Ion Exchange
Because results were favorable, considerable effort was given to
experiments on the reclamation of metals by solvent extraction (also
referred to as liquid-liquid extraction or liquid ion exchange).
Because of the current relatively limited application of this technique
on a commercial scale, such techniques are not commonly familiar.
Therefore, a short description follows.
In this method, an aqueous solution of two or more metal ions is
agitated vigorously with an organic solvent that is immiscible with
water. If conditions, particularly pH, are adjusted correctly, a
single metal ion will transfer to the organic phase. In this way the
extracted metal ion is selectively separated from other metal ions in
the aqueous phase. The pH required for extraction is the pH at which
the particular metal hydroxide begins to precipitate (in the aqueous
phase).
The metal ion being extracted is distributed in the two phases according
to the equation
where D is the distribution coefficient, Co is the concentration of the
metal ion in the organic phase, and C is the <
ion in the aqueous phase. The higher the valui
is the given organic solvent as an extractant.
metal ion in the organic phase, and C is the concentration of the metal
ion in the aqueous phase. The higher the value of D the more efficient
-------
After the two phases have been separated the organic phase is treated
in such a way that the metal ions are stripped from the organic medium
and the organic solvent is restored to its original condition (except
for the water that it dissolves) and can be recycled. If it is desired
to separate a second metal ion from the original aqueous phase, the pH
is changed and the aqueous solution is again agitated vigorously with
organic solvent. For each metal ion present more than one extraction
may be necessary. Many organic solvents have been investigated. It
was not possible to study all of these in connection with this project.
The first organic agent selected for study is naphthenic acid. The
naphthenic acids have the general formula:
R
where R is hydrogen or an alkyl group.
The reaction between a metal ion, e.g., copper, and naphthenic acid
may be expressed as follows :
Cu (aq.) + 2 RCOOH
extraction
stripping
Cu(RCOO>2 + 2H (aq.).
This is actually a liquid ion-exchange reaction in which hydrogen ions
are given up in exchange for copper ions.
The naphthenic acid used for the work here is a mixture of acids
rather than a single compound. Prior to extraction the naphthenic
acid was dissolved in kerosene to form a 1M solution. The general
scheme for separating and electrowinning copper and nickel is shown in
Figure D-l . The nephthenic acid solution was added to a sulfuric acid
leach solution of sludge F, whose pH had been adjusted to 5.2, the pH
for copper extraction. The agitation was done in a 500 ml separatory
funnel. Three extractions were made to get all the copper. Sludge F
was used ia this preliminary work because it had 2-6 percent copper
and practically no nickel.
(1) Chevron Chemical Company, Des Plaines, Illinois, Grade E
Naphthenic Acid.
65
-------
Dissolved in Sulfuric Acid
T
Aqueous Phase
with Copper
Electrowin
Copper
Filtrate (CD , »i , Cr )
I
Adjust pH to 5.2 with NaOH
Org.
Ph.
jse
Organic Phase
with Copper
Add H,SO
to Strip
Agif
in Nap!
Agitate with Equal Volume
in Naphthenic Acid in Kerosene
Recycle
Adjust pH to 6,0 with NaOH
Agitate with Equal Volume in
Naphthenic Acid in Kerosene
Aqueous ftiase
with Nickel
Adjust to pH
7.5 with NaOH
FIGURE D-l. Flow Diagram for Proposed Method for Separating
Copper and Nickel Using Naphthenic Acid
Agitate with Equal Volume
1M Naphthenic -Ac id in
Kerosene
i —
Aqueous
Phase
Aqueous Phase
-n *
Electrowin
Nickel
— 1
Add HjSO^ to Strip
Organic
Phase
66
-------
A series of thirteen runs were then made using leach liquors prepared
from Company A dried sludge. Table D-2 compares the materials used
in Runs 7 through 14. Runs 1 through 5 were discontinued before
completion for various reasons such as precipitate formation, excessive
emulsion problems and loss of part of one or more fractions in the runs.
Figure D-2 shows a generalized flow chart for the liquid-liquid
extraction processes used and the nomenclature and code referring to
the processing steps. The sample code used in the general form XNYM
where X is the run number, N is the process step designation (E for
extraction products, R for stripping products, etc.), Y is the
extraction step number and M is a letter A, B, C, etc., to show first
stripping, 2nd, 3rd, etc. The sample Code 7R3B, for example, means
Run Number 7, enriched stripping liquor (R) is that the raffinate from
the second (B) strip of the organic fraction from the third (3) extrac-
tion step.
In most of the runs, the pH of the leach liquor filtrate was first
adjusted to one of four ranges: below 2.5, 2.5 to 4.5 (extraction
range for Cu) , 5.0 to 7,0 (extraction range for nickel), and above
7.0. A typical run would start by adjusting the leach liquor filtrate
to 2.5 to 4.5. This would be followed by filtration if needed. One or
more extractions then were made on the adjusted filtrate with naphthenic
acid solution, until repeated successive extraction steps no longer
significantly reduced the pH of the partially depleted feedstock. The
copper was next stripped from the organic phase by treating with dilute
H SO, solution. The pH of the partially depleted feedstock then is
aajusted to the 5.0 to 7.0 range or higher, filtered if necessary, and
nickel extracted with 100 ml of approximately 1.0 M naphthenic acid in
kerosene. The naphthenic acid solution(s) was stripped with 50 ml
fractions of 0.5 M H SO, . After one or more strippings, the naphthenic
acid solution was recycled into the next extraction step. Repeated
extractions were necessary at pH 5 to 7 or higher, if much ammonia was
present, since concentrated solutions of ammonium sulfate inhibit the
nickel extraction. Extractions were terminated when a spot test of the
partially depleted feedstock for nickel showed no nickel or only a
trace present.
All runs were made using 500 ml separatory funnels as contacting vessels,
with 2 minutes of vigorous agitation (shaking) followed by settling
until two separate layers formed and stabilized. Then the denser
aqueous phase was drawn off (separated) and further processed according
to the specifications of each particular run.
Runs 7, 8, 9, 10, and 12 were made using feedstock from an (NH.) CO.,
leach of Company A dried sludge. Fifty grams of the latter were added
to 1 liter of ammonia carbonate solution composed of 300 g (NH.) C0_
in distilled water. Leaching was carried out in a 2-liter beaker at
115 to 120 F for 4 hours with stirring. The slurry settled overnight
67
-------
Table D-2. MATERIALS USED IN LIQUID-LIQUID EXTRACTION RUNS 7 THROUGH 14
Oo
Run Initial Feedstock
Number (ml) (type)(a)
7 92.5 I
8 92.5 I
9 92,5 I
10 92.5 I
11 95.0 II
12 80.0 I
13 III
14 100 IV
Acids
1:1 H S04
(mlj
21.65
None
5.00
10.0
None
12.0
0.42
2.25
or Bases Added
Ca(OH).
(R)
None
None
None
None
l,78+(c)
None
None
10.0
1 : 1 NH.OH
(ml)
2.5
None
None
None
7.9
None
19.2
9.6
Extraction Liquor
Naphthenic Stripping
Acid E in Kerosene Liquor
Maximumv"'
Metals Available
for Recovery
Acid Total Volume 0.5 M H,SO, Cu
(R) (tnl) (mir * (E)
27.5
82.5
55.0
55.0
55.0
100
100
100
100
300
200
200
200
300
300
200
460
700
500
450
1300
300
400
500
0.25
0.25
0.25
0.25
0.50
0.22
0.25
0.34
Ni
0.25
0.25
0.25
0.25
0.50
0.22
0.25
0.34
Cr
0.75
0.75
0.75
0.75
1.50
0.65
0.75
1.00
(a) I is filtrate from ammonium carbonate leach; II is filtrate from 1:1 H_SO. leach; III is an alkaline ammonium
sulfate leach; and IV is also an alkaline ammonium sulfate leach with small amounts of ammonium phosphate and
ammonium carbonate added to reduce magnesium and calcium in solution.
(b) Based on the approximate analysis of Company A sludge as reported previously (3-5% Cu, 3-57; Ni, 5-157= Cr) ,
(c) Amount used during run. More than 30 grams of Ca(OH), was used to neutralize the acid leach and adjust the
pH to 1.70 for the feedstock.
-------
ON
VO
1 1
1:1 H SO , ,
I Ca(OH,) or '
I NH4OH 1
, r--J
Leach P.
Liquor , Adjust pH
Filtrate Filter
1
Filter
Cake
1
' 0.5 M
. Strippi
Filtrate
j. (Process
Feetstoc
1
' H2S°4 J
n£ Liquor ' »
\
\
\
\
%
*T
Naphthenic ' '
Acid E in 1 1
Kerosene | >
LiquiJ-
Liquid
Extract io
^
t
Liquid-
Liquid
Stripping
\
Liq
Liq
Stri
*
f
uid-
nid
pping
f
Partially
.^.Depleted ,
Feedstock
i
XR1A
(Enriched
stripping
liquor)
— XR1B
Ca(OH)
or NH.OH
Ad jus? pH
and
Filter
I
Filter
Cake
1
J
/
(Enriched
stripping
liquor)
Recycle Naphthenic
Acid in Kerosene
1 *"
f
Partially
t£j
— i
H
r
Liquid -
Liquid
Extraction
E2
^
t
Liquid-
Liquid
Stripping
' Liquid-
Liquid
Stripping
>
Str
Org
Pha
t
Lpped
anic
se
depleted
feedstock
—)( Designated
"Raffinate"
following
last
extraction)
__^XR2A (Enriched
stripping
liquor)
_^XR2B (Enriched
stripping
liquor)
—I
FIGURE D-2. Generalized Flowshart of Liquid-Liquid Extraction and Recovery Process
-------
and was filtered the next day using a Number 2 Whatman Filter paper, a
Buchner funnel, and a vacuum filtering flask. The volume of filtrate
obtained and subsequently used for feedstock was 925 ml. Each run
except Number 12 used 92.5 ml of this filtrate. Run Number 12 was
made using 80 ml of filtrate.
One major difference in these five runs was the amount of acid added
to the feedstock before any extractions were made. (See Table D-2.)
Feedstock in Run Number 7 was acidified to pH 3.54 using 21.65 ml of
1:1 H SO with considerable evolution of CO gas. After CO evolution
stopped, 0.25 ml of 1:1 NH.OH was added to Bring the pH up to 4.43 to
promote greater copper extraction. Six extraction steps were performed
(see Figure D-3)* using 100 ml of about 1 molar Naphthenic Acid E
(27.5 g) in kerosene and recycling the naphthenic acid after stripping.
Nine stripping steps were performed yielding nine enriched stripping
liquors (see Table D-3). Sample 7R1A showed high copper and no nickel
and the raffinate showed no copper with moderate to high nickel, but
intermediate extraction samples contained various mixtures of the two.
Better separation is desirable.
Run 8 feedstock, from the same filtered leach liquor as Run 7, was
made beginning at a pH of 8.95 with no acid added. Considerable gas,
presumably CO , was evolved during extraction operations, but little
or none during stripping. This suggests the naphthenic acid was
reacting with the ammonium carbonate, producing CO . The pH of the
raffinate was 8.26 and it contained no copper or nickel, so both were
totally extracted. Fifty milliliters of about 1 molar Naphthenic Acid
E were used in each of the six extraction steps, with no recycling.
This smaller quantity was chosen to try to improve selective stripping
since selective extraction, requiring precise pH adjustment, was not
used. Attempts to obtain selective stripping from the enriched organic
fractions were generally unsuccessful (see Table D-3). None of the
enriched stripping liquors was high in either copper or nickel with
only a trace or none of the other.
Run 9 was made using 92.5 milliliters of the same leach filtrate as
Runs 7 and 8, but with the addition of 5.0 ml of 1:1 H SO, before any
extractions were performed. Considerable gas, presumably CO , was
evolved both then and during extraction steps. Two 100-milliliter
fractions of 1 molar Naphthenic Acid E were used for extractions and
were recycled alternately. Examination of the enriched stripping
liquors from Run 9 (see Table D-4) shows that the raffinate was
totally depleted of copper and nickel. Sample 9RB3 was moderate to
This figure illustrates an actual run, showing each step. Subsequent
runs followed similar patterns and figures for them have not been
included in this report.
TO
-------
100 ml 1M
Naphthenic Acid 1:1 NH^OH
in Kerosene '
Feedstock—^
™
rT
T
co,
50 ml fractions
of 0.5 molar
H2S04
JI
I
R1A
P
R2A
7R1A
Recycle
Naph. Acid
•9R2B
T
Recycle
Napho Acic
Recycle
Napho Acid
R4A
R4B
•>7R4B
Recycle
Naph. Acid
7 Raffinate
-V7R5A
Recycle
Naph. Acid'
rt R6A
7R6A
•! R6B
7R6B
FIGURE D-3. Run 7 Flow Chart
-------
Table D-3. STRIPPING LIQUOR ANALYSES
Sample
Number
16R1A
16R1B
16R2A
16R2B
16R2C
16R2D
16R2E
16 raffinate
17R1A
17R1B
17R2A
17R3A
17R3B
17R3C
17R3D
17R3E
17 raffinate
0.5(M)H2S04
Nickel
Estimate
Trace- low
None-trace
None
Very high
Very high
Low
None
Low
None
None
Moderate
None
None
Very high
Moderate-high
None
Moderate
None
(a) In a few samples, some
(b) Re check
of pH of acid
Copper
Estimate
Very high
Low
None
Trace
Trace
None
None
None
Very high
Low-moderate
None-trace
None
None
Trace
None
None
None
None
precipitate
and measured
PH
Measurement
0.73
0.53
6.53
5.62
0.55
0.43
0.44
6.67
0.71
0.46
0.55
6.70
6.24
1.10
0.52
0.48
7.20
0.42(b>
formed after
Quantity of
Precipitate
Much
None
None
Much
Some
None
None
None
Much
None
None
None
None
Much
None
None
None
None
several hours.
0.42. Earlier measurements
averaged 0.53.
-------
high in nickel with only a trace of copper and Sample 9R3B was high in
copper with only a trace of nickel. However, Samples 9R1B and 9R2B
were moderate to high in both. Since 9R3A and 9R3B were both from the
third extraction, they demonstrate that under some circumstances,
selective stripping appears to be possible, but more information is
needed.
Run 10 is the same as Run 9 except that 10.0 ml of 1:1 H SO, was added
to 92.5 ml ammonium carbonate leach filtrate to make the feedstock.
Also the fifth extraction was eliminated since the resulting stripping
liquor in Run 9 (9R5A) was low in both nickel and copper. Two 100-ml
fractions of about 1 molar Naphthenic Acid E were used and recycled
alternately as in Run 9. None of the enriched stripping liquors was
high in copper with only a trace or no nickel (see Table D-4) , but two
samples, 10R1C and 10R2A, were high in nickel with very little or no
copper. Again, several samples contained both. Also, it should be
noted that Sample 8R1A, 8R2A, 9R1A, 9R2A, and 10R1A show little or no
copper and relatively little nickel, although the pH for each stripping
liquor is above 7.0, indicating that the sulfuric acid has been
exchanged for something since the 0.5 molar H SO, used in all the
stripping operations shows a pH of about 0.53. The material extracted
in these samples has not as yet been identified.
Run 11 was made using 95 ml of filtrate from a sulfuric acid leach of
Company A dried sludge. The acid leach was made using 1.00 liter of
1:1 H-SO, (by volume) and 100 g of Company A dried sludge. This was
stirred for 2 hours at 150 F, allowed to cool and settle, and filtered
through a fritted-glass Buchner funnel. The total filtrate was 950 ml.
The only problem was that of considerable froth formation during initial
mixing of acid and sludge, but the froth was broken by overnight
settling, prior to heating.
The 95 ml of acid leach filtrate used in Run 11 was first raised to a
pH of 1.7 by adding Ca(OH) and filtering. The considerable mass of
precipitate required several rinses to recover the soluble metal values.
This dilute filtrate and rinse was concentrated by evaporation under
vacuum. The concentrated liquor was diluted to 95 ml and the pH found
to be 1.71. Sixteen consecutive extractions were performed using two
100-ml batches of about 1 molar Naphthenic Acid E, alternately, and 26
stripping steps were performed (see Table D-4). Samples 11R4A through
11R7B contained nearly all the copper recovered and relatively little
of the nickel. Samples 11R8A through 11R16A contained most of the
nickel and almost no copper. The total amount of thes:e is presumably
greater than in previous runs since the feedstock probably contained
about twice as much of nickel and copper feedstock as for the other
runs described. Nearly all of the copper and nickel was extracted from
the feedstock as shown by the fact that the raffinate showed no copper
T3
-------
Table D-4. ROUGH QUANTITATIVE COMPARISON OF
ENRICHED STRIPPING LIQUORS
DH(a) of
Partly Extracted Enriched Stripping
Feedstock
Sample
Number
7R1A
7R2A
7R2B
7R3A
7R4A
7R4B
7R5A
7R6A
7R6B
7 raffinate
8R1A
8R1B
8R1C
8R2A
8R2B
8R2C
8R3A
8R3B
8R3C
8R4A
8R4B
8R5A
8R5B
8R6A
8 raffinate
9R1A
9R1B
9R1C
9R2A
9R2B
9R2C
9R3A
9R3B
9R4A
9R5A
9 raffinate
Before
Extraction
4.43
5.03
5.03
5.39
_-
5.71
6.95
-_
5.81
8.95
--
--
N.M.
--
__
N.M.
--
__
N.M.
-_
N.M.
-_
N.M.
8.26
8.23
--
--
N.M.
--
—
N.M.
—
N.M.
N.M.
7.51
After
Extraction
4.90
3.08
3.32
4.41
--
4.83
5.81
--
--
N.M.
--
--
N.M.
--
--
N.M.
--
--
N.M.
--
N.M.
—
8.26
--
N.M.
--
--
N.M.
--
—
N.M.
--
N.M.
7.51
--
PH
0.70
0.66
0.61
0.57
0.59
0.56
0.56
0.53
0.53
__
7.89
0.92
0.70
7.53
0.74
0.70
4.60
0.57
0.69
1.65
0.71
1.34
0.79
1.37
--
7.63
1.18
0.61
7.13
0.63
0.62
4.65
0.58
1.42
1.08
--
Liquor
Cu
High
Moderate
None
Low-mod .
Moderate
None
Low
None
None
(None)
Trace
None
None
Trace
Moderate
None
High
Moderate
None
Low -mod.
None
Low -mod.
None
Low-mod .
(None)
None
Mod . -high
None
None
Mod . -high
None
Trace
High
Low
Low
(None)
Ni
None
Trace
Trace
None
None
Trace
Moderate
Moderate
Low
(Low-mod.)
Low
Low
Low
Trace
Low
Trace
Low-mod .
None
Trace
Low-mod .
Low
Low -mod .
Trace
Low -mod .
(None)
Trace
Mod . -high
Low
Low
Mod . -high
Low -mod .
Mod . -high
Trace
Low
Low
(Rone)
-------
Table D-4. ROUGH QUANTITATIVE COMPARISON OF
ENRICHED STRIPPING LIQUORS (Continued)
pH('
of Partly Extracted Enriched Stripping
Feedstock
Sample Before
Number Extraction
10R1A
10R1B
10R1C
10R2A
10R2B
10R2C
10R3A
10R3B
10R4A
10 raffinate
11R1A
11R1B
11R2A
11R2B
11R3A
11R3B
11R4A
11R4B
11R5A
11R5B
11R6A
11R6B
11R7A
11R7B
11R8A
11R9A
11R9B
11R10A
11R11A
11R11B
11R12A
11R13A
11R14A
11R14B
11R15A
11R16A
11 raffinate
7.04
--
--
N.M.
—
—
N.M.
--
N.M.
7.55
1.71
—
2.01
--
3.20
--
3.82
--
3,50
--
4.18
—
4.81
—
4.87
5.65
5.03
7.35
--
5.78
N.M.
7.63
—
5.83
9.23
6.50
After
Extraction
N.M.
--
--
N.M.
--
--
N.M.
--
7.55
--
1.71
--
2.01
--
3.10
--
3.80
--
3.33
--
3.86
-_
3.70
--
4.28
3.50
4.20
5.78
--
4.32
4.80
5.83
--
N.M.
6.50
--
PH
7.03
0.81
0.65
3.64
0.64
0.60
1.20
0.63
0.96
--
0.57
0.60
0.53
0.57
0.56
0.53
0.55
0.59
0.55
0.54
0.72
0.58
0.58
0.57
0.58
0.58
0.58
0.61
0.85
0.59
0.61
0.70
0.83
0.63
0.74
0.85
--
Liquor
Cu
None
High
None
Very low
Low-mod .
Trace
Low- mod .
None
None
(None)
None
None
None
None
Low
None
Mod . -high
None
Low- mod .
None
V. High
Low
Mod. -high
None
Trace
Low-mod .
None
None
Trace
None
None
Trace
None
None
None
None
(None)
Ni
Moderate
High
High
High
Low
None
High
None
Low -mod .
(None)
Mod. -high
None
Mod. -high
Trace
Mod. -high
None
Low-mod.
None
Low-mod .
None
Trace
Trace
Low-mod .
Trace
Low-mod .
Trace
Trace
Low -mod .
High
Mod.
Mod.
High
High
Mod.
Trace
High
(Trace)
75
-------
Table D-4. ROUGH QUANTITATIVE COMPARISON OF
ENRICHED STRIPPING LIQUORS (Continued)
PH(
of Partly
Extracted
Enriched Stripping
Feedstock
Sample
Number
12R1A
12R1B
12R1C
12R1D
12R1E
12R2A
12 raffinate
13R1A
13R2A
13R3A
13R4A
13R5A
13R6A
13R6B
13R6C
13 raffinate
14R1A
14R2A
14R3A
14R3B
14R4A
14R4B
14R5A
14R5B
14R5C
14R5D
14 raffinate
Before
Extraction
6.68
—
--
--
—
N.M.
7.02
5.27
5.60
5.72
6.33
8.02
8.50
—
--
7.02
3.53
4.96
6.43
—
7.95
--
9.63
—
—
--
7.23
After
Extraction
N.M.
--
—
—
--
7.02
--
2.75
3.58
3.98
3.13
3.36
7.02
--
--
--
3.70
3.23
5.72
--
3.60
—
7.23
—
--
—
PH
7.06
1.60
0.67
0.67
0.71
1.01
--
0.70
0.72
0.77
0.72
1.03
6.87
1.37
0.73
_-
0.75
0.78
0.70
1.02
0.78
6.97
6.39
1.18
0.71
Liquor
Cu
Trace
V. High
Low-mod .
Trace
None
None
(None)
None
None
Mod . -High
Mod . -High
Low-Mod .
None
Trace
None
(None)
Low
High
Mod.
Trace
Low-mod .
Trace
None
None
Low
None
(None)
Ni
Low-mod .
V. High
High
Low -mod.
None
Low-mod .
(Low)
Trace
Low
Trace
None
Mod . -High
None
High
Trace
(Low)
Low
Trace
Low -Mod .
Trace
V. High
Mod . -High
None
Low
High
Mod.
(Trace)
(a) N.M. means not measured.
76
-------
and only a trace of nickel. Problems with precipitation were
encountered around pH 5-6 and required filtering after each pH adjust-
ment step after the 7th extraction. The precipitate probably is
mostly chromium compounds. A small amount of this precipitate formation
was found in the runs on ammonium carbonate leach; also, in even small
amounts of precipitate tend to interfere with separation of the
aqueous and organic fractions. Run 11 yielded several usable samples,
but some improvements in separation may still be desirable.
Runs 13 and 14 were made using alkaline ammonium sulfate leach of
Company A dried sludge. This was selected with hopes of recycling the
(NH,)_SO, and to reduce process material costs for acids and bases,
which, according to Fletcher and Flett, constitute a significant
fraction of the total costs involved.
Run 12 was made using only 80 ml of the ammonium carbonate leach
filtrate and 12 ml of 1:1 H SO, (equivalent to using 15.4 ml of acid
with 92.5 ml of filtrate). Only two extraction steps were made using
300 ml of about 1.2 molar Naththenic Acid E in kerosene and recycling.
This was done to reduce the variables pertaining to selective stripping,
with the objective of extracting all the nickel and copper in one
fraction of enriched organic phase, then selectively stripping. This
was not successful (see Table D-4). None of the enriched stripping
liquor samples was high in copper or nickel with only a trace or none
of the other. The controlling factors for selective stripping have
still not been determined, through pH and stoichiometric balance seem
to be important.
For Run 13, 5.0 g of Company A dried sludge was stirred overnight with
80 ml distilled water and about 20-30 grams of (NH ) SO, with one drop
of 1:1 NH OH. This was filtered and the filtrate (pH 7.00) was
acidified to pH 3.90 using 0.42 ml of 1:1 H SO,. The organic extract
was composed of 100 g Naphthenic Acid E dilated to 300 ml with kerosene
(about 1.2 molar) which was recycled through a total of six estractions.
Eight enriched stripping liquors were produced (see Table D-4) with
good separation of nickel and copper in all samples except 13R5A.
Run 14 was similar to Run 13. The leach liquor was produced by
stirring overnight 10.0 g of Company A dried sludge with 90 ml distilled
water, 50 g ammonium sulfate, 2 g ammonium carbonate, 1 g dibasic
ammonium phosphate, and 3 drops of 1:1 NH.OH. This was filtered and
the filtrate diluted to 150 ml. Of this, 100 ml were used in Run 14.
The initial pH of 8.20 was reduced to 3.53 using 1.65 ml of 1:1 H SO, .
Five extractions were performed with the pH raised between each by
addition of 1:1 NH.OH or Ca(OH)9. The organic phase was composed of
100 g Naphthenic Acid E diluted-to 200 ml with kerosene (about 1.8
molar). This higher molarity was chosen to see if phase separation
problems and/or viscosity problems would result. They did not, and,
77
-------
in fact, the separation of organic and aqueous fractions in Run 14 was
better than the other runs, but this is probably due more to the
apparent total lack of chromium compounds to produce precipitates,
more than the change in molarity of tnaphthenic acid used. The nickel
and copper separation obtained was actually not quite so good as in
Run 13, since both 14R3A and 14R4A have moderate amounts of both.
Runs 7, 11, 13, and 14 are the best of the series. Each had some
problems, but it is anticipated that application of Fletcher and
Flett's suggestions from the source previously cited will further
improve separations of Cu and Ni. The various leaching liquors used
in these four runs support the feasibility of more than one approach.
The examples of selective stripping, especially in Run 9, suggest this
may perhaps be used as a refinement in conjunction with selective
extraction, should it be required (after further study to isolate the
controlling variables). Numerous samples have been produced that seem
amenable to electrowinning of copper and nickel separately. The
chromium may be recovered as a precipitated compound in the case of the
acid extraction, or recovered from the cake in the case of extraction
in ammoniacal systems.
Run 16 feedstock was prepared from a slightly alkaline ammonium sulfate-
ammonium hydroxide leach of Company A dried sludge. Two extractions
were performed using Naphthenic Acid E in kerosene. The first extrac-
tion (for nickel) was adjusted to a pH of 5.4 and the second (for
copper) to a pH of 6.7. Table D-3 shows the estimated levels of nickel
and copper in the various stripping liquor samples. The two extractions
seem to give good separation of nickel and copper. A small amount of
nickel was still present in the raffinate.
Run 17 was performed using a feedstock prepared from an acidic
ammonium sulfate-sulfuric acid leach of Company A dried solids. Three
extractions were made with Naphthenic Acid E in kerosene at final pH
values of 5.3, 5.2, and 7.1. Table D-3 shows the results. There is
more nickel in the Run 17 raffinate than in that of Run 16 because
NH.OH was used to adjust the pH in Run 17. This verifies the earlier
finding that the use of some Ca(OH) to adjust pH in Run 6 precipitates
calcium sulfate and seems to permit more complete recovery of nickel.
The results of these last two runs indicate that two extractions, one
for copper and a second for nickel, are probably adequate for separ-
ation and recovery of these two metals from Company A sludge if the
pH is carefully selected and adjusted and if the level of ammonium
sulfate is not so high that it interferes with the nickel recovery.
78
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APPENDIX E
ECONOMIC STUDIES
In view of all the considerations and physical variables identified in
this study, it is clear that a general "sludge treatment process" is
beyond consideration at this point. However, to gain some idea of
the economics that might be encountered in such a process, estimates
were made to determine the capital and operating costs for a small
batch type plant. In designing this process, the order of priorities
in process selection was:
(1) detoxification of wastes
(2) simplicity of process
(3) the generation of products amenable to
consumption in current metallurgical processes
and waste products exhibiting the maximum
practical chemical stability.
The plant design and cost estimates were based on batch operation,
using the following design basis:
Plant Capacity: 100 Ib dry solids/12-hour day/batch
Daily Sludge Feed: 3,330 Ib total (3 percent solids)
3,230 Ib water
100 Ib solids
5 Ib Cu
0.3 Ib Ni
10 Ib Cr
1 Ib Fe
0.6 Ib Zn
10 Ib Ca
A flowsheet of the sludge-treatment process is shown in Figure E-l.
A sludge feed (3 percent solids) is preconcentrated by filtration to
a solids concentration of 15 percent. The concentrated sludge is then
dried in a pan dryer to 41 percent solids. The latter concentration
was arrived by assuming that acid leaching is to be carried out with
30 gallons of 20 percent sulfuric acid on 100 Ib of dry solids.
70
-------
SLUDGE
I OPTIONAL I
' NEUTRALIZATION |
FOR ACID OR |
ALKALINE SLUDGES I
I I
H SO .'. - t
"24 *
Cake
* (CaSO.)
Na2C03(1)
Sale Cake
or ^-_. Cr(OH)
Waste Fe(OH)^
Cu, Ni
Cake (Residual
FILTER
DRYER
LEACH
FILTER
PRECIPITATION
FILTER
ELECTROLYSIS
TTNTT
NEUTRALIZATION
FILTER
water ^
,
I
water
f
To Dump
Figure E-l. Flowsheet of proposed sludge treatment pilot process
(1) The efficiency of this separation of iron and chromium from copper
and nickel remains to be tested. On carbonate solutions it is
easily possible to separate chromium from Fe and Ni. Data indicate
that in a carefully controlled system of carbonate precipitation it
might be possible to effect a reasonably good separation of Fe
from copper„ Fe is reported to precipitate at near pH 2.5; copper
at pH 5 to 5.5. The available data on carbonate precipitation does
not show the behavior of Cr
80
-------
In the leaching step, 58 Ib of 98 percent sulfuric acid is added to the
dried sludge. The quantity of acid is 1.43 times (43 percent excess)
the stoichiometric requirement for neutralization of metal hydroxides
(excluding Ca) present in the sludge.
The slurry from the leaching step is filtered to remove solids (calcium
sulfate). The filtrate is then neutralized with soda ash to pH 3 to
precipitate out Cr(OH)., and Fe(OH)_. The latter oxides are removed by
filtration, and the filtrate is fea into an electrolysis unit removal
and recovery of metallic copper and nickel.
After electrolysis is completed, lime is added to the electrolysis
solution to neutralize the acid formed from electrolysis. Calcium
sulfate and zinc hydroxide precipitate out as the result of neutral-
ization. The lime-treated liquor is filtered in the final step of
the process to remove the precipitates.
Valuable materials recovered from the process include: 5 Ib of Cu,
0.2 Ib of Ni, and 20 Ib of Cr(OH)3 (mixed with 2 Ib of Fe(OH)3>.
The batch operation could probably be accomplished over a 12-hour
period according to the following time schedule:
Operation Duration, hr
Drying 2.0
Leaching 1.0
Precipitation 0.3
Electrolysis 4.0
Neutralization 0.2
Filtration 2.0
Materials handling 1.5
Start-up and Shut-down 1.0
Total 12.0
A list of equipment and estimated costs is given in Table E-l. The
total direct plant cost, exclusive of a trailer for mounting the plant,
was estimated as about $15,000. Estimates of operating costs are shown
in Table E-2. The daily operating cost was estimated to be about $61
per day. The labor cost is the major cost item.
The plant capacity can be doubled to treat 200 Ib/day of sludge on
dry-solids basis by operating the plant 24 hours a day.
81
-------
Table E-l. EQUIPMENT COST SUMMARY
Item
Electrolysis Unit
Pan Dryer, 304
Capacity
1.3 Ib Cu+Ci/hr
50 ft2, 10 h.p.
Number
1
1
fa\
Purchased v Cost,
Dollars (1972)
3,500
3,430
stainless
Pressure Filter
Leaching/Neutrali-
zation Tank, rubber
lined
1 ft"
100 gal
Precipitation Tank, 100 gal
rubber-lined
Agitators, 304
stainless
Pump, magnetic-
coupled
1 h.p.
13 gpm @ 6' HO
1/4 h.p. Z
Purchased Equipment Cost (E)
Equipment Installation, 40 percent of E
Piping (hoses), estimate
Instrumentation, estimate
Total Direct Plant Cost
600
200
200
1,280
100
$9,310
3,720
100
2,000
$15,130
(a) Cost data obtained from: (1) Peters, M. S., and Timmerhaus, K. D.,
"Plant Design and Economics for Chemical Engineers", McGraw-Hill
(1968), (2) Laney, L. E., and Forbes, C. A., "Brass Wire Mill
Process Changes and Waste Abatement, Recovery and Reuse", Water
Pollution Control Research Series, 12010 DPF 11/71 (1971).
82
-------
Table E-2. OPERATING COST SUMMARY
Daily Cost,
Basis: 200 days/year Operation do liars/day
Materials
H2S04, 98% @ 1.7c/lb 0.99
Na2C03 @ 2.5
-------
APPENDIX F
BIBLIOGRAPHY
(1) A State-of-the-Art Review of Metal-Finishing Waste Treatment,
prepared by Battelie-Columbus Laboratories, for the FWQA and
the Metal Finishers' Foundation, November, 1968.
(2) Selected Water Resources Abstracts, Water Resources Scientific
Information Center, Office of Water Resources Research, U. S.
Department of Interior.
(3) Pollution Abstracts, 6811 La Jolla Blvd., La Jolla, California
92037.
(4) Water Pollution Abstracts, Dept. of the Environment, Her Majesty's
Stationery Office, London.
(5) Grune, W. N. Plating and Cyanide Wastes. J. Water Pollution
Control Federation, 43_:1024, 1971. (Note: This is a yearly
feature in the journal.)
(6) Foulke, D. Gardner. Bibliobraphy Metal-Finishing Wastes.
Plating. 51:289, 1970.
(7) Kirk-Othmer, Encyclopedia of Chemical Technology, John Wiley &
Sons, 1963, 2nd Edition, 22 vols.
(8) Mellor, J. W. A Comprehensive Treatise on Inorganic and
Theoretical Chemistry. (London). Longmans Green & Co., Ltd.,
1927. 16 vols.
(9) tFriend, J. Newton, Editor. A Testbook of Inorganic Chemistry.
(London) Charles Griffin & Co., Ltd., 1927. 11 vols.
(10) Gmelins Handbook der Anorganischem Chemie Verlag Chemie, G.M.B.H.
(Gmelins Handbook of Inorganic Chemistry, Published by Chemie
Corporation). (Berlin) 8th Edition (Still being published).
(11) Linke, William F. Weidell's Solubilities of Inorganic and Metal
Organic Compounds. ACS (Washington, D.C.). 1958. 2nd Edition.
(12) Brasted, Robert C. Comprehensive Inorganic Chemistry. D. Van
Nostrand, 1961. 11 vols. (Vol. VIII).
(13) Ephram, Fritz. Inorganic Chemistry. Interscience, 1946.
4th Edition Revised.
8U
-------
(14) Sidgwick, N. V. The Chemical Elements and Their Compounds.
(Oxford) 1950. 2 vols.
(15) Cotton & Wilkinson. Advanced Inorganic Chemistry. Interscience,
1962.
(16) McAlpine and Soule. Prescott and Johnson Qualitative Chemical
Analysis. D. Van Nostrand, 1933.
(17) Latimer and Hildebrand. Reference Book of Inorganic Chemistry.
MacMillan, 1951. 3rd Edition.
(18) Mantell, C. L. Industrial Electrochemistry. McGraw-Hill, 1950.
3rd Edition.
(19) Hampel, Clifford A., Editor. The Encyclopedia of Electrochemistry.
Reinhold, 1964.
(20) Lowenheim, Fred A., Editor. Modern Electroplating. John Wiley
& Sons, 1963. 2nd Edition.
(21) Allmand, H. J. and H. J. T. Ellingham. The Principles of Applied
Electrochemistry. Longmans, Green & Co., 1924.
(22) Glasstone, S. Electrolytic Oxidation and Reduction. D. Van
Nostrand Co., 1936.
(23) Sand, Henry J. S. Electrochemistry and Electrochemical Analysis.
Blackee & Son Ltd., 1941. Vols. I, II and III.
(24) Perry, Robert H., Cecil H. Chilton and Sidney D. Kirkpatrick.
Chemical Engineers Handbook. McGraw Hill, 1963. 4th Edition.
(25) Pepperl, H. Incineration of Electroplating Sludges by Means of A
Fusion Cone. (In German - Translation on file at Battelie-Columbus
Laboratories). Brennst.-Warme-Kraft. 1_6_:399, 1964.
(26) Rehm, F. and S. R. Nietz. Critical Study of Effluent Sludge
Disposal Through Low Temperature Incineration. (In German -
Translation on file at Battelle-Columbus Laboratories.) Galvano-
technik. 56:462, 1965.
(27) Weiner, Robert F. Acute Problems in Effluent Treatment. Plating.
54:1354, 1967.
(28) Niemetz, W. Disposal of Industrial Sludges. (Presented at the
4th International Congress of the International Research Group
on Refuse Disposal (Basel). Schweiz. Z. fur Hydrologie.
June 2-5, 1969.) 31,:425, 1969.
-------
(29) McGrath, J. J. Treatment of Brass Mill Effluents at Anaconda
Toronto Plant. (Presented at the 16th Ontario Ind. Waste Conf
June 15-18, 1969.) 82 p.
(30) Jones, J. L. Electrolytic Recovery of Copper and Sulfuric Acid
from Copper Mill Pickling Solutions. Trans. Amer. Electrochem.
Soc. 32.: 229, 1917.
(31) Whistance, D. J. and E. C. Mantle. Effluent Treatment in the
Copper and Copper Alloy Industries. The British Non-Ferrous
Metals Research Assn. January, 1965.
(32) Sierp, F. Electrolytische Verfahren, Handbuch der Metallbeizerei
(Nichteisenmetalle) (Electrolytic Methods. Handbook of Metal
Pickling (Non-Ferrous Metals). Verlag-Chemie (Published by Chemie
Corporation). 1951. 100 p. (In German)
(33) Boer, P. Die Ruckgewinnung von Kupfer aus Beizbadern (The Recovery
of Copper from Pickling Bath). Metall. 14_:1107, 1960. (In
German).
(34) Keetman, E. Modern Pickling and Yellow Pickling Baths. Wire and
Wire Products. 34:1085, 1959.
(35) Anon. Recovery of Copper and Sulfuric Acid from Pickling Baths
Electroplating. 5_:227, 1952.
(36) Anon. Copper Recovery from Pickling Solutions. Metal Industry.
8^:370, 1952. y
(37) Muller, Richard L. Pollution in the Non-Ferrous Metals Industry-
Metal Plating and Electroplating Industries. Waste Disposal
Problems in Selected Industries. Edited by John E. Ullmann,
Hofstra University Yearbook of Business, Series 6, Vol. 1, 1969.
(38) Wise, William S. The Industrial Waste Problem-IV Brass, Copper,
and Electroplating Wastes. Sewage Works J. 20_:96, 1948.
(39) Wise, William S. and B. F. Dodge. Industrial Wastes—Brass and
Copper Industry. Ind. Eng. Chem. 3_9:632, 1947.
(40) McKee, Ralph H. and S. T. Leo. A Continuous Process for Electro-
lytic Regeneration of Chromic Acid. J. of Industrial and Ene.
Chem. 12.-.16, 1920.
(41) Mitter, G. C. and S. G. Dighe. J. Sci. Ind. Res (India).
1:11, 1943.
86
-------
(42) Barnard, Paul G., et al. Recovery of Metallurgical Values from
Industrial Wastes. (Presented at Proceedings of the Second
Mineral Waste Utilization Symposium, U. S. Bureau of Mines and
IIT Research Institute. Chicago, March 18-19, 1970.)
(43) George, L. C. and Andrew A. Cochran. Recovery of Metals from
Electroplating Wastes by the Waste-Plus-Waste Method, Bureau of
Mines Solid Waste Research Program, Technical Progress Report
27. August, 1970.
(44) George, L. C., et al. The Waste-Plus-Waste Method for Recovering
Metals from Electroplating Wastes. (Presented at the A.E.S.
Waste Treatment Conference, Post Junior College, Waterbury, Conn.
May 21-22, 1974.)
(45) Potter, G. M. Recovery of Nonferrous and Precious Metals from
Electrical and Electronic Scrap. (Presented at the Effective
Technology and Research for Scrap Metal Recycling Symposium,
conducted by NASMI and the U. S. Department of Interior, Bureau
of Mines, Washington, D.C. January 7, 1971.)
(46) Brooks, P.T., et al. Chemical Reclaiming of Superalloy Scrap.
Bureau of Mines RI 7316. November, 1969.
(47) Brooks, P.T., et al. Processing of Superalloy Scrap. J. Metals.
22_:25, 1970.
(47a) Anon. New Recovery Process Can Yield Both Electrolytic Nickel
and Copper. Engineering and Mining J. January, 1971. 94 p.
(48) Van Arsdale, George D., Editor. Hydrometallurgy of Metals.
McGraw Hill, 1953. 1st Edition. Cementation, p. 188; Leaching,
pp. 11-29 and 70-102.
(49) Burkin, A. R. The Chemistry of Hydrometallurgical Processes.
E. & F. N. SPON, Ltd. (London), 1966.
(50) Habashi, Fathi. Principles of Extractive Metallurgy - Vol. 2.
Hydrometallurgy. Gordon and Breach, 1969. pp. 227-245.
(51) Smithson, G. R., Jr. et al. Liquid-Liquid Extraction of Nonferrous
Metals-Review of the Art. J. of Metals. JJ3;1037, 1966.
(52) Habashi, Fathi. Principles of Extractive Metallurgy-Vol. ,2.
Hydrometallurgy. Gordon and Breach, 1969. Chap. 17, Solvent
Extraction.
(53) Green, H. Separations Using Solvent Extraction. Metallurgia.
7jO:143, 1964.
87
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(54) Zakarias, M. J. and M. J. Cahalan. Solvent Extraction for Metal
Recovery. Trans. Inst. Min. Met. 75_:C245, 1966.
(55) Fletcher, A. W. and J. C. Wilson. Naphthenic Acid as a Liquid-
Liquid Extraction Reagent for Metals. Trans. Institution of
Mining and Metallurgy. 7£:355, 1960-61.
(56) Fletcher, A. W. and K. D. Hester. A New Approach to Copper-Nickel
Ore Processing. Trans. Soc. Mining Engrs. 229:282, 1964.
(57) Fletcher, A. W. and D. S. Flett. Carboxylic Acids as Reagents for
the Solvent Extraction of Metals. Proc. of the Int. Conf. on
Solvent Extraction Chemistry of Metals, Edited by McKay, H. A. C.,
et al., CRC Press, Cleveland, Ohio, September 27-30, 1965.
(58) Fletcher, A. W., et al. Separation of Zinc and Cadmium by Solvent
Extraction. Advances on Extractive Metallurgy, Proc. of Symp. by
Institution of Mining and Metallurgy (London), Elsevier Publishing
Company, April 17-20, 1967.
(59) Baggott, E. R. et al. Recovery of Valuable Metals from Nickel-
Cobalt Alloy Scrap. Mineral Processing and Extractive Metallurgy.
Proc. of Ninth Commonwealth Mining and Metallurgical Congress.
Edited by M. J. Jones. 1969, Vol. 3.
(60) Fletcher, A. W. Metal Extraction from Waste Materials. Chemistry
and Industry. 1971. p. 776.
-------
TECHNICAL REPORT DATA
(Please read Instructions on the reverse before completing}
1. REPORT NO.
EPA-670/2-75-018
3. RECIPIENT'S ACCESSIO[*NO.
4. TITLE AMDSUBTiTLE
RECLAMATION OF METAL VALUES FROM METAL-FINISHING
WASTE TREATMENT SLUDGES
5. REPORT DATR
April 1975;
Issuing Date
6. PERFORMING ORGANIZATION CODE
7. AUTHOR(S)
Arch B. Tripler, Jr., R. H. Cherry, Jr., and
G. Ray Smithson, Jr.
8. PERFORMING ORGANIZATION REPORT NO
9.PERFORMING ORGANIZATION NAME AND ADDRESS
Battelle Memorial Institute
Columbus Laboratories
505 King Avenue
Columbus, Ohio U3201
10. PROGRAM ELEMENT NO.
1BB036;ROAP 21 AZO;Task
11.SBRXKAKWGRANT NO.
12010 FXD
12. SPONSORING AGENCY NAME AND ADDRESS
National Environmental Research Center
Office of Research and Development
U.S. Environmental Protection Agency
Cincinnati, Ohio ^5268
13. TYPE OF REPORT AND PERIOD COVERED
Final
14. SPONSORING AGENCY CODE
15. SUPPLEMENTARY NOTES
16. ABSTRACT " ~~ ~~~~ ~~
The efforts of this program have included the determination of the vorth of recovering
metal values from metal-finishers' vastevater treatment sludges, the definition and
research of processes for such recovery, and the selection, design, and costing of a
recommended process. The study included a survey of the literature to determine the
state-of-the-art regarding the generation, disposal, and recovery treatment practices
relevant to metal-finishers' sludges, and to identify metal recovery processes possi-
bly applicable to those sludges. This information was supplemented with a survey "by
questionnaire to determine the current status of relevant practices and conditions.
Field investigations provided detailed examples of plant practices, sludge storage
conditions, and sludge characteristics. The extraction of metal values from waste
sludges by various leaching agents, and the recovery of metal values by techniques
including electrowinning, cementation, and liquid-liquid ion exchange were studied.
A portable pilot process for the treatment of waste sludges and recovery of metal
values was selected and equipment and operating costs developed.
KEY WORDS AND DOCUMENT ANALYSIS
DESCRIPTORS
b.IDENTIFIERS/OPEN ENDED TERMS
c. COSATI Field/Group
Industrial wastes, Waste water, Waste
treatment, *Sludge, *Metal finishing,
'Materials recovery, Electrowinning,
Ion exchanging, Sludge disposal
*Waste recovery
13B
8. DISTRIBUTION STATEMENT
RELEASE TO PUBLIC
19. SECURITY CLASS (This Report}
UNCLASSIFIED
21. NO. OF PAGES
91
20. SECURITY CLASS (This page)
UNCLASSIFIED
22. PRICE
EPA Form 2220-1 (9-73)
.S. GOVERNMENT PRINTING OFFICE.- 1975-657-592/5361 Region No. 5-11
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