United States Patent [19]
Hazen
[11] 3,767,543
[45] Oct. 23, 1973
[54] PROCESS FOR THE ELECTROLYTIC
RECOVERY OF COPPER FROM ITS
SULFIDE ORES
805,969
415,738
333,815
11/1905
11/1889
1/1886
Hybinette 204/112
Seegall 204/107
Body 204/107
[56]
3,464,904
1,726,258
1,539,713
1,485,909
1,456,784
1,441,063
1,435,891
1,434,088
1,128,315 7 Claims, 1 Drawing Figure
Primary Examiner-John H. Mack
Assistant Examiner-R. L. Andrews
Attorney....:.·Sheridan, Ross & Fields
[57] ABSTRACT
An improvement in the ferric chloride leach recovery
of copper from its sulfide ores which comprises recovery
of the copper from the leach solution after removal
of sulfur by electrolysis rather than by conventional
cementation with added iron.
FOREIGN PATENTS OR APPLICATIONS
66,547 1/1893 Germany
OTHER PUBLICATIONS
Principles of Electroplating & Electroforming by Blum
et aI., 3rd ed.; 1949, pgs. 68-69.
The Electromotive Series, "Simple Methods for Analyzing
Plating Solutions," 7th ed, 1949, Hanson-Van
Winkle-Munning Co., p. 20.
References Cited
UNITED STATES PATENTS
9/1969 Brace 204/105
8/1929 Christensen 204/117
5/1925 Christensen 204/111
3/1924 Christensen 75/104
5/1923 Christensen 204/117
1/1923 Christensen 75/1 04
11/1922 Christensen 75/104
10/1922 Christensen 75/104
2/1915 Hybinette 204/106
Inventor: Wayne C. Hazen, Denver, Colo.
Assignee: Hazen Research, Inc., Golden, Colo.
Filed: June 28, 1971
Appl. No.: 157,281
U.S. CI. 204/107,204/52 R
Int. CI. C22d 1/16, C23b 5/20
Field of Search 204/1 07, 52 R;
75/117
[52]
[51 ]
[58]
[75]
[73]
[22]
[21 ]
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AW
IRON
OPPER
::0
ICHALCOPYRITE I
-i LEACH ING1
PURIFICATION FOR
IFILTRATION1 SOLUTION REMOVAL OF
~
BISMUTH, LEAD,
SOLIDS
SILVER
ELEMENTAL
SULPHUR
SULPHUR RECOVERY
~
BY SOLVENT EXTRACTION
ELECTROLYSIS IN
FLOTATION, MELTING ' DIAPHRAGM CELL C
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,
RESIDUE TO BE
TREATED FOR
ELECTROLYSIS IN
RECOVERY OF GOLD DIAPHRAGM. CELL
FERRIC CHLORIDE SOLUTION r
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3,767,543
2
copper in solution as copper chloride. To the extent
that iron is present in the sulfide mineral, as for example
in chalcopyrite, this iron is dissolved as ferrous
chloride and the associated sulfur is oxidized to ele-
In the present methods of treating copper sulfide ores 5 mental sulfur as is the case with the copper sulfide.
for the recovery of copper the general practice involves The ferric chloride leaching has very little effect
smelting the sulfides and through a complex series of upon the mineral pyrite when present by itself but, in
smelting operations driving off the sulfur as sulfur diox- general, will attack most sulfides such as pyrrhotite,
ide to produce a crude copper metal which is subse- chalcopyrite, chalcocite, sphalerite, and the like. Once
quently refined by electrolysis. This process is an an- 10 the leaching reaction has taken place and the metallic
cient one which has been responsible for the produc- elements are in solution as the chloride the sulfur retion
of most of the copper in the history of the world mains in the residue as elemental sulfur. In this reaction
but suffers from some serious drawbacks. One very se- the ferric chloride is reduced to ferrous chloride. The
rious drawback of a smelting operation is that the sulfur slurry is removed from the leaching vessel and sulfur is
dioxide produced from some steps of the operation is 15 separated by any standard method such as filtration or
of such low grade that it is very uneconomical to pro- countercurrent decantation thereby separating it from
duce sulfuric acid from it by means of the contact pro- the metal. It can be seen that this is obviously a great
ce.ss and, accordingly, most smelting operations dis- advantage because the sulfur has been converted to a
charge this low grade sulfur dioxide waste gas into the form that can be recovered and sold while the metallic
atmosphere. This creates a serious pollution problem 20 elements remain in solution as chlorides for subsequent
and for this reason the smelting process is the subject treatment.
of intense investigative efforts at the present time to In the process described in the Canadian Department
find ways of avoiding this pollution of the atmosphere. of Mines publications, the leach solution containing
In addition to the above disadvantage a smelting op- copper chloride and ferrous chloride is treated by ceeration
is economical only if done on a large scale and 25 mentation in which metallic iron is added to the soluthis
accounts for the growth of central smelting facili- tion and the copper thereby precipitated as cement
ties with the result that the capital investment for a copper. This procedure, of course, increases the quansmelting
complex is large. Small mine operators who tity of iron in solution which must be removed eventuproduce
copper sulfide concentrate cannot afford the ally and produces a copper of relatively low grade that
heavy capital investment of the pyrometallurgical 30 must be purified and retreated. The solution remaining
smelters and must therefore sell the output of their after electrolytic removal of copper and contained lead
mines to companies already possessing such facilities. is then subjected to a second electrolysis where the iron
This oftentimes leads to high shipping charges and high is removed by deposition on a cathode in an electrosmelter
toll fees. lytic cell and some ofthe ferrous chloride in the anolyte
Over the years many efforts have been made to find 35 is oxidized to ferric chloride, thus regenerating the
means of treating copper sulfide minerals by other leaching reagent for reuse.
methods than the ordinary smelting practice. Among In this flowsheet the amount of electrolytic iron
these are processes in which the copper sulfide is first which must be recovered is the total amount dissolved
roasted to a mixture of copper oxide and copper sulfate from the original concentrate feed and the amount
with the production of by-product sulfur dioxide gas of 40 which was added during the cementation for removal
a concentration high enough to permit its economic of the copper. The disadvantage of the recovery of the
conversion into sulfuric aci& copper in this way is that the copper which is produced
Certain features of this process have disadvantages. is of relatively low grade and is made expensive be-
The calcine from the furnace which contains the cop- causeofthe cost of theiron which is added. In addition,
per as a mixture of oxide and sulfate is leached with an 45 this added quantity of iron must then be removed from
acidic solution thereby producing a strong copper sul- the circuit by electrolysis in the next stage thereby refate
solution. This copper sulfate solution is then sepa- quiring a considerable increase in the size of the elecrated
from the insoluble materials and subjected to trolytic circuit that is required for the total iron. reelectrolysis
to produce copper metal and sulfuric acid. moval with consequent increase in the cost for recov-
The latter is in such weak solution that it cannot be 50 ery of the copper.
economically utilized, so that the loss of this by- In the flowsheet of this invention the teaching of the
product is a disadvantage. Furthermore, the disposal of Canadian Department of Mines' process can be folthe
weak sulfuric acid solution presents an economic lowed as far as the leaching and separation of liquids
problem. . . 55 and solids but the improvement lies in the method of
Another effort to solve the problem of recovery of recovering the copper. It has been found, surprisingly,
copper values from copper sulfide minerals is the ferric that it is possible to remove copper from the ferric
chloride leaching method. This method has been stud- chloride leach solution by electrolysis. Ordinary elecied
at great length in various laboratories and particu- trolysis of copper universally uses sulfuric acid solution
larly in the Canadian Department of Mines many years 60 as the electrolyte from which the copper is deposited
ago. (See "Investigations in Ore Dressing and Metal- on a cathode. It has been found that by conducting the
lurgy 1924,"Canada, Department of Mines, Mines electrolysis in a diaphragm-type cell, it is possible to re-
Branch, John McLeish, Director, No. 643.) In this pro- move the copper at quite high efficiency and high curcess,
the copper sulfide minerals are agitated with a hot rent density and at the same time generate ferric chlosolution
containing a high concentration of ferric chlo- 65 ride in the anolyte compartment. This not only removes
ride. The ferric chloride acts as an oxidizing agent to the requirement for the addition of iron in order to preattack
the copper bearing sulfide minerals thereby con- cipitate the copper but it regenerates the leaching reverting
the sulfide to elemental sulfur and putting the agent, ferric chloride, whereas the previous process pe-
1
PROCESS FOR THE ELECTROLYTllC RECOVERY
OF COPPER FROM ITS SULFIDE ORES
SUMMARY OF THE INVENTION
3,767,543
4
Perlodle Cumulative
60 0 0
66 5.5 5.5
67 12. !I 18. 4
65 25.7 44.1
68 25.3 69.4
71 21. 9 91. 3
69 5.9 97.2
72 .. __ .. .. .... _
71 .. .. _.. _
64.8
61.1
52.5
35.4
18.5
3.9
0.01
0.002
0.004
Catholyte
a./I., Temp.,
Elapsed time, hours Cu 0 C.
iron to ferric at the anode. If insufficient ferrous iron
is available for oxidation, chlorine is also produced at
the anode.
In the first test the cell was filled with electrolyte medium
which contained 65 g/l copper and 135 g/l iron.
Approximately 100 cc/hr of 135 g/1 ferrous chloride
solution was added to the cathode compartment and 12
amps were passed through the cell. After 8 hours of operation
when practically all of the copper was plated
out the test was terminated. The cathode had increased
in weight by 116 grams. The voltage used was, of
course, below the voltage requirement for the deposition
of iron. The bulk of the cathode deposit was hard
and coherent, but about 10 grams was spongy and contained
only 70% copper. Data collected during the run
are shown in the following table.
o..._.. ...... __ _..
1. .. _.. __ _
2 . __ ... _
3 __ _
4. . ....
5 .. .. _
6 .. .. __
7 __ .......... .. _
8 .. ...... _
30
The above results show that the process produces over
35 99 percent recovery of the copper in solution.
The iron in solution, of course, is electrolytically recovered
in a subsequent stage by conventional electrolysis
with simultaneous formation of ferric chloride
which is recycled to the initial leach stage in a commer-
40 cial process for treatment of copper sulfide concentrate.
Tests were run at various current densities to determine
the most favorable current density ranges for
forming spongy or coherent copper deposits, and to
45 test the efficiency of the process as applied to low copper
content electrolytes which correspond to electrolytes
resulting from leaching low copper content concentrates.
In the second test the copper content of the used
50 catholyte from the first test was increased to 2 g/l copper
by the addition of cupric chloride. The solution
added to the cathode compartment during electrolysis
contained 180 gil ferrous iron and 36 g/l copper. The
solution was added at approximately 160 cc/hr while
55 12 amps were passed through the cell. Approximately
36 grams of copper were added to the cell. At no time
was the catholyte ab,ove 0.66 g/l copper. Approximately
all of the copper in solution was deposited at the
60 ~;:~~~' half of which was coherent and the remainder
In the third test all conditions were the same as the
second test except that 24 amps were used and the solution
flow was increased to 335 cc/hr. In this test sub65
stantially all of the copper in solution was deposited as
a spongy deposit.
The fourth test was also similar to those of the second
test except that only 6 amps were used and the solution
nalized the circuit by the addition of iron above that in
the concentrate which had to be removed subsequently.
The net result of the present procedure is a significant
reduction in the capital expense of a plant and
the operating cost thereof for producing copper from 5
sulfide minerals.
In ordinary electrowinning practice for copper a considerable
effort is made to restrict the quantity of iron
which is in solution with the copper. If the iron content
of the solution becomes high (for example, 10g/1), a 10
lowering of the current efficiency is observed and a
change in the physical character of the deposited copper
may take place. Also, if the iron content is very
high, the current efficiency for the electrodeposition of
the copper may drop to such a low value as to render 15
the operation uneconomic. Therefore, the fact that the
copper can be deposited from a solution which contains
above 100 g/l of iron as ferric chloride is quite
surprising. The reason that the iron does not appreciably
interfere with the current efficiency is probably be- 20 _
cause the electrolysis is done in a diaphragm cell so that
the ferric iron which is formed at the anode cannot mix
and come in contact with the deposited copper to redissolve
it. Therefore, by performing the electrolysis in
this manner, it is possible to take advantage of the elec- 25
tric current being used to deposit the copper and at the
same time regenerate the leaching reagent without paying
the penalty of low current efficiency ordinarily encountered
when electrolysing a copper solution high in
iron.
The iron in the solution after the electrolytic removal
of copper is recovered by conventional electrolysis.
This amount of iron is, of course, restricted to that
which existed in the concentrate. Other metals are recovered
by conventional means.
The process of the invention as illustrated in the accompanying
drawing, a flow sheet of the process, will
now be described, along with the equipment used.
The tests for which results are given below were performed
using a ferrous chloride electrolyte medium
having various adjusted concentrations of iron and to
which cupric chloride was added in amounts to provide
various concentrations of copper. The electrolyte medium
simulated the leach solution which exists after
copper sulfide concentrate is leached with ferric chloride.
The equation representing the chemical reaction
in the leaching step is:
4 Fe Cb + Cu Fe 82 ...... Cu CI2 + 5 Fe Cl2 + 28.
After the sulfur is removed the electrolyte media contains
essentially ferrous chloride and cupric chloride as
does the electrolyte medium used in the tests.
A diaphragm cell was used in all tests. It consisted of
a 1,500 cc cathode compartment separated from a 300
cc anode compartment by a Dynel filter cloth diaphragm.
The anode compartment was equipped with a
solution overflow. The anode was high purity graphite
and had approximately a 0.1 sq.ft. area immersed. The
cathode was a sheet of 22-gauge copper with approximately
0.33 sq.ft. submerged. During each of the tests,
ferrous chloride electrolyte medium was continuously
added to the cathode compartment. This solution
flowed through the diaphragm into the anode compartment
and out of the cell. The pH of the electrolyte was
maintained at about 2.3 or below.
The electrolytic reactions are reduction of the copper
to metal at the cathode and oxidation of ferrous
3,767,543
5
flow was reduced to only 85 cc/hr. Practically all of the
copper deposited was in a coherent form.
The data from the tests where the catholyte copper
composition was kept low could be summarized as fol_
lOWS:
These results indicate that copper can be quantitatively
recovered from concentrated ferrous chloride solutions
( 180 gil iron and above) by electrowinning. At current
densities below about 20 asf the deposit is coherent and
above 20 asf the deposit contains a gradually increasing
amount of spongy material with practically all of it
being spongy above about 70 asf. Copper solutions can
be plated down to less than 0.1 gil copper. The percentage
recovery of copper is not affected appreciatively
by the concentration of the copper solution.
Wltile the invention has been ilIustrated by its application
to a flow sheet applied to calcopyrite, it is obviously
not limited to this mineral as it is equally applicable
to all concentrates of copper sulfide ores which can
be leached with ferric chloride.
I claim:
1. The process for recovering copper from chalcopy-
Amps
24 _
12 _
6 _
Effective
current
densityha.s.f. Deposit
(cat ode) description
80 All spongY.
40 Half spongy.
20 Coherent.
6
rite concentrate which comprises:
a. leaching the concentrate with ferric chloride,
b. separating sulfur from the leach solution of (a),
c. subjecting the leach solution of (b) to electrolysis
5 in a diaphragm cell having an anode and cathode
to electroplate copper at the cathode while continuously
flowing ferrous chloride through the diaphragm
from the cathode to the anode to prevent
ferric ions from contacting the cathode.
10 2. The process of claim 1 in which the ferric chloride
is recovered at the anode and is recycled to step (a).
3. The process of claim 1 in which iron in solution is
recovered in a second electrolysis at the cathode and
ferric chloride is recovered at the anode and recycled
15 to step (a).
4. The process of claim 1 in which the pH of the leach
solution subjected to electrolysis is maintained at about
2.3 or below.
20 S. The process of claim 1 in which the iron content
of the leach solution subjected to electrolysis is up to
about 180 grams per liter.
6. The process of claim 1 in which a cathode current
density below about 20 asf is used for the electrolysis
25 of copper to produce a coherent deposit on the cathode.
7. The process of claim 1 in which a cathode current
density inexcess of about 70 asf is used in the electrolysis
of copper to provide a spongy deposit on the cath30
ode.
* * * * *
35
40
45
50
55
60
65
UNITED STATES PATENT OFFICE
CERTIFICATE OF CORRECTION
Patent No. 3,767,543 Dated October 23, 1973
Inventor(s)--W-a-yne-C-.-H-az-en------------------
It is certified that error appears in the above-identified patent
and that said Letters Patent are hereby corrected as shown below:
Column 2,
--cementation--;
line 31 ~electrolyticn should be
line 32, "asecond" should be cancelled.
Signed and sealed this 15th day of October 1974.
(SEAL)
Attest:
McCOY M. GIBSON JR.
Attesting Officer
C. MARSl·L\.LL Di\NN
Commissioner of Patents
FORM PO-lOSO 00-69)
USCOMM-DC l10376'Pll~
'Jt u, S, GOVERNMENT "RINTIMG OFFIC[ ~ 19ft O-J'6-33",
Roman\�Tei��0�reast-font-family: HiddenHorzOCR'>values therefrom;
(e) extracting rhenium 'Values from the mother liquor
of (d) with a liquid water insoluble amine ion exchange
agent;
3,681,016
25 HERBERT T. CARl1ER, Primary Examiner
u.s. Cl. X.R.
8
ing the pH of rthe solution' to a pH between about
2.5-3.5;
(d) separating the crystallized ammonium tetramolybdate
from the mother liquor of (c);
(e) extracting rhenium values from the mother liquor
of (d) with a liquid water insoluble amine ion exchange
agent;
(f) stripping rhenium values from the loaded agent of
(e) with sodium hydroxide;
(g) recovering rhenium values from the strip solution
of (f) by extracting with pyridine; and
(h) recovering rhenium from the pyridine extractant
by distilling off the pyridine.
References Cited
UNITED STATES PATENTS
7/10969 Litz , 23-15 W
7/1969 !Platzke et al. 23-15W
3/1959 Zimmerley et al. 23-18 X
7/1960 Zimmerley et al. 23-24
4/1966 Churchward , 23~15 W
211970 Ziegenbaly et al. 23-23 X
1/1971 Proter et al. 23-22
3,455,'677
3,45:8,277
2,876,065
2,945,743
3,244,475
3,495,934
3,558,268
23-23, 24 R, 51 R
7
(f) stripping the loaded agent of (e) with an alkali
metal hydroxide;
(g) extracting rhenium values from the strip solution
of (f) with pyridine or pyridine derivative; and
(h) recovering rhenium from the pyridine extractant 5
by distilling off the pyridine.
2. The process of claim 1 in which metal ion impurities
are removed from the strip solution of .(b) before
crystallizing ammonium tetramolybdate in (c).
3. The process of claim 1 in which the anion exchange 10
agent in (a) is a tertiary amine ion exchange resin and the
stripping solution of (b) is ammonium hydroxide.
4. A process for recovering molybdenum and rhenium
values from pregnant acid leach solutions containing these
values together with other metal impurities and derived 15
from dusts and flue gases resulting from roasting relatively
impure molybdenite concentrate, said process comprising:
(a) extracting molybdenum and rhenium values from
the pregnant acid solution with a liquid water in- 20
soluble amine ion exchange agent;
(b) stripping the molybdenum and rhenium values
from the exchange resin with ammonium hydroxide
solution to form a strip solution containing the molybdenum
as ammonium molybdate and the rhenium
as ammonium perrhenate;
(c) crystallizing the molybendum from the strip solution
in (b) as ammonium tetramolybdate by adjust