Note: Descriptions are shown in the official language in which they were submitted.
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FIELD OF THE INVENTION
The present invention relates to the elec-trolytic re-
covery of cobalt, and more specifically to its recovery from a
cobaltic oxide hydrate which is contaminated with chloride ions.
BACKGROUND OF THE INVENTION
The recovery of cobalt from various process streams can
conveniently be carried out by first precipitating ~he cobalt
as a hydrated oxide and thereafter redissolving to produce an
electrolyte from which cobalt can be electrowon. Where, as
is common, the process stream contains significant amounts
of other metals, most notably nic~el, significant upgrading of
the relative amount of cobalt present is achievable by pre-
c-pitating the cobalt under oY~idative conditions which ensure
the formation of its trivalent o~ide hydrate, sometimes
referred to as cobaltic hydroxide, Co(OH)3. Such an oxidative
precipitation is achieved if the process stream is treated
with sodium hypochlorite or chlorine in the presence of a base.
While from an economic viewpoint the procedure of
forming a cobaltic oxide hydrate with the aid of chlorine is
attractive, an impediment to its commercial application stems
from the fact that the resulting filter cake is contaminated
with chloride ions. Two undesirable consequences flow from
such èntrained chloride which con-taminates the electrolyte
from which cobalt is to be electrowon. Firstly it necessitates
the use of relatively expensive anodes for the electrowinning
operation since the commonly used lead alloy anodes would
corrode rapidly in a chloride-containing electrolyte. More-
over, electrowinning from chloride-containing electrolytes is
accompanied by chlorine evolution which is environmentally
objectionable and necessitates the use o more elaborate cells
with means for containing and exhaustiny the atmosphere.
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OBJECT OF T~-IE INVENTION
An object of the presen-t invention is to provide a pro-
cedure by which a chloride-contaminated filter cake of cobaltic
oxide h~drate can be used to produce a solution from which
electrowinning can be carried out without the above-mentioned
impediments,
SUMM~RY OF THE INVENTION
The present invention provides a process in which a
feed which consists of a cobaltic ox:ide hydrate precipi-tate is
dissolved in a spent sulfate electrolyte from a cobalt electro-
winning operation and in which at least one of the feed and
spent electrolyte is contaminated with chloride ions, wherein
the improvement comprises mixing the feed with the spent
electrolyte to form a slurry/ sparging air through the slurry
for a period sufficient to liberate as gaseous chlorine sub-
stantially all of the chloride ions in the slurry, and there-
after treating the dechlorinated slurry with a reducing agent
selected from the group consisting of sulfur dioxide, hydrogen
pero~ide and organic reagents capable of reducing cobalt to
its divalent state; whereby substantially all of the cobalt
in the feed is dissolved to provide a substantially chloride-
free cobalt-containing solution from which pure cobalt can be
electrowon.
Particularly useful materials for preparing the cobalt
feed precipitate used in the process of the invention are
mixed cobalt-nickel basic precipitates which are formed as
intermediates in various nickel recovery schemes. Providing
the mi~ed precipitate contains nickel in an amount at least
equal to its cobalt content, the baslc nickel compound can
be relied on as the base needed to precipitate the desired
cobaltic compound as follows. The mixed precipitate is
separated into two fractions. A first of these fractions is
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dissolved in a dilute mineral acid-containiny aqueous solution.
This solution is then treated with chlorine while the second
fraction of mixed precipitate is added in a controlled manner
such as to ensure that a pH between about 2.5 and 4.5 is
maintained. In this way substantially all of the cobalt in
the mixed precipitate reports as a cobaltic oxide hydrate
which contains only a minor portion of the nickel present in
the mixed precipitate. This method of preparing the cobaltie
precipitate feed inevitably results in chloride contamination
thereof.
It has been found that the proeedure of slurrying the
feed precipitate with ~he aeidic sulfate solution which con~
sists of spent electrolyte and subjecting the slurry to an air
sparge is an effective means of liberating substantially all
of the entrained ehloride ions as gaseous chlorine. The
deehlorination step is necessary and is equally effective,
regardless of whether the contaminant ehloride was present in
the feed precipitate or in the spent eleetrolyte used. It is
essential that this dechlorination of the slurry takes plaee
prior to the leaehing operation, i.e., prior to introduction
of the redueing agent into the slurry. This is because the
eobaltic preeipitate plays a role in the dechlorination which
is believed to proeeed by virtue of the following reaction:
2Co(OH)3 + 2H2S04 + 2E + 2C1 ~ 2CoS04 ~ C12 + 6H20
The above reaction ean be carried out at room temperature/ but
for kinetie reasons it is preferred to perform it at a temp-
erature of the order of 60-65C. Under such conditions a
residual ehloride eoneentration of less than 20 mg/l ean be
attained in about 30 minutes by sparging air and also impart-
ing meehanieal agitation to the slurry.
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Subsequent to this pre-leachiny operation, a reducing
agent is used to induce dissolution of the cobaltic precipi-
tate. The reducing agent can be sulfur dioxide, in which case
the leach is believed to involve the following reaction:
2Co(OH)3 + H2S04 ~ S02~ ~ 2CoS04 ~ 4H20
Because dissolution in this manner adds to the sulfate ion
concentration, it becomes necessary to control i-ts build up.
This can be done by bleeding a portion of the cobalt sulfate
solution, preferably after it has been purified, and treating
it with sodium carbonate to precipitate cobalt carbonate,
part of which is used to treat impure electrolyte to remove
iron therefrom while the remainder is redissolved in the puri-
fied electrolyte to adjust the pH thereof.
Alternatively, the cobaltic precipitate can be dissolved
without sulfate build-up if use is made of a reductant other
than sulfur dioxide. Hydrogen peroxide can be used as a re-
ducing agent for this purposej though its cost makes it less
attractive than other reagents. It is known, however, that
many organic reagents are oxidized by cobaltic hydroxide,
i.e., such reagents are capable of reducing the cobalt to its
divalent state. Thus the use of alcohols, aldehydes and
ketones for this purpose is suggested by publications such as
"Oxidation of Some Organic Compounds By Cobaltic Hydroxide"
by S. Ludwik, in ~oczniki Chemii, 1973, 47, p. 43, and
"Reactions of the Cobaltic Ion, Part III : The Kinetics o
the Reaction of the Cobaltic Ion wlth Aldehydes and Alcohols"
by C. E. H. Bawn and A. G. White, in J. Chem. Soc., 1951, p.343.
We particularly prefer to use methanol as the reducing
agent for cobalt, in which case the reaction which takes
place is believed to be as follows:
6Co(OH)3 ~ 6H2SO~ ~ CH30H ~ 6CoS04 ~ C02 ~ 17H20
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This reaction proceeds rapidly in the initial stages of dis-
solution. However we have found that after cobalt dissolution
has proceeded to the extent of 85-90% completion, the leach
becomes slower and it is preferable for that reason to resort
to a different reductant to complete the leach. The final
treatment can be conveniently performed with sulfur dioxide
or with hydrogen peroxideO When this is done, an overall
dissolution of 98-99% of the cobalt can be achieved with a
total leaching time similar to that needed when sulfur
dioxide is used as the sole reductan-t. ~hen S02 is used for
all or part of the leach, it is desirable to employ a brief
air sparge subsequent to the leach, to elimlnate any excess
S2 from the solution.
While the dechlorination operation carried out in
accordance with the invention removes from the slurry chloride
ions which were present as such in either the feed cake or
the recycled spent electrolyte, a problem may arise if the
electrolyte contains chlorate ions. The latter may result
from chloride contamination of pregnant electrowinning electro-
lyte, due to impure reagents added, for example, to control
the pH. The anodic conditions during electrowinning can result
in formation of chlorate ions from any chloride in solution.
If this chlorate contamination is left unchecked, the air
sparging treatment would not succeed in removing the chlorate,
and the subsequent reduction leach of cobalt would be accom
panied by reduction of the chlorate so that the "dechlorinated"
slurry would be recontaminated with chloride ions. Accordingly,
where chlorate ions may be present in the rPcycled spent
electrolyte, we avoid the above-described problem by reducing
the chIorate ions to chloride ions prior to slurrying that
electrolyte with~the cobaltic feed. This can be accomplished
in many known ways~ such as by means of a short sparge with
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sulfur dioxide. Thus the sequence of operations in such a
case is:
i) treat the spent electrolyte with sulfur dioxide to
reduce C103 to Cl ;
ii) slurry the treated electrolyte with the cobaltic
feed precipitate;
iii) subject the slurry to an air sparge to li~erate as
gaseous chlorine the Cl present in both the precipitate and
solution; and
iv) introduce the reducing agent needed to effect cobalt
dissolution.
While at the end of the leaching operation some residue
will remain undissolved, ik is unnecessary to separate it at
this point of the procedure since it is economically desirable
to mlnimize the number of solid~liquid separations in any
commercial operation. Accordingly, the residue can be left
with the solution until the latter has been treated to remove
lead and iron therefrom and the combined residue and precipi-
tated impurities can then be separated from the solution.
The removal of lead and iron from the cobalt solution
can be carried out in any conventional manner. We prefer to
treat the solution with barium carbonate for lead removal
and thereafter with cobalt carbonate for iron removal. Where
copper is present as impurity, some of the copper may be pre-
cipitated in the iron removal stage.
- According to a preferred embodiment of the invention
the impurities zinc, copper and nickel are removed by means of
ion exchange resins. Such a manner of purification is
rendered economically acceptable by its use in conjunction
with an electrowinning operation which is carried out with a
high cobalt bite, i.e., a depletion of the cobalt concentration
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in the elctrolyte by at least about 35 grams/liter. ~y operatlng with such
a bite, use is made of concentrated electrolytes, so that purification of a
given amount of cobalt entails treating a relatively small volume of electro-
lyte which can be treated in a relatively small resin bed.
For zinc removal we prefer to use a resin which contains di(2 -
; ethylhexyl) phosphoric acid (which is hereinafter referred to as D2EHPA).
One such resin which is available commercially from Bayer AG is known by the
designation: Lewatit* OC1026 and is a macroporous copolymer of styrene-
divinylbenzene containing about 150 grams of D2EHPA per liter of resin bed.
The use of this resin for zinc removal from sulfate solutions containing 40
g/l of cobalt has been reported by others, and we have found it effective for
treating the more concentrated solutions preferred in the process of the
present invention wherein the cobalt level is of the order of 100 g/l or
more.
Re val of nickel is preferably carried out in the manner described
and claimed in copending application, Serial Number 333,728, filed August 1,
1979, and assigned in common with the present invention. The procedure makes
use of a resin having bis (2 - picolyl)amine functional groups, such a resin
being available from Dow Chemicals under the designation: XF4195. While the
reported selectivity of this resin between cobalt and nickel is less `~
attrac-tive than the selectivity reported for many other resins, the XF4195
resin was found to be much more effective than any of such other resins for
re ving nickel down to low levels from concentrated cobalt solutions.
Currently used procedures for electrowinning cobalt invariably
entail use of bag-free cells and operation with
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relatively low cobalt bite (i.e., less than 15 g/l) in order
to attain acceptahle current efficiency. While the use of
cathode boxes to achieve hlgher bites is well known in the art
of nickel electrowinning, cobalt producers have been prevented
from adopting such procedures by the tendency for cobalt
oxide slimes and possibly also gypsum to precipitate and hence
cause clogging of the diaphragms. We have found surprlsingly
that if the bite sought is not slightly but substantially
higher, i.e., between about 35 and 60 g/l, e.g., 45 g/l, then
provdding diaphragm cells are employed, the electrowinning
can proceed satisfactorily with high current efficiency and
without slime formation problems. We prefer to use cells in
which the diaphragm is provided in the form of a bag surround-
ing each anode, the anodes being o conventional lead based
material since chloride ions are absent from the electrolyte.
To aid in understanding the present invention, some
examples of the steps for recovering pure cobalt irom cobaltic
oxide hydrate precipita~es will now be described.
EXAMPLE 1
A wet Co(OH)3 cake was used, which contained 23~ Co and
0.1~ Cl . (Unless otherwise specified all percentages herein
quoted are percentages by weight). Preliminary tests showed
that if such a cake is merely dissolved in an aaidic sulfate
solution to produce an electrolyte containing of the order
of 50 g/l of Co, the resulting electrolyte contains at least
0.2 g/l of Cl , which is very much higher than can be toler-
ated. The dechlorination procedure in accordance with the
invention was carried out as ollows:
3.45 kg of the cake were slurried with 14.5 liters of
spent cobalt sulfate electroly-te which contained 42 g/l of Co,
0 23 g/l of ~i, 3.3 mg/l of Pb a~d 85 g/l of Hz504. The
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slurry was agitated mechanically and maintained at 65C while
air was sparged through it for 30 minutes. At the end of that
time assays showed that about 93~ of the chloride ions had
been eliminated, while only about 4~ of the cobalt in the
cake was dissol~ed. The filtrate at the ena of the dechlori-
nation assayed 20 mg/l Cl , 41 g/l Co and l g/l Ni.
To further test the efficiency of the dechlorination
procedure, the above procedure was repeated with a solution
which had been spiked with Cl so that the initial concentra~
tion in the slurry liquid was l g/l Cl . The resulting slurry
was sparged with air at 60C and sampled at various intervals
to determine the chloride content of the solution which was
found to be as shown in Table l.
TABLE 1
ISparging Time ~ Cl Assay
(min) (mg/l)
O 1,000.
:
. 60 20
150 13
It is clear that dechlorination proceeds rapidly at
this temperature and that an acceptable level of 20 mg/l of
Cl is attainable with a sparge duration of 30 minutes or less
even when the solution as well as the feed cake is highly con-
taminated with Cl . Of course, as explained above, if the
solution is contaminated with Cl03 rather than Cl , air
sparging alone will not remove the Cl03 and it is necessary
to reduce the Cl03 to Cl prior to formation of the slurry.
The dechlorinated slurry obtained in the manner de-
scribed above was then subjected to a reductive leach by intro-
ducing sulur dioxide into it at a sparging rate o 0.21 moles
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of S2 per liter per hour. Progress of the leach was followed by monitor-
ing the redox potential (relative to a saturated calomel electrode). The
initial redox potential of ~900 mV had dropped to +200 mV after 130 minutes
of leaching, corresponding to a sulfur dioxide consumption of 0.5 moles per
mole of cobalt in the feed cake. At this point an assay showed that the so-
lution contained:
Cobalt : 91 g/l
Nickel : 1.5 g/l
Lead : 12 mg/l
Iron : 0.2 g/l
Copper : 15 mg/l
Zinc : 5 mg/l
Lead was removed from the slurry by adding barium carbonate in an amount
corresponding to 0.5 g/l of slurry, which reduced the lead in solution after
30 minutes to less than 0.1 mg/l.
The slurry was thereafter neutralized to pH 5.5 by adding to it
2.3 liters of a CoCO3 slurry containing about 150 g/l of cobalt. The latter
constituted a partial recycle of cobalt in that it had been prepared by
treating a bleed stream of purified electrolyte with sodium carbonate.
After filtration of the neutralized slurry, 17.8 liters of filtrate were ob-
tained which analyzed:
Cobalt : 99.5 g/l
Nickel : 1.27 g/l
Lead : ~0.1 mg/l
Iron : 0.3 mg/l
Copper : 1.8 mg/l
Zinc : 3.8 mg/l
Cl : 30 mg/l
The leach residue separated from the above filtrate contained an
amount of cobalt representing 1.5% of the cobalt present in the feed cake.
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EXAMPLE 2
A similar test to that described in the previous example
was carried out to investigate the use of methanol as reduc-
tant during the leach. In this case 8.2 kg of wet cobaltic
oxide hydrate cake containing 25% Co and 0.1~ Cl were slurried
with 36 liters of spent cobalt sulfate electrolyte containing:
Cobalt : 5406 g/l
Nickel : 0.18 g/l
Lead : 2 mg/l
Sulfuric
Acid : 85 g/l
After a 30 minute air sparge at 65C it was found that
90~ of chloride had been liberated leaving an electrolyte which
contained 45 g/l Co, 0.8 g/1 Ni and 20 mg/1 Cl .
Pure methanol was added to the dechlorinated slurry at a
rate of 15 ml/hr/1 of slurry. After 20 minutes, corresponding
to a methanol addition of 0.17 moles of CH30H per mole of
cobalt in the feed cake, the pH of the slurry had increased
from 0.6 to 1.6. At this point a sample of ~iltrate Erom the
slurry assayed 89.5 g/l Co indicating extraction of about
88% of the cobalt present in the feed. The methanol introduc-
tion was discontinued and substituted by sulfur dioxide
sparging at a rate of 0.21 les/hr/l of slurry for lO0 minutes
at which time completion of the leach was evidenced by a drop
in the redox potential from ~700 mV to +440 mV at a pH of 2.3.
The leach was ~ollowed by a 20 minute air sparge during which
the redox potentlal rose to +690 mV. Analysis of the leach
solution gave the following results:
Cobalt : 98.2 g/l
` I~ickel : 1.12 g/1
; Lead : lO mg/l
Iron : 0.11 g/l
` Copper : 15 mg/1
~ Zinc : 4 mg/1
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~ emoval of lead from this solution required two consecutive
additions of BaCO3. The first addition of 0.5 g of BaC03 per liter of slurry
reduced the lead content to 0.4 mg/l after 15 minutes. An identical addition
reduced the lead to 0.2 mg/l in a further 15 minutes.
The resulting slurry was neutralized to pH 5.4 by adding to it 1
liter of a CoC03 slurry containing about 100 g/l of cobalt. After filtration
40 liters of electrolyte were obtained which assayed 0.3 mg/l ~e and 3 mg/l
Cu, while the separated residue represented 0.95% of the cobalt in the feed
cake.
E~AMPLE 3
The following tests illustrate the removal of zinc by ion exchange
from electrolytes having high concentrations of cobalt and high ionic strength.
A cobalt sulfate electrolyte containing 120 g/l Co, 1.2 g/l Ni,
0.020 g/l Zn and about 50 g/l Na2SO4 and having a pH of 5.5 measured at 22 C
was treated with 50 ml of "Lewatit* OC1026" resin. The resin, which as
stated earlier comprises a copolymer containing 150 grams of D2EHPA per liter
of resin bed, was contained in a columnar bed 1.7 cm in diameter and 20 cm
deep. The column was operated with a flow of 2 cubic meters of solution
per hour per square meter of bed cross section (m3/m /hr) and maintained at
20 50C. After processing 0.5 liters of solution, i.e., 10 bed volumes (B.V.)
the column effluent was found to analyze less than 0.2 mg/l Zn. This
represents a ratio of Co to Zn in the purified electrolyte greater than
6 x 10 .
On a larger scale a solution containing 100 g/l Co, 1 g/l Ni,
0.005 g/l Zn and 100 g/l Na2SO4 and having a pH of 5.0 measured at 22 C
was purified in a column 4.1 cm in diameter, 79 cm deep and containing 1
liter of the "Lewatit* OC1026" resin. The column was operated at 60 C and at
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a rate of 4.5 m3/m2/hr to process 213 ]iters (i.e., 213 B.V.) of the
solution, at the end of which time the effluent assayed only 0.7 mg/l Zn.
This represents a Co/Zn ratio higher than 1 x 105 in the purified
electrolyte.
It is clear that a D2EHPA containing resin provides an effective
method of removing zinc from the concentrated electrolytes of the invention.
The extraction should be performed at a pH which is not less than about 2.5
and preferably the pH should be initially adjusted, if necessary, to a value
in the range 4 to 6. The resin bed, once loaded, can be eluted with a dilute
mineral acid, and thereafter re-used for further zinc removal.
After passage through the resin bed, the electrolyte may contain
some D2EHPA due to the slight solubility of the latter in aqueous solutions.
To remove the small quantities of extractant from the electrolyte, the latter
is preferably passed through a column of activated carbon.
EXAMPLE 4
The following tests illustrate the removal of nickel by ion
exchange from electrolytes having high concentrations of cobalt and high
ionic strength.
A solution containing 125 g/l Co, 2.04 g/l Ni and about 50 g/l
Na2S04 and having a pH of 6.1 measured at 22 C was treated with Dow Chemicals
Company's bis (2 - picolyl) amine resin in a fixed bed of 5 cm diameter and
3 meter depth containing 6 liters of the resin. The solution to be purified
was passed upwards through the resin column at 50 C and at a rate of 3m3/m /
hr. Samples of the effluent taken after processing of various bed volumes of
solution were analyzed for nickel. The results showed that after two bed
volumes of solutions had been processed, the sample of effluent exiting from
the bed contained
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only 33 mg/l of nickel, i.e., it represented an electrolyte
with a Co/Ni ratio higher than 3000. A sample analysis of
effluent exitin~ after about 4 B.V. had been processed showed
a nickel content of 490 mg/1. By determining the average
assay of the total ef~luent collected, it was determined that
with such a starting solution as much as 6 B.V. could be
treated in the column while ob-taining a purified electroly-te
in which the nickel did not exceed about 0.5 g/l, i.e., the
Co/Ni ratio was at least 200. Such an electroly~e enables
cobalt of very high purity to be electrowon.
The above described ion exchange treatment is effective
to remove not only nickel but also any copper present in the
electrolyte. Stripping of the copper from the picolylamine
resin is more difficult to accomplish than stripping of nickel
from the loaded resin~ Fox this reason we preer to remove
copper prior to using this resin. The removal of copper can
be accomplished in various known ways, such as by using
specific copper-selecti~e ion exchange resins or solvent extrac-
tants such as carboxylic acid or oxime types of extractant.
EXAMPLE 5
A first purified electrolyte which contained:
Cobalt : 89 g/l
Nickel o 20 mg/l
Lead : 0.4 mg/l
Zinc : 1.3 mg/l
Iron : 0.3 mg/l
Copper : 0.1 mg/l
.
was used to electrowin cobalt in 2.5 liter laboratory cells as
well as in larger 16 liter cells. The electrodes for these
cells were lO x 15 cm and lO x 35 cm respectively, the anodes
being made of a lead alloy containing 0.05~ calcium and 0.5~
tin. The cathodes were made of stainless steel. Each anode
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was surrounded by a diaphram defining an anolyte compartment
while the bulk of the cell space constituted a comnon catholyte
compartment.
A small amount, 20-30 mg/l, of sodium lauryl sulfate
was added to the electrolyte to act as antipitting and anti-
misting agent. The electrolyte was then fed through the cells
at a rate such that the cobalt bite was 41 g/l, the current
density used being about 200 A/m . The cell temperature was
55C and the pH of the catholyte was 2.1. Cobalt was plated
for 171 hours with a current efficiency of 92%. The plates
deposited were found to contain the following impurities, in
parts per million:
Nickel : 87 ppm
Lead o 5 ppm
2inc : 11 ppm
Iron ~ 16 ppm
Copper : 2.5 ppm
A second electrowinning test was done using similar
procedure except for the following differences~ The electrolyte
had a higher level of impurities, its assay being:
Cobalt o 93 g/l
Nickel ~ 0.24 g/l
Lead : 0.2 mgjl
Zinc : 1 mg/l
Iron 0.2 mg/l
. Copper : 0.3 mg/l
The cathodes were masked in this case to expose only
circular islands on which discrete cobalt deposits were formed.
The bite achieved was 46 g/l of cobalt with a catholyte pH
of 2.3 and a current efficiency of 91%. After 167 hours of
electrowinning the deposit was found to contain:
Nickel : 252 ppm
Lead : ~4 ppm
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Zinc : 10 ppm
Iron : <8 ppm
Copper : <7 ppm
By -testing the effect of variations in the eleckrowin-
ning parameters, the following were established as desired and
preferred conditions:
Feed composition: 85-105 g/l Co preferably about
100 g/l Co;
Cobalt Bite: 35-60 g/l, preferably about 45 g/l;
Spent ElectrOlyte > 40 g/l Co, preferably > 50 g/l Co;
Composltion:
Current Density: 100-300 A/m2, preferably about Z00 A/m2;
Temperature: 50-60C, preferably about 55C;
Ca-tholyte pH: 2-3, preferably about 2.5.
The cobalt bite has to be at least 35 g/1 in order to
avoid slime formation problems. This need for a minimum bite
is illustrated by the results of the following comparative
test. Identical electrowinning experiments were carried out
under conditions similar to those of the second of the afore-
mentioned electrowinning tests except that the bite was
arranged to be 45 g/l in one case and only 12 g/l in the other
case. The feed electrolytes were chosen to ensure that in
both cases the same cobalt level (50 g/l) was present in the
spent electrolyteO After 167 hours of electrodeposition the
total amount of slimes collected from the anode box and spent
electrolyte was found to be 13.0 grams when the low bite was
used, but only 2.0 grams when the high bite was used.
The blte cannot, however, be chosen to be above about
60 g/l without detriment to the operation. This is because
at excessivel~ high bites the catholyte pH drops to ~2 and
this causes both a lowering of current e~iciency and pitting
of the deposits. Thus a comparative test wherein a 75 g/l
cobalt bite was achieved showed a current efficiency of only
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55~. Moreover, a large nurnber of pits were discovered in the
deposits after 100 hours despite the presence of the sodium
lauryl sulfate.
The present invention has been described with reference
to preferred embodiments thereof. It will be appreciated that
various modifications may be made to the details of such
embodiments without detracting from the benefits of the
present invention. Thus other reagents may be used for puri-
fying the electrolyte, and other procedures adopted for the
purification and eventual electrowinning. Furthermore, the
steps which have been described as batch operations may be
performed in a continuous or semi-continuous manner. These
and other modifications are within the scope of the invention
whi~h is d~fin~d by 'he appe~ded claims.
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