U.S. patent number 3,917,519 [Application Number 05/446,314] was granted by the patent office on 1975-11-04 for process for the manufacture of electrolytic copper.
This patent grant is currently assigned to Freeport Minerals Company. Invention is credited to Bernard M. Fisher, Robert C. Hills, Freddie J. Touro.
United States Patent |
3,917,519 |
Fisher , et al. |
November 4, 1975 |
Process for the manufacture of electrolytic copper
Abstract
A process is described for making electrolytic copper of high
quality from chalcopyrite flotation concentrates, and the like. The
concentrates are first slurried and then subjected to a high
temperature-high pressure oxidation-leaching operation in an
autoclave system under controlled conditions. The resulting slurry
from the oxidation-leaching operation is flashed to atmospheric
pressure in order to remove heat and then subjected to a
solids-liquid separation. The separated liquor from the
solids-liquid separation system, containing the recovered copper
values, is cooled and then fed to an electrolytic deposition
operation where metallic copper cathodes of high purity are
produced. The spent electrolyte is further subjected to a sulfide
treatment to precipitate and recover, as copper sulfide, the
residual amounts of copper not deposited in the electrolytic cells.
The copper sulfide is then recycled in the form of a slurry to the
initial stages of the process.
Inventors: |
Fisher; Bernard M. (New
Orleans, LA), Hills; Robert C. (Metairie, LA), Touro;
Freddie J. (New Orleans, LA) |
Assignee: |
Freeport Minerals Company (New
York, NY)
|
Family
ID: |
23772125 |
Appl.
No.: |
05/446,314 |
Filed: |
February 27, 1974 |
Current U.S.
Class: |
205/584; 423/35;
423/41; 423/87; 423/158; 423/555; 423/633; 423/37; 423/140;
423/522; 423/602 |
Current CPC
Class: |
C22B
15/0086 (20130101); C22B 15/0089 (20130101); C22B
15/0071 (20130101); Y02P 10/20 (20151101) |
Current International
Class: |
C22B
15/00 (20060101); C22d 001/16 (); C22b
015/00 () |
Field of
Search: |
;75/117
;423/140,158,633,602,37,87,522,555,41 ;204/108 |
References Cited
[Referenced By]
U.S. Patent Documents
Primary Examiner: Andrews; R. L.
Attorney, Agent or Firm: Flynn; Lawrence W.
Claims
What is claimed is:
1. Process for recovering copper from a copper-bearing sulfide ore
containing iron comprising the steps of:
passing an aqueous slurry of said ore through an oxidation-leaching
operation in an autoclave system while maintaining the temperature
within said autoclave system at between 350.degree.F and
450.degree.F and providing an oxygen partial pressure within said
autoclave system of between about 50 and 500 pounds per square inch
to form sulfuric acid and to solubilize the metal constituents in
said slurry as sulfate solutions,
reducing the acidity of said solutions in said autoclave system by
the addition of a neutralizing agent during said oxidation-leaching
operation to precipitate substantially all of the solubilized iron
in said solutions as insoluble ferric oxide,
withdrawing from said autoclave system the treated slurry
comprising a liquid phase of acidic copper sulfate and solids
comprising ore gangue and the precipitated ferric oxide,
flashing the treated slurry to atmospheric pressure,
separating said liquid phase from said solids,
washing said separated solids with an aqueous wash medium to
recover retained copper sulfate solution and recycling said
recovered copper sulfate solution to said autoclave system,
cooling said separated liquid phase,
electrolytically treating said separated liquid phase to produce
electrolytic copper and a spent electrolyte comprising sulfuric
acid and a residual amount of copper,
treating said spent electrolyte with a sulfiding agent to
precipitate substantially all of said residual amount of copper
from said electrolyte as copper sulfide,
thickening said treated spent electrolyte to separate said
precipitated copper sulfide from said treated spent electrolyte as
a thickened copper sulfide slurry, and
recycling said thickened copper sulfide slurry to said autoclave
system.
2. The process as defined in claim 1 wherein water vapor produced
during said oxidation-leaching operation within said autoclave
system is removed from said autoclave system whereby the
concentration of said sulfate solutions is increased.
3. The process as defined in claim 1 wherein the acidity of said
solutions in said autoclave system is reduced to less than 40 grams
per liter of sulfuric acid by the addition of said neutralizing
agent.
4. The process as defined in claim 1 further comprising the step of
cooling the treated slurry withdrawn from said autoclave system to
between about 275.degree.F and 400.degree.F while maintaining the
same total pressure used in said oxidation-leaching operation prior
to said flashing operation.
5. The process as defined in claim 1 further comprising the step of
cooling the treated slurry to approximately 150.degree.F prior to
said separating step.
6. The process as defined in claim 5 wherein the concentration of
copper in said separated liquid phase following said cooling and
water removal step is about 75 grams per liter of copper.
7. The process as defined in claim 1 wherein water is removed from
said separated liquid phase during said cooling step.
8. The process as defined in claim 1 wherein a portion of said
cooled separated liquid phase is blended with said treated slurry
prior to separating said liquid phase from said solids.
9. The process as defined in claim 1 wherein the temperature of
said separated liquid phase is reduced to about 120.degree.F by
said cooling step.
10. The process as defined in claim 1 wherein said cooling of said
separated liquid phase is carried out using a barometric
condenser.
11. The process as defined in claim 1 wherein said spent
electrolyte has a concentration of between about 5 and 15 grams per
liter of copper and between about 100 and 170 grams per liter of
sulfuric acid.
12. The process as defined in claim 1 wherein said sulfiding agent
is hydrogen sulfide.
13. The process as defined in claim 1 wherein said sulfiding agent
is selected from the group consisting of hydrogen sulfide, ammonium
sulfide, sodium sulfide, ammonium hydrosulfide, sodium
hydrosulfide, potassium sulfide, and potassium hydrosulfide.
14. The process as defined in claim 1 wherein said treated spent
electrolyte is thickened to a solids content of between about 3 and
15 percent.
15. The process as defined in claim 1 wherein the copper content of
said liquid phase is reduced from between about 60 and 80 grams per
liter to about 5 grams per liter in said electrolytic
treatment.
16. The process as defined in claim 1 wherein the acidity of said
solutions in said autoclave system is reduced to between about 10
and 20 grams per liter of sulfuric acid by the addition of said
neutralizing agent.
17. The process as defined in claim 1 wherein the
oxidation-leaching operation is carried out at a temperature
between 425.degree.F and 450.degree.F.
18. The process as defined in claim 1 wherein the
oxidation-leaching operation is carried out using an oxygen partial
pressure of between about 100 and 200 pounds per square inch.
19. The process as defined in claim 1 wherein said neutralizing
agent is selected from the group consisting of the hydroxides,
oxides and carbonates of calcium, strontium and barium.
20. The process as defined in claim 1 wherein said neutralizing
agent is calcium carbonate, wherein the calcium in said calcium
carbonate forms insoluble calcium sulfate and wherein said calcium
sulfate is withdrawn from said autoclave as part of the solids in
said treated slurry.
21. The process as defined in claim 20 wherein said copper bearing
sulfide ore also contains arsenic, wherein a first part of said
calcium in said calcium carbonate forms insoluble calcium sulfate,
wherein said arsenic, solubilized to a metal sulfate in said
oxidation-leaching operation, is precipitated with at least a part
of said calcium as insoluble iron and calcium arsenates upon the
addition of said calcium carbonate, and wherein said iron and
calcium arsenates are withdrawn from said autoclave as part of the
solids in said treated slurry.
22. The process as defined in claim 1 wherein the electrolytic
copper produced is about 99.9 percent pure.
Description
BACKGROUND OF THE INVENTION
Increasingly, in recent years, consideration has been given to
employing hydrometallurgical processes for the recovery of metal
values from sulfide ores. One of the reasons for this is that
hydrometallurgical processes, for the most part, do not involve the
generation of sulfur dioxide and the consequent problems of air
pollution characteristic of pyrometallurgical process. Another
reason for considering hydrometallurgical processes is the extreme
flexibility afforded with regard to the virtually infinite sets of
unique conditions of temperature, pressure, retention time,
specific solvents and additives and procedures that can be resorted
to for whatever particular metallurgical separation one is
confronted with.
Where the final step in a hydrometallurgical copper recovery
process is to be based upon electrowinning, it would normally be
desirable that the preliminary steps for producing the electrolyte
include the best possible solids-liquid separation, reduction of
dissolved iron and arsenic to minimal levels, and concentration of
copper, as copper sulfate, to maximum levels in the feed to the
electrowinning step. While a variety of methods are available for
carrying out each of these steps, in most known processes these
individual methods do not always complement each other, i.e., they
lack in technological compatibility with regard to making up an
efficient, economical, pollution-free process. It is an object of
this invention to provide a process wherein each individual step is
uniquely suited to the others and to the overall process for
manufacturing a high quality metallic copper product efficiently
and economically. A secondary object is to circumvent the
production of sulfur dioxide and thereby negate the possiblity of
air pollution.
BRIEF SUMMARY OF THE INVENTION
In accordance with and furtherance of this stated objective the
present invention provides a process wherein a copper sulfide
flotation concentrate, at least a portion of which is chalcopyrite,
CuFeS.sub.2, is first mixed with water and recycle process liquors
and slurries. Alternatively, the concentrate may be slurried with
recycle process liquors and slurries only. The resulting slurry is
then subjected to a high temperature-high pressure operation in an
autoclave system, wherein the temperature is closely controlled and
an oxygen partial pressure (from industrial oxygen or an
oxygen-bearing gas) is maintained. The autoclave system may be a
single vessel or a series of vessels. The oxidation in the high
temperature-high pressure operation results in the formation of
sulfuric acid and the solubilization of the metal constituents of
the concentrate in the form of sulfates. In addition to copper,
iron and arsenic also are dissolved by the acid, but the latter two
are removed from solution, after solubilization of virtually all of
the copper, by the neutralization of the greater portion of the
sulfuric acid with a neutralizing agent such as limestone. The
neutralization of the sulfuric acid results in the precipitation of
virtually all of the dissolved iron and arsenic.
Exiting the high temperature-high pressure operation, the slurry is
flashed to atmospheric pressure and fed to a solids-liquid
separation system. The separated liquor from the solids-liquid
separation system is then cooled and fed to an electrolytic
deposition operation where end-product copper cathodes are
produced. The solids from the solids-liquid separation operation
are normally washed with water and the return wash water,
containing residual copper values, is recycled back to the initial
stages of the process. The washed solids, composed of the ore
gangue and the materials precipitated in the autoclave--Fe.sub.2
O.sub.3 (hematite), iron and calcium arsenates, and CaSO.sub.4
(anhydrite)--are sent on for further processing or to waste. The
spent electrolyte from the electrolytic deposition operation is
then treated with a sulfiding agent, such as hydrogen sulfide, to
precipitate, as copper sulfide, the copper values not deposited
with the copper cathodes. The copper sulfide slurry is then
thickened to a solids content of between about 3 and 15 percent and
recycled back to the initial stages of the process.
It is significant of the process of this invention that the various
steps required are uniquely compatible with each other and with the
overall process. For example, the high-temperature precipitation of
iron and calcium as hematite and anhydrite, respectively, makes the
subsequent solids-liquid separation much easier than if the
precipitated materials were ferric hydroxide and gypsum as would be
obtained in a low-temperature process. Additionally, the steps
wherein water vapor is removed from the liquor in the process of
venting noncondensable gases, or controlling the temperature, or
flashing, etc., are compatible with the overall object of the
leaching step, that is, providing a liquor with a maximum
concentration of copper as copper sulfite and a very low
concentration of dissolved iron.
These and other aspects of the invention will be understood more
thoroughly in the light of the following description, as
illustrated in the accopanying drawing.
BRIEF DESCRIPTION OF THE DRAWING
The drawing is a flow sheet illustrating the preferred embodiment
of the invention.
DETAILED DESCRIPTION
The preferred embodiment of the invention involves the treatment of
a chalcopyrite flotation concentrate and is best illustrated with
reference to the drawing. Referring to the drawing, than, a
chalcopyrite flotation concentrate 1, containing between about 20
and 30% copper, is fed into a slurry tank and mixed with recycle
wash water 16 containing residual amounts of copper sulfate and
sulfuric acid, and with a recycle acidic slurry of copper sulfide
22. The resultant slurry 2 is fed continuously to a high
temperature-high pressure oxidation-leaching operation which, in a
preferred embodiment, is carried out in a multi-compartment,
horizontal autoclave system equipped with agitators.
Withn the autoclave system the slurry temperature is maintained
between 350.degree. and 450.degree. F, and preferably between
425.degree. and 450.degree.F. The total pressure is maintained at
between about 300 and 1000 psig, and preferably between about 400
and 600 psig, by the introduction of a high quality
oxygen-containing gas 3 so as to provide an oxygen partial pressure
of between about 50 and 500 psi, and preferably between about 100
and 200 psi. The oxygen-containing gas used should have an oxygen
content of at least about 95% and preferably in the order of
99%.
The oxidizing chemical reactions in the autoclave system result in
the formation of sulfuric acid and the dissolution of the copper
and iron as sulfates. Because the oxidizing reactions are
exothermic, heat must be removed to maintain the temperature within
the desired range. Water may be removed from the system during the
oxidation-leaching operation, resulting in a desirable increase in
the concentration of the sulfates in the slurry liquor. The
autoclave system may be a single pressurized vessel or may include
several pressurized vessels.
A suitable process and apparatus for the removal of heat and water
during the oxidation-leaching operation is described in copending
application for U.S. Pat. Ser. No. 446,412, filed Feb. 27, 1974 by
Freddie J. Touro, one of the present inventors.
Removal of water also serves to increase the concentration of the
acidity of the slurry, which is desirable, since increased acidity
results in increased copper extraction rates when the
oxidation-leaching operation is carried out under the preferred
temperature conditions of 425.degree.-450.degree.F. Control of the
acidity during the oxidation-leaching of chalcopyrite forms the
subject matter of another copending application for U.S. Pat. Ser.
No. 446,315, filed Feb. 27, 1974 by Freddie J. Touro, one of the
present inventors.
To insure maximum removal of iron from the leach liquor, a
neutralizing agent such as limestone slurry 4 is added in the
autoclave system so as to reduce the acidity to less than about 40
grams per liter H.sub.2 SO.sub.4, preferably between about 10 and
20 grams per liter H.sub.2 SO.sub.4, and permit the hydrolysis of
Fe.sub.2 (SO.sub.4).sub.3 and precipitation of iron as Fe.sub.2
O.sub.3. Neutralization of the acidity at a temperature between
425.degree.-450.degree.F in the autoclave is a particularly
preferred embodiment of the process because in this temperature
range the iron is precipitated as hematite (Fe.sub.2 O.sub.3) and
the calcium as anhydrite (CaSO.sub.4), both of which are
crystalline in nature and their precipitation in this form results
in the satisfactory separation of the solids from the liquid later
in the process. In contrast, the precipitates obtained at low
temperatures (iron as ferric hydroxide, and calcium as gypsum) are
much more difficult to separate from the liquid phase.
While limestone, i.e., calcium carbonate, has been described as the
neutralizing agent used in the preferred embodiment, other
neutralizing agents may be utilized. For example, the hydroxides,
oxides and carbonates of calcium, strontium and barium, all of
which form insoluble sulfate, may be used as the neutralizing
agent.
Following the addition of the neutralizing agent, the slurry is
provided additional retention tim in the autoclave system at the
preferred temperature of between 425.degree. and 425.degree.F, and
at the stated oxygen partial pressure with constant mechanical
agitation. During this time, additional iron is precipitated and
further oxidation and dissolution of copper may occur. A
vapor-space bleed-stream may be vented to atmosphere from the
autoclave to purge the gradual buildup of contaminant nitrogen
(introduced with the oxygen-containing gas). Some oxygen is lost
with this bleedstream as well as some water vapor. The loss of the
latter again serves to increase the copper concentration in the
leach liquor and to remove some heat.
In the preferred embodiment the treated slurry stream 5 leaving the
autoclave is made up of a liquid phase composed of a solution of
acidic copper sulfate and solids composed of gangue, anhydrite
(CaSO.sub.4), and hematite (Fe.sub.2 O.sub.3). Before this liquid
phase is separated from these solids, the slurry is subjected to a
flashing operation. The flashing operation may optionally be
preceded by an indirect cooling operation. In the preferred
embodiment, the flashing operation is preceded by an indirect
cooling operation, as indicated in the drawing. Exiting the
autoclave, then, the slurry is first cooled with water 6 in an
indirect heat exchanger to approximately 275.degree.-400.degree.F
while maintaining the same pressure as used in the autoclave. The
heat exchanger may be a waste heat boiler, for example, generating
15 psig steam. This reduction in temperature is provided so that
when the cooled slurry 7 is finally flashed to atmospheric
pressure, the volume of flashed steam produced 8 is significantly
less than that which would be produced if flashing occurred at the
original temperature of between 425.degree. and 450.degree.F. As a
result, the velocity of the slurry-steam mixture through the
let-down valve is decreased and the erosion of the valve reduced in
severity.
In the process of being flashed to atmospheric pressure, additional
water vapor is lost and the copper concentration of the liquor
further increased. After flashing, the temperature of the slurry is
approximately 230.degree.F (somewhat elevated for atmospheric
pressure due to the content of dissolved salts). Further cooling of
the liquor, down to approximately 150.degree.F, may be carried out
before the separation of the liquid and solids by blending recycle
liquor 13, cooled to about 120.degree.F in a subsequent cooling
operation (described below), with the 230.degree.F slurry. The
resultant slurry 9, cooled to about 150.degree.F, is then fed to a
solids-liquid separation system with separated liquor 10 going to a
cooling operation and solids 14 going to a washing system. In a
preferred embodiment a barometric condenser is used to cool stream
10 to about 120.degree.F. A portion of the cooled separated liquor
13 is recycled to cool the slurry, as it leaves the flashing
operation, as described above. In the barometric condenser
additional water 11 is removed under vacuum at about 120.degree.F
to further increase the copper concentration of the leach liquor.
Removal of water at this point has the beneficial effect of
increasing the wash water-to-tailings ratio that may be used in the
washing system.
In the washing system the solids are washed with water 15 to
recover the retained copper sulfate solution. The return wash water
16, containing copper values and dilute sulfuric acid, is recycled
back to the slurry preparation facilities. The washed solids
tailings 17 may be sent to a tailings pond or to further treatment
for the recovery of gold or silver, etc. It is to be noted that any
type of solids-liquid separation techniques, such as filtration,
centrifuging, etc., may be employed in these operations; however, a
thickener or a combination of thickeners is preferred. In the
preferred embodiment, the portion of the 120.degree.F liquor from
the barometric condenser that is not recycled 12 is fed to the
electrolytic cells. At this point, the concentration of copper in
the separated copper sulfate solution has been increased to about
75 grams per liter.
In the electrolytic deposition process copper cathodes 18 of as
high as 99.9 percent purity can be produced as final product. Heat
is liberated during this operation and should be removed in order
to keep the temperature in the electrolytic cells at about
150.degree.F or less. Sulfuric acid is also produced to the
stoichiometric extent of the copper deposited so that the spent
electrolyte 19 normally contains between about 5 and 15 grams per
liter of copper and between about 100 and 170 grams per liter of
sulfuric acid.
The spent solution 19 is then treated with a sulfiding agent 20 to
precipitate the copper as copper sulfide. Any one of a number of
sulfiding agents which cause the precipitation of the copper from
spent electrolyte 19 as copper sulfide, such as hydrogen sulfide
(H.sub.2 S), ammonium sulfide (NH.sub.4).sub.2 S), sodium sulfide
(Na.sub.2 S), ammonium hydrosulfide (NH.sub.4 HS), sodium
hydrosulfide (NaHS), potassium sulfide (K.sub.2 S) and potassium
hydrosulfide (KHS), may be used. Preferably hydrogen sulfide
(H.sub.2 S) is used. Preferably an amount in excess of that
stoichiometrically required is employed in order to assure that
precipitation of virtually all of the copper present in the spent
electrolyte. The sulfided slurry 21 is processed in a thickener and
the thickened acidic copper sulfide (CuS) slurry 22, having a
solids content of between about 3 percent and 15 percent and
preferably higher than 5 percent, is recycled to the feed slurry
tank. About 70-90% of spent electrolyte stream 19 can be bled from
the process via stream 23 by this method with minimal loss of
copper. This bleed serves to eliminate from the process certain
soluble impurities such as calcium, magnesium, nickel and cobalt.
The acidic supernatant liquor stream 23 from the thickener may be
treated with a limestone slurry 24, as shown in the drawing, and
the resultant gypsum slurry 25 pumped to waste, i.e., to holding
ponds or land-fill areas, or it may be used in any other process
which requires the use of a dilute sulfuric acid solution.
Example
The following example illustrates the manner in which the process
of this invention may be operated.
Chalcopyrite concentrate 1 may be fed to the system of the drawing
and mixed with recycle streams 16 and 22 as previously indicated.
The composition of stream 1 is given below, in Table 1, together
with the typical compositions of other pertinent streams of the
process.
Oxidation-leaching of the slurried concentrate 2 is carried out at
425.degree.F and 100 psi of oxygen partial pressure in an autoclave
system. Total pressure is 445 psig. Limestone 4 is used to
partially neutralize the acidity during oxidation-leaching so that
the exiting slurry 5 going to the indirect cooler has an acidity of
about 20 grams per liter H.sub.2 SO.sub.4. The temperature of
slurry 5 is decreased in the indirect cooler to about 320.degree.F.
Partially cooled slurry 7 is then flashed to atmospheric pressure
and to a temperature of about 230.degree.F, and blended with cooled
recycle liquor stream 13 for further cooling to 150.degree.F before
entering the solids-liquid separation system.
The solids-liquid separation is carried out in one thickener, and
the bottoms 14 from this thickener are washed with water in a
countercurrent fashion using three thickeners. Tailings 17 from the
last washing thickener may be sent to waste. The overflow 10 from
the solids-liquid separation thickener is further cooled to
120.degree.F using a barometric condenser and then divided into
stream 13, which is used for blending with the flashed slurry, as
described above, and stream 12, which is fed to the electrolytic
deposition operation.
Copper cathodes 18 of high purity are produced in the electrolytic
cells. The temperature in this operation normally tends to rise and
so an external cooler (not shown) is provided to maintain it at
about 150.degree.F. In the electrolytic deposition operation the
copper content of the electrolyte is reduced from about 60-80 grams
per liter (in stream 12) to about 5 grams per liter (in stream 19).
Depletion of the copper, during electrolysis, to below about 5
grams per liter is not deemed economically attractive in view of
the convenience of scavenging the copper values not deposited
during electrolysis which is provided by the subsequent sulfiding
operation of our process.
Spent electrolyte 19 is treated with hydrogen sulfide 20 in an
amount sufficient to precipitate as copper sulfide virtually all of
the copper present in solution at this point. Sulfided slurry 21 is
thickened and recycled as stream 22 to the slurry feed tank. The
overflow stream 23 from the sulfide thickener contains most of the
impurities not rejected in the oxidation-leaching as precipitated
solids and is conveniently treated with limestone 24 and pumped to
waste 25. It is significant of the process of this invention that
these impurities are rejected from the system in the manner just
described and without losing any significant amount of copper
values. By contrast, a process in which the spent electrolyte 19 is
sent to waste would lose significant amounts of copper values as
dissolved copper in the waste spent electrolyte.
It is also significant of the process of this invention that the
copper values not deposited in the electrolytic deposition
operation are conveniently recovered by the recycling of a
relatively concentrated copper sulfide slurry 22 to the initial
stages of the process without the recycling of impurities or large
quantities of undesirable water to the initial stages of the
process. By contrast, a process which chose to recycle the spent
electrolyte 19 from the electrolyte deposition operation in order
to recover the copper values not deposited during electrolysis
would do so at the expense of recycling the impurities associated
with the spent electrolyte and, of course, of not being able to
remove water from the system at this point. The inability to remove
water from the system at this point would have the effect, in such
case, of decreasing the wash water-to-tailings ratio which may be
used in the washing operation of the process and this, in turn,
would increase the number of thickeners necessary to effectively
carry out the washing.
TABLE 1
__________________________________________________________________________
STREAM COMPOSITIONS OF A TYPICAL RUN OF PROCESS FOR THE MANUFACTURE
OF ELECTROLYTIC COPPER Stream* 1 12 17 19 22 23 Composition** % gpl
% gpl % gpl
__________________________________________________________________________
Copper (Cu) 26.1 75.1 0.70 5 0.001 Iron (Fe) 28.6 1.0 27.0 0.05 1.0
Total Sulfur 33.9 13.6 Free Acid (H.sub.2 SO.sub.4) 20.0 130 140.0
Calcium 0.18 0.57 13.5 0.06 0.57 Magnesium 0.33 0.013 0.32 Aluminum
0.55 0.01 0.55 Lead 0.01 0.12 0.005 Antimony 0.007 0.05 0.007
Bismuth 0.009 0.04 0.009 Nickel 0.06 0.008 0.06 Cobalt 0.05 0.002
0.05 Zinc 0.25 0.003 0.24 Manganese 0.015 0.015
__________________________________________________________________________
*All stream numbers refer to the drawing. **Compositions of streams
No. 1, 17 and 22 are given on a dry basis.
The terms and expressions which have been used here are used as
terms of description and not of limitation, and there is no
intention, in the use of such terms and expressions, of excluding
equivalents of the features shown and described, or portions
thereof, it being recognized that various modifications are
possible within the scope of the invention claimed.
* * * * *