U.S. patent number RE37,251 [Application Number 09/358,555] was granted by the patent office on 2001-07-03 for chloride assisted hydrometallurgical extraction of metal.
This patent grant is currently assigned to Cominco Engineering Services Ltd.. Invention is credited to David L. Jones.
United States Patent |
RE37,251 |
Jones |
July 3, 2001 |
Chloride assisted hydrometallurgical extraction of metal
Abstract
A process for the extraction of a metal from an ore or
concentrate comprises subjecting the ore or concentrate to pressure
oxidation in the presence of oxygen and an acidic solution
containing halogen ions and a source of bisulphate or sulphate
ions, such as H.sub.2 SO.sub.4. The metals which can be extracted
by the process comprises copper as well as non-cuprous metals such
as zinc, nickel and cobalt. During pressure oxidation the metal may
be precipitated as an insoluble basic salt, such as basic copper
sulphate, or substantially completely solubilized and precipitated
later as the basic metal salt.
Inventors: |
Jones; David L. (Delta,
CA) |
Assignee: |
Cominco Engineering Services
Ltd. (Vancouver, CA)
|
Family
ID: |
27378677 |
Appl.
No.: |
09/358,555 |
Filed: |
July 22, 1999 |
Related U.S. Patent Documents
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Application
Number |
Filing Date |
Patent Number |
Issue Date |
|
|
425117 |
Apr 21, 1995 |
5645708 |
|
|
|
098874 |
Jul 29, 1993 |
5431788 |
|
|
Reissue of: |
488128 |
Jun 7, 1995 |
05650057 |
Jul 22, 1997 |
|
|
Current U.S.
Class: |
205/579; 205/574;
205/602; 205/604; 205/606; 205/607; 423/27; 423/36 |
Current CPC
Class: |
C22B
3/06 (20130101); C22B 15/0067 (20130101); C22B
23/043 (20130101); C22B 3/10 (20130101); C22B
15/0071 (20130101); C22B 3/08 (20130101); C22B
15/0089 (20130101); C22B 15/0084 (20130101); C22B
19/26 (20130101); C22B 3/20 (20130101); C22B
23/0423 (20130101); C22B 3/26 (20210501); C22B
23/0415 (20130101); C22B 19/22 (20130101); C22B
3/24 (20130101); C22B 23/0453 (20130101); C22B
3/3842 (20210501); C22B 15/0069 (20130101); C22B
11/00 (20130101); Y02P 10/20 (20151101) |
Current International
Class: |
C22B
3/26 (20060101); C22B 11/00 (20060101); C22B
3/06 (20060101); C22B 3/08 (20060101); C22B
3/20 (20060101); C22B 3/38 (20060101); C22B
3/00 (20060101); C22B 15/00 (20060101); C22B
3/10 (20060101); C25C 001/00 () |
Field of
Search: |
;205/574,579,602,604,606,607 ;423/27,36 |
References Cited
[Referenced By]
U.S. Patent Documents
Foreign Patent Documents
Other References
Subramanian et al., "Oxygen Pressure Leaching of Fe-Ni-Cu Sulfide
Concentrates at 110.degree. C.--Effect of Low Chloride Addition",
Hydrometallurgy Section, Jan. 5, 1976, pp. 117-125..
|
Primary Examiner: Phasge; Arun S.
Attorney, Agent or Firm: Oliff & Berridge, PLC
Parent Case Text
CROSS REFERENCE TO RELATED APPLICATION
This application is a continuation-in-part of U.S. patent
application Ser. No. 08/425,117 filed Apr. 21, 1995, .Iadd.now U.S.
Pat. No. 5,645,708, .Iaddend.the entire contents of which is
incorporated herein by reference which is a continuation-in-part of
Ser. No. 08/098,874 filed Jul. 29, 1993 now U.S. Pat. No.
5,431,788.
Claims
What is claimed is:
1. A process for the extraction of .[.a.]. .Iadd.at least one
.Iaddend.non-cuprous metal .Iadd.selected from the group consisting
of nickel and cobalt .Iaddend.from a metal ore or concentrate
.Iadd.containing the non-cuprous metal.Iaddend., comprising the
.[.step.]. .Iadd.steps .Iaddend.of
subjecting the .Iadd.metal .Iaddend.ore or concentrate
.Iadd.containing the non-cuprous metal selected from the group
consisting of nickel and cobalt .Iaddend.to pressure oxidation in
the presence of oxygen and an acidic solution containing halogen
ions and a source of bisulphate or sulphate ions to form a
.Iadd.product .Iaddend.solution of said non-cuprous metal, wherein
said source of bisulphate or sulphate ions is .Iadd.at least one
member .Iaddend.selected from the group consisting of sulphuric
acid and a metal sulphate .[.which.]. .Iadd.that
.Iaddend.hydrolyzes in said acidic solution.Iadd.,
recovering the non-cuprous metal from the product solution, and
recycling a portion of the product solution remaining after the
recovering step to the pressure oxidation step.Iaddend..
2. The process according to claim 1, wherein said halogen .[.is.].
.Iadd.ions are .Iaddend.selected from chlorine and bromine
.Iadd.ions.Iaddend...[.
3. The process according to claim 1, wherein the non-cuprous metal
is zinc and further comprising the steps of precipitating zinc from
said solution of non-cuprous metal in the form of a basic zinc
sulphate; separating the basic zinc sulphate from the remainder of
said solution; and leaching zinc from said basic zinc sulphate to
produce a solution of zinc ions..]..[.
4. The process according to claim 3, further comprising the steps
of extracting zinc from said solution of zinc ions by means of an
organic extractant; and stripping zinc from said extractant to
produce a concentrated solution of zinc ions for
electrowinning..]..[.
5. The process according to claim 3, wherein said solution of
non-cuprous metal also contains copper ions and further comprising
the step of removing said copper ions from solution prior to said
precipitation of zinc as the basic zinc sulphate..]..[.
6. The process according to claim 5, wherein the copper ions are
removed by solvent extraction..]..[.
7. The process according to claim 5, wherein the copper ions are
removed by selective precipitation of copper..]..[.
8. The process according to claim 1, wherein the non-cuprous metal
is nickel and further comprising the steps of precipitating nickel
from said solution of non-cuprous metal in the form of a basic
nickel sulphate; separating the basic nickel sulphate from the
remainder of said solution; and leaching nickel from said basic
nickel sulphate to produce a solution of nickel ions..]..[.
9. The process according to claim 8, wherein said solution of
non-cuprous metal also contains copper ions and further comprising
the step of removing said copper ions from solution prior to said
precipitation of nickel as the basic nickel sulphate..]..[.
10. The process according to claim 9, wherein the copper ions are
removed by selective precipitation of copper..]..[.
11. The process according to claim 1, wherein the non-cuprous metal
is cobalt and further comprising the steps of precipitating cobalt
from said solution of non-cuprous metal in the form of a basic
cobalt sulphate; separating the basic cobalt sulphate from the
remainder of said solution; and leaching cobalt from said basic
cobalt sulphate to produce a solution of cobalt ions..]..[.
12. The process according to claim 11, wherein said solution of
non-cuprous metal also contains copper ions and further comprising
the step of removing said copper ions from solution prior to said
precipitation of cobalt as the basic cobalt sulphate..]..[.
13. The process according to claim 12, wherein the copper ions are
removed by selective precipitation of copper..]..[.
14. The process according to claim 11, wherein the halogen ion
concentration is about 12 g/L..]..[.
15. The process according to claim 1, wherein the non-cuprous metal
is a mixture of nickel and cobalt and further comprising the steps
of precipitating nickel and cobalt from said solution of
non-cuprous metal in the form of basic nickel and cobalt salts;
separating the basic nickel and cobalt salts from the remainder of
said solution; and leaching nickel and cobalt from said basic
nickel and cobalt sulphates to produce a solution of nickel and
cobalt ions..]..[.
16. The process according to claim 15 further comprising the step
of separating said nickel and cobalt ions by selective solvent
extraction to produce separate nickel and cobalt solutions for
electrowinning..]..[.
17. The process according to claim 1, wherein the halogen
concentration is in the range of about 8 g/L to about 20
g/L..]..[.
18. The process according to claim 1, wherein said source of
bisulphate or sulphate ions is generated by the metal ore or
concentrate during said pressure oxidation..]..Iadd.
19. The process according to claim 2, wherein the halogen ions are
chloride ions and the concentration of chloride is in the range of
about 8 g/L to about 20 g/L..Iaddend..Iadd.
20. The process according to claim 19, wherein the chloride
concentration is about 12 g/L..Iaddend..Iadd.
21. The process according to claim 1, wherein the pressure
oxidation is carried out at a temperature between about 130.degree.
C. and 150.degree. C..Iaddend..Iadd.
22. The process according to claim 1, wherein the pressure
oxidation is carried out under a total oxygen and steam pressure of
about 690 kPa to about 1380 kPa..Iaddend..Iadd.
23. The process according to claim 1, wherein the non-cuprous metal
is nickel, and the recovering of the nickel comprises precipitating
the nickel from the product solution to form a nickel precipitate
and leaching nickel from the nickel precipitate to produce a nickel
solution..Iaddend..Iadd.
24. The process according to claim 23, wherein the product solution
also contains copper and the process further comprises removing the
copper from the product solution prior to the precipitating of the
nickel..Iaddend..Iadd.
25. The process according to claim 23, wherein the metal ore or
concentrate also contains copper that is leached into the product
solution during the pressure oxidation and the process further
comprises removing the copper from the product solution prior to
the precipitating of the nickel..Iaddend..Iadd.
26. The process according to claim 25, wherein the copper is
removed by solvent extraction..Iaddend..Iadd.
27. The process according to claim 25, wherein the copper is
removed by precipitation..Iaddend..Iadd.
28. The process according to claim 1, wherein the non-cuprous metal
is cobalt, and the recovering of the cobalt comprises precipitating
the cobalt from the product solution to form a cobalt precipitate
and leaching cobalt from the cobalt precipitate to produce a cobalt
solution..Iaddend..Iadd.
29. The process according to claim 28, wherein the product solution
also contains copper and the process further comprises removing the
copper from the product solution prior to the precipitating of the
cobalt..Iaddend..Iadd.
30. The process according to claim 28, wherein the metal ore or
concentrate also contains copper which is leached into the product
solution during the pressure oxidation and the process further
comprises removing the copper from the product solution prior to
the precipitating of the cobalt..Iaddend..Iadd.
31. The process according to claim 30, wherein the copper is
removed by solvent extraction..Iaddend..Iadd.
32. The process according to claim 1, wherein the at least one
non-cuprous metal is both nickel and cobalt, and the recovering of
the nickel and cobalt comprises precipitating the nickel and cobalt
from the product solution to form a nickel and cobalt precipitate
and leaching nickel and cobalt from the nickel and cobalt
precipitate to produce a nickel and cobalt
solution..Iaddend..Iadd.
33. The process according to claim 32, wherein the product solution
also contains copper and the process further comprises removing the
copper from the product solution prior to the precipitating of the
nickel and cobalt..Iaddend..Iadd.
34. The process according to claim 32, wherein the metal ore or
concentrate also contains copper that is leached into the product
solution during the pressure oxidation and the process further
comprises removing the copper from the product solution prior to
the precipitating of the nickel and cobalt..Iaddend..Iadd.
35. The process according to claim 34, wherein the copper is
removed by solvent extraction..Iaddend..Iadd.
36. The process according to claim 34, wherein the copper is
removed by precipitation..Iaddend..Iadd.
37. The process according to claim 32, wherein the process further
comprises separating the nickel and cobalt from the nickel and
cobalt solution by selective solvent extraction to produce separate
nickel and cobalt solutions for electrowinning..Iaddend..Iadd.
38. A process for the extraction of a non-cuprous metal from a
metal ore or concentrate containing the non-cuprous metal,
comprising the steps of
subjecting the metal ore or concentrate containing the non-cuprous
metal to pressure oxidation at a temperature of between about
130.degree. C. and 150.degree. C. in the presence of oxygen and an
acidic solution containing halogen ions and a source of bisulphate
or sulphate ions to form a product solution of said non-cuprous
metal, wherein said source of bisulphate or sulphate ions is at
least one member selected from the group consisting of sulphuric
acid and a metal sulphate that hydrolyzes in said acidic
solution,
recovering the non-cuprous metal from the product solution, and
recycling a portion of the product solution remaining after the
recovering step to the pressure oxidation step..Iaddend..Iadd.
39. A process for the extraction of at least one non-cuprous metal
selected from the group consisting of nickel and cobalt from a
metal ore or concentrate containing the non-cuprous metal,
comprising the steps of subjecting the metal ore or concentrate
containing the non-cuprous metal to pressure oxidation at a
temperature of between about 130.degree. C. and 150.degree. C. in
the presence of oxygen and an acidic solution containing halogen
ions and a source of bisulphate or sulphate ions to form a product
solution of said non-cuprous metal, wherein said source of
bisulphate or sulphate ions is at least one member selected from
the group consisting of sulphuric acid and a metal sulphate that
hydrolyzes in said acidic solution, and recovering the non-cuprous
metal from the product solution..Iaddend.
Description
FIELD OF THE INVENTION
This invention relates to the hydrometallurgical treatment of metal
ores or concentrates. In particular, it relates to the extraction
of metals from ores in the presence of halogen ions, such as
chloride ions. It also relates to the extraction of nickel and
cobalt from laterite ores.
BACKGROUND OF THE INVENTION
Hydrometallurgical treatment of copper sulphide ores, such as
chalcopyrite (CuFeS.sub.2), is problematical because the severe
conditions required in a pressure oxidation step for the effective
leaching of copper from these ores results in oxidation of the
sulphide in the ore to sulphate, resulting in the generation of
large amounts of acid which requires expensive neutralization.
Attempts have been made to render the sulphide concentrate
leachable under relatively milder conditions under which the
sulphide would only be oxidized to elemental sulphur and not all
the way through to sulphate. These attempts include the
pretreatment of the concentrate prior to the pressure oxidation
step to render the sulphide concentrate more readily leachable, and
the leaching of the concentrate in the presence of chloride ions,
such as described in U.S. Pat. No. 4,039,406. In this process, the
copper values in the concentrate are transformed into a solid basic
copper sulphate from which the copper values must then be
subsequently recovered, as described in U.S. Pat. No. 4,338,168. In
the process described in U.S. Pat. No. 4,039,406 a significant
amount (20-30%) of sulphide in the ore or concentrate is still
oxidized to sulphate, resulting in greater oxygen demand during the
pressure leach and the generation of sulphuric acid. This is
particularly unfavourable for low grade concentrates, where the
S/Cu ratio is high.
The present invention provides a process for the hydrometallurgical
extraction of copper and other metals in the presence of halogen
ions, such as chloride and bromide in solution.
SUMMARY OF THE INVENTION
According to the invention, there is provided a process for the
extraction of metal from a sulphide ore or concentrate, comprising
the steps of subjecting the ore or concentrate to pressure
oxidation in the presence of oxygen and an acidic chloride solution
to obtain a resulting pressure oxidation filtrate and an insoluble
basic metal sulphate salt, characterized in that the pressure
oxidation is conducted in the presence of a source of bisulphate or
sulphate ions which is selected from the group consisting of
sulphuric acid and a metal sulphate which hydrolyzes in the acidic
solution and wherein the amount of the source of bisulphate or
sulphate ions which is added contains at least the stoichiometric
amount of sulphate or bisulphate ions required to produce the basic
metal sulphate salt less the amount of sulphate generated in situ
in the pressure oxidation.
According to one particular embodiment of the invention, the
process further comprises the steps of recycling the pressure
oxidation filtrate to the pressure oxidation step; leaching the
basic metal sulphate salt produced by the pressure oxidation in a
second leaching with an acidic sulphate solution to dissolve the
basic metal salt to produce a leach liquor containing metal
sulphate in solution and a resulting solid residue; separating the
leach liquor from the solid residue; subjecting the leach liquor to
a solvent extraction process to produce metal concentrate solution
and a raffinate; and recycling the raffinate to the second leaching
step. In this embodiment, the pressure oxidation may be carried out
at a temperature of from about 115.degree. C. to about 175.degree.
C. The pressure oxidation may further be carried out under an
oxygen partial pressure of from about 50 psig (345 kPa) to about
250 psig (1725 kPa).
Also according to the invention, there is provided a process for
the extraction of a non-cuprous metal from a metal ore or
concentrate, comprising the step of subjecting the ore or
concentrate to pressure oxidation in the presence of oxygen and an
acidic solution containing halogen ions and a source of bisulphate
or sulphate ions to form a solution of said non-cuprous metal,
wherein said source of bisulphate or sulphate ions is selected from
the group consisting of sulphuric acid and a metal sulphate which
hydrolizes in said acidic solution. In this specification, M
represents the metal being extracted, such as copper, zinc, nickel
or cobalt.
The halogen concentration in the pressure oxidation filtrate, which
is recycled to the pressure oxidation step, is preferably
maintained in the range of from about 8 g/L to about 20 g/L,
preferably about 11 g/L to about 14 g/L, and more preferably at
about 12 g/L.
Reference is made to the use of chloride in the specification.
However, it will be appreciated that the chloride could be
substituted with bromide, if desired.
The term "non-cuprous metal" as used herein refers to a metal other
than copper.
The second leaching is preferably effected at a pH in the range of
from about 1.3 to about 2.2. It has been found that this maximizes
the solution of base metal and minimizes the solution of iron. More
preferably, the second leaching is effected in a pH range of from
about 1.6 to about 1.9.
The second leaching may be carried out at a temperature of from
about 20.degree. C. to about 70.degree. C., preferably, from about
35.degree. C. to about 45.degree. C.
For the second leaching, residence times of up to one hour or less,
such as 15 to 20 minutes, have been found to be adequate.
The raffinate may be split into a first portion comprising about
two-thirds of the raffinate and a second portion comprising about
one-third of the raffinate and the first portion may be recycled to
the second leaching and the second portion may be subjected to a
secondary solvent extraction to produce a secondary lixiviant and a
secondary raffinate. The secondary lixiviant may be used as
extractant in the solvent extraction of the leach liquor.
In another embodiment of the invention, the pressure oxidation is
carried out at a predetermined molar ratio of H.sup.30 /M, where
H.sup.+ represents the hydrogen ions in the acidic chloride
solution and M represents the metal in the ore or concentrate, so
that the pressure oxidation filtrate contains a first portion of
the metal in the ore or concentrate and the basic metal salt
contains a second portion of the metal in the ore or concentrate
and further comprising the steps of separating the pressure
oxidation filtrate and the basic metal salt; leaching the basic
metal salt in a second leaching step with an acidic sulphate
solution to dissolve the metal salt to produce a second metal
solution and a solid residue; and subjecting the pressure oxidation
filtrate and the second metal solution to solvent extraction to
produce concentrated metal solution for electrowinning of metal
therefrom.
The invention also extends to copper, zinc, nickel and cobalt
whenever produced by the process according to the invention.
Further objects and advantages of the invention will become
apparent from the description of preferred embodiments of the
invention below.
BRIEF DESCRIPTION OF THE DRAWINGS
FIG. 1 is a flow diagram of a hydrometallurgical copper extraction
process according to one embodiment of the invention, which is
suitable for the treatment of high grade copper ores or
concentrates.
FIG. 2 is a flow diagram of a hydrometallurgical copper extraction
process according to another embodiment of the invention, which is
suitable for the treatment of medium and lower grade copper ores or
concentrates.
FIG. 3 is a flow diagram of a hydrometallurgical copper extraction
process according to a further embodiment of the invention, which
provides for the extraction of zinc in addition to copper.
FIG. 4 is a flow diagram of a hydrometallurgical copper extraction
process according to another embodiment of the invention, which
provides for the extraction of nickel in addition to copper.
FIG. 5 is a flow diagram of a hydrometallurgical process for the
extraction of metals for a copper-nickel sulphide concentrate
according to another embodiment of the invention.
FIG. 6 is a flow diagram of a hydrometallurgical process for the
extraction of metals from a nickel-copper sulphide concentrate
according to another embodiment of the invention.
FIG. 7 is a flow diagram of a hydrometallurgical process for the
extraction of metals from a nickel laterite ore according to
another embodiment of the invention.
FIG. 8 is a flow diagram of a hydrometallurgical process for the
extraction of metals from a copper-zinc sulphide concentrate
according to another embodiment of the invention.
DETAILED DESCRIPTION OF PREFERRED EMBODIMENTS
The process according to the invention is flexible enough to treat
a range of copper concentrates in which the grade of copper varies
from low, i,e. about 15% copper or less, to high grade, i.e. about
35% copper or more.
Broadly, the process comprises a pressure oxidation stage, an
atmospheric leach stage, one or more solvent extraction stages and
an electrowinning stage. Different grades of concentrate require
different treatment in the pressure oxidation stage, requiring
different modes of operation. These modes of operation are termed
Mode A and Mode B, respectively. In Mode A, which is effective when
high grade copper ores are leached, copper is not leached in the
pressure oxidation stage. In Mode B, which is effective when medium
and low grade copper ores are leached, copper is leached in the
pressure oxidation stage.
Each of the two modes of operation will now be described in
turn.
Process Mode A
FIG. 1 is a flow diagram of Mode A. The process comprises a
pressure oxidation stage 12 in a pressure oxidation vessel or
autoclave, an atmospheric leach stage 14, primary and secondary
solvent extractant stages 16 and 18, respectively, and an
electrowinning stage 20.
In the pressure oxidation stage 12, all copper minerals are
converted to basic copper sulphate, CuSO.sub.4.2Cu(OH).sub.2. The
treatment is carried out with oxygen in the presence of an acidic
chloride solution. Oxygen, as well as HCl and H.sub.2 SO.sub.4 are
introduced into the autoclave for this purpose. The temperature in
the autoclave is about 130.degree.-150.degree. C. and the pressure
about 100-200 psig (1380 kPa). This is total pressure comprising
oxygen pressure plus steam pressure. The retention time is about
0.5-2.5 hours and the process is normally carried out in a
continuous fashion in the autoclave. However, the process can also
be carried out in a batch-wise fashion, if desired.
The solids content in the autoclave is maintained at about 12-25%,
i.e. 150-300 g/L solids as determined by the heat balance and
viscosity limitations.
The slurry produced in the autoclave is discharged through a series
of one or more flash tanks 22 to reduce the pressure to atmospheric
pressure and the temperature to 90.degree.-100.degree. C. The
liquid part of the slurry is referred to as the product solution
from the pressure oxidation stage 12 and is indicated by reference
numeral 21.
The slurry from the flash tank(s) 22 is filtered, as shown at 24,
and the resultant filter cake is washed thoroughly to remove
entrained liquor as much as possible.
The pressure oxidation filtrate from the filtration 24 is recycled
to the pressure oxidation stage 12 but there is a small bleed of
about 5%, as shown at 26. This bleed 26 is determined by the
concentration of the soluble metals in the ore or concentrate which
may dissolve during the pressure oxidation stage 12. The bleed 26
is treated at 28 with lime to remove metals such as zinc and
magnesium as solid residues, which are present in the copper
concentrate, and to counteract buildup of these metals in the
pressure oxidation circuit. The pressure oxidation circuit is the
circuit from the pressure oxidation stage 12 to the flash tank(s)
22 to the filtration 24 to the bleed 26 and back to the pressure
oxidation stage 12. It is indicated by reference numeral 23.
The bleed 26 is subject to a solvent extraction, as shown at 27,
prior to the bleed treatment 28. The solvent extraction 27 is
carried out by means of a suitable organic extractant to remove
copper from the bleed 26. This solvent extraction is associated
with the solvent extraction stages 16 and 18 and will be referred
to again when the latter two solvent extraction stages are
described.
Prior to the pressure oxidation stage 12, the copper concentrate is
first subjected to a regrind, as shown at 30, to reduce the
particle size to about 97% minus 325 mesh, which corresponds to P80
(80% passing) 15 micron. The regrind 30 is carried out in solution
recycled from the bleed treatment 28. Thus, the slurry from the
bleed treatment 28 is subjected to a liquid/solid separation, as
shown at 32, and the solution is recycled to the regrind 30 and the
zinc/magnesium bleed residue is discarded, as shown at 17.
The solution which is recycled to the regrind 30 is an alkaline
chloride liquor at about pH 10. Use of this liquor minimizes water
input into the pressure oxidation circuit 23 which is important in
maintaining heat balance and in preserving the chloride solution in
the pressure oxidation circuit 23 as much as possible.
As stated above, copper is not leached in the pressure oxidation
stage 12 but is converted to an insoluble basic copper salt. The
feed solution to the pressure oxidation stage 12, which is the
leach liquor being recycled from the filtration 24 is indicated by
reference numeral 25. Although there is copper present in the feed
solution 25, there is no additional copper leached, i.e. the
process is operated so that the copper concentration in the feed
solution 25 to the pressure oxidation stage 12 is equal to the
copper concentration in the product solution 21 from the pressure
oxidation stage 12. This is indicated as: .DELTA.[Cu.sup.2+
]=0.
The feed solution 25 to the pressure oxidation stage 12 contains
about 15 g/L Cu and 12 g/L Cl, together with about 30-55 g/L
sulphuric acid. The acid is added in the form of make up H.sub.2
SO.sub.4 (usually 93%). The product solution 21 from the pressure
oxidation stage 12 also contains about 15 g/L Cu and 11-12 g/L Cl
but is at about pH 3. There is substantially no acid left in the
product solution 21 as it is all consumed in the pressure oxidation
stage 12 to form the basic copper salt.
As referred to above, the liquid pressure feed 25 to the pressure
oxidation stage 12 is made up partly of recycled filtrate to which
H.sub.2 SO.sub.4 is added. The immediate effect of adding the acid
to the filtrate is to increase the acidity of the filtrate which is
fed to the autoclave for the pressure leaching stage 12, but the
most important effect, surprisingly, has been found to be that the
addition of the acid, or more specifically the sulphate ions,
actually suppresses the oxidation of sulphur emanating from the
concentrate in pressure oxidation stage 12.
Typically the oxidation of sulphur that is experienced if no acid
is added is about 25-30% of the feed sulphur in the concentrate, as
is the case with the process described in U.S. Pat. No. 4,039,406.
However, if acid is added, it has been found that the sulphur
oxidation to sulphate is reduced to about 5-10%. This improvement
has substantial beneficial effects on the hydrometallurgical
extraction process. The oxidation of sulphur to sulphate creates
addition costs in several ways, such as additional oxygen required
for the reaction, additional reagent required to neutralize the
acid so formed by the oxidation and provision must be made for heat
removal due to the oxidation of sulphur to sulphate which is very
exothermic. This actually limits the throughput of the autoclave in
which the pressure leaching stage 12 takes place.
The chemistry of the reaction in the pressure oxidation stage 12 is
believed to be altered by the addition of the acid as follows:
No acid addition:
With acid addition:
In both reactions, the copper is precipitated in the form of a
basic copper salt, which has been found to comprise mostly basic
copper sulphate.
In the first reaction it appears that the sulphate of the basic
copper sulphate is supplied by oxidation of the feed sulphur in the
concentrate, whereas in the second reaction it appears to be
supplied by the sulphate ions in the acid which is added to the
autoclave, thus obviating the need for the oxidation of sulphur to
sulphate. Thus, in the second reaction, there is a nett consumption
of sulphate ions to form the basic copper salt. The amount of
sulphuric acid needed to suppress sulphur oxidation has been found
experimentally to be about 25 to 75 grams per liter, depending on
the type of concentrate and the percentage solids in the
concentrate.
In actual test work, there is more sulphur oxidation than is
predicted by either reaction. The first reaction predicts one sixth
or 16.7% of the sulphur to be oxidized, whereas experimentally
about 25%-30% is found. With acid addition, experiments indicate
that about 2-16% sulphur is oxidized to sulphate, rather than the
zero oxidation that would be predicted if the second reaction as
written was the only reaction taking place. Therefore, these
reaction equations do not reflect exactly what is happening in the
pressure leaching stage 12 but are only an approximation.
Chloride is conserved as much as possible in the pressure oxidation
circuit 23 but typically about 3-10% chloride is lost per pass into
the solid product at the filtration 24. Thus, the chloride must be
made up by the addition of HCl or another source of chloride to
provide 12 g/L chloride in the feed solution 25. The chloride
losses are minimized by thorough washing of the solids from the
pressure oxidation stage 12 on the filter 24. The amount of wash
water is constrained by the requirement to maintain a water balance
in the pressure oxidation circuit 23. The only water loss from the
circuit 23 is in the steam 29 from the flashing step 22 and in the
filter cake after the filtration 24. Hence, the need to use the
recycled solution from the bleed treatment 28 to slurry up the
concentrate in the grinding step 30, and thus minimize fresh water
input from the concentrate to the pressure oxidation step 12.
It has been found to be advantageous to maintain at least 15 g/L Cu
in the product solution 21 from the pressure oxidation stage 12 so
as to counteract chloride loss in the form of solid basic copper
chloride, CuCl.sub.2.3Cu(OH).sub.2, which can occur if insufficient
copper is present in solution to allow basic copper sulphate to
form:
This reaction can be counteracted by the addition of sufficient
acid into the autoclave during the pressure oxidation stage 12 to
maintain at least enough copper in solution to satisfy the
stoichiometric requirements for Cl as CuCl.sub.2. For 12 g/L Cl in
solution, the stoichiometric amount of Cu is: ##EQU1##
Thus, 15 g/L Cu is a safe minimum to prevent a significant chloride
loss in the form of the basic copper salt.
On the other hand, the copper concentration in the product solution
21 from the pressure oxidation stage 12 should be kept as low as
possible to counteract the formation of CuS by the reaction of
elemental sulphur with aqueous copper sulphate. This reaction can
occur during the pressure oxidation stage 12 or in the slurry after
discharge from the autoclave but before the filtration step 24:
This reaction is particularly undesirable because CuS is insoluble
in the dilute acid conditions of the atmospheric leaching stage 14.
Thus, the copper is not recovered and this results in the loss of
copper to the final residue.
To counteract the formation of CuS it is necessary to keep the
copper concentration in the product solution 21 as low as possible,
i.e. below 30 g/L for some concentrates. The tendency to CuS
formation is apparently related to the type of concentrate being
treated, with the medium to high grade concentrates being more
susceptible to CuS formation. Thus, although a high copper
concentration in the product solution 21 does not present a problem
with the low grade concentrates, it cannot be tolerated with the
higher grade concentrates.
As is known to date, high grade concentrates, i.e. above 35%
copper, are best treated to produce as low a copper concentration
in the product solution 21 as possible, i.e. below 25 g/L Cu.
Given the need to maintain at least 15 g/L Cu in solution in the
pressure oxidation circuit 23, there is an optimum range of copper
concentration of from 15 to 25 g/L Cu for high grade concentrates.
With medium grade concentrates, the upper limit can be stretched
considerably and for low grade ore, the copper concentration does
not play a significant role.
The copper concentration in the pressure oxidation filtrate 29 can
be controlled simply by adding the required amount of acid into the
feed solution 25 to the pressure oxidation stage 12. More acid
results in a higher copper concentration due to the dissolution of
the basic copper sulphate:
The addition of about 1 g/L acid results in an increase in copper
concentration of about 1 g/L. The actual concentration of acid
required is determined empirically by comparing the assays of feed
solution 25 to the pressure oxidation stage 12 and the product
solution 21 from the pressure oxidation stage 12 to satisfy
.DELTA.[Cu.sup.2+ ]=0. The volume of solution in the circuit 23,
however, is determined by the heat balance.
The percentage by weight of solids in the feed of copper
concentrate slurry to the pressure oxidation stage 12 can be varied
at will. The weight of concentrate solid fed to the pressure
oxidation stage 12 is determined by the amount of copper to be
recovered. The weight of the solution is determined mainly by the
heat balance in the pressure oxidation stage 12.
The desired operating temperature in the pressure oxidation stage
12 is about 150.degree. C. and the heat must be supplied largely by
the heat of reaction of the sulphide minerals with the high
pressure oxygen in the autoclave. For high grade concentrates, such
as will be treated by the Process Mode A currently being described,
this means a relatively low S/Cu ratio and thus a smaller heat
production per tonne of copper treated in the autoclave. Much of
the heat evolved is due to oxidation, not of copper, but of the
other two main elements in the concentrate, iron and sulphur. If
the grade of the concentrate is high, then the ratio of S/Cu and
Fe/Cu is low, hence a lower heat production.
To reach operating temperature from a starting temperature of say
50.degree. to 80.degree. C., which is typical for the pressure
oxidation filtrate 29 which is recycled after the filtration 24, it
is necessary to control the amount of water that must be heated,
since this is the main heat sink in the pressure oxidation stage
12. It is impractical to cool or heat the slurry inside the
autoclave by indirect means, such as by means of heating or cooling
coils, because of rapid scale formation on all surfaces,
particularly heat exchangers, leading to very poor heat transfer
characteristics. Direct heating or cooling by injection of steam or
water is also impractical due to water balance considerations.
Therefore, it is required that the heat balance be maintained by
balancing heat production from reaction heat with the heat capacity
of the feed materials, i.e. the feed solution 25 being recycled and
the concentrate slurry. The main variable that can be controlled
here is the volume of the feed solution 25. This is one of the
distinguishing features between Modes A and B. In Process Mode B,
still to be described, the heat evolution is much greater,
expressed as heat per tonne of copper product. Therefore, it is
possible to use more solution volume in the feed 25 to the pressure
oxidation stage 12.
Once the solution volume is fixed, the acidity of the solution can
be determined, since the total mass of acid is determined by the
need to maintain .DELTA.[Cu.sup.2+ ]=0. Typically, for a high grade
concentrate, about 35-55 g/L acid will be required.
It has been found to be beneficial to add small concentrations of
certain surfactants which change the physical and chemical
characteristics of liquid elemental sulphur (S.degree.) in the
autoclave during the pressure oxidation stage 12. Surfactants such
as lignin sulphonate and quebracho added to the pressure oxidation
feed solution 25 in small amounts, i.e. 0.1 to 3 g/L can reduce the
viscosity of the liquid sulphur and also change the chemistry in
the autoclave.
Additions of surfactants can reduce the sulphur oxidation in ways
that are not well understood, but are beneficial to the process. It
is believed that this is due to lower viscosity, resulting in
lowered tendency for liquid sulphur and solids to be held up within
the autoclave, thus reducing the retention time for these
materials, and hence the reduced tendency for sulphur oxidation to
occur.
Also it has been found that more complete reaction of the copper
minerals takes place if surfactants are added, apparently because
of lower viscosity sulphur, which does not "wet" unreacted sulphide
minerals, and thus allows the desired reaction to proceed to
completion.
Reaction (5) describes how adding sulphuric acid to the pressure
oxidation feed 25 will control the copper concentration in the
pressure oxidation filtrate 29. The overall reaction for the
pressure oxidation with sulphuric acid addition for a chalcopyrite
ore is given by reaction (2) above.
A similar reaction can be written using CuSO.sub.4 as the source of
sulphide ions instead of H.sub.2 SO.sub.4 :
It is noteworthy that there are 3/2 moles of sulphate required as
copper sulphate in reaction (6) compared to one mole of sulphuric
acid in reaction (2). Therefore, if CuSO.sub.4 is to be used as the
source of sulphate ions instead of sulphuric acid, it is necessary
to use 1.5 times as many moles of CuSO.sub.4. To take this into
account, the inventor has developed the concept of Excess Sulphate
Equivalent, which allows the calculation of how much acid to add to
the pressure oxidation feed solution 25 in order to achieve a
target copper concentration and still take into account reaction
(6).
By taking reaction (6) into account, it is possible to calculate "a
priori" the amount of acid required for constant copper
concentration in the pressure oxidation filtrate 29. The concept of
Excess Sulphate Equivalent is helpful:
Excess Sulphate Equivalent is equal to the sulphate available in
the pressure oxidation feed solution 25 for formation of basic
copper sulphate during the pressure oxidation stage 12. The
sulphate available is that which is in excess of a defined Base
Level of CuSO.sub.4 and CuCl.sub.2.
Base Level of CuSO4 and CuCl.sub.2 is sufficient to support
chloride in solution at 12 g/L in the form of CuCl.sub.2 and, in
addition, about 4.3 g/L Cu as CuSO.sub.4. The concentration of
CuCl.sub.2 corresponding to 12 g/L chloride in solution is
134.5/71*12=22.7 g/L CuCl.sub.2, which contains 10.7 g/L Cu in
solution. The additional 4.3 g/L copper therefore means a total of
15 g/L copper combined as CuCl.sub.2 and CuSO.sub.4 in the Base
Level.
Sulphate available is then the total sulphate as CuSO.sub.4 less
the Base Level. For instance, if the total copper concentration is
28 g/L in the pressure oxidation filtrate 29, then the sulphate
available is 28-15=13 g/L Cu * 98/63.5=20 g/L H.sub.2 SO.sub.4 as
available sulphate from CuSO.sub.4.
Excess Sulphate Equivalent (ESE) is then calculated from the
available sulphate from CuSO.sub.4 by dividing by 1.5:
Thus, in the example of 28 g/L total copper concentration or 20 g/L
available sulphate from CuSO.sub.4, there is 20/1.5=13.3 g/L ESE
from CuSO.sub.4.
Finally, if the target free acid equivalent is, say, 52 g/L H.sub.2
SO.sub.4 in the pressure oxidation feed solution 25, then the
amount of acid required is 52 less the ESE (13.3 g/L) or 38.7 g/L
H.sub.2 SO.sub.4. This is the amount that must be added to the feed
solution 25 to the pressure oxidation stage 12 to produce a
constant copper concentration in the pressure oxidation filtrate
29, i.e. the Base Level of 15 g/L Cu.
Other reactions can be written using Fe.sub.2 (SO.sub.4).sub.3 and
ZnSO.sub.4 as the source of sulphate ions instead of H.sub.2
SO.sub.4. In the case of ZnSO.sub.4, the zinc is assumed to
hydrolyze to basic zinc sulphate, ZnSO.sub.4.3Zn(OH).sub.2, which
is a basic salt of Zn analogous to basic copper sulphate. These
reactions are given below as reactions (7) and (8).
The solids from the pressure oxidation stage 12 after the
filtration 24, are treated in the atmospheric leaching stage 14 at
about pH 1.5 to pH 2.0 using raffinate from the primary leaching
stage 16, which is acidic, to dissolve the basic copper sulphate.
The leaching 14 takes place at a temperature of about 40.degree. C.
for a retention time of about 15-60 minutes. The percentage solids
is typically about 5-15% or about 50-170 g/L, although it is
possible to operate the process outside this range.
During the atmospheric leaching stage 14, the basic copper salts
dissolve almost completely with very little of the iron present in
the concentrate going into solution.
Typically, the leach liquor 33 produced after the liquid/solid
separation 34 contains about 10-20 grams per litre copper,
depending on the percentage solids feed to the leach 14, with
0.1-1.0 g/L iron and about 0.1-1.0 g/L chloride. Much of this iron
and chloride are derived from the feed raffinate 37 rather than the
solids from pressure oxidation, i.e. they are recycled. Typically
about 0.1-0.2 g/L iron and chloride dissolve per pass.
The copper extraction has been found to be about 95-98% based on
the original feed to the pressure leaching stage 12. Iron
extraction to solution has been found to be less than about 1%.
The slurry 31 from the atmospheric leaching stage 14 is difficult
if not impossible to filter, but settles well. In view of the need
to wash the leach solids very thoroughly, the slurry 31 is
therefore pumped to a counter current decantation (CCD) wash
circuit, symbolically indicated as a solid/liquid separation 34 in
FIG. 1. In the CCD circuit 34, the solids are fed through a series
of thickeners with wash water added in the opposite direction. By
this method, the solids are washed and entrained solution removed.
About 3 to 5 thickeners (not shown) are required with a wash ratio
(water to solids) of about 5 to 7 to reduce entrained liquor down
to less than 100 ppm Cu in the final residue.
The thickener underflow from the last thickener is the final
residue stream 35 at about 50% solids. This can be treated for the
recovery of precious metals, such as gold and silver, or sent to
tailings. Precious metals may be recovered by known methods, such
as cyanidation. The main constituents of the stream 35 are hematite
and elemental sulphur, which may be recovered by flotation if
market conditions warrant.
The thickener overflow from the first thickener is the product
solution 33 which is fed to the primary solvent extraction stage
16, as shown. This solution contains about 12 g/L Cu, 1 g/L Cl and
0.5 g/L Fe.
The optimum copper concentration is determined by the ability of
the solvent extraction stage 16 to extract the maximum copper from
the solution 33. Since a fraction of about one-third of the
raffinate from the solvent extraction stage 16 is eventually
neutralized, it is important to minimize the copper content of this
raffinate.
Solvent extraction performs best on dilute copper solutions due to
the fact that a concentrated copper solution results in a higher
acid concentration in the raffinate which tends to lower extraction
efficiency. More concentrated solutions are, however, cheaper to
treat from a capital cost point of view, since the volume is less.
Above a certain point, though, the increased concentration does not
reduce the size of the solvent extraction unit, since (i) there is
a maximum organic loading and (ii) aqueous volume is generally kept
equal to organic volume for mixing purposes by means of aqueous
recycle. Therefore, the total volume of organic extractant and
aqueous solution is only determined by the volume of organic
extractant. The maximum organic loading and hence volume of organic
is determined by the concentration and characteristics of the
particular organic solvent selected. For the typical solvent, e.g.
LIX.TM. reagent from Henkel Corporation, the maximum loading per
pass at 40% volume concentration in diluent is about 12 g/L Cu.
Therefore, the product solution 33 also should contain about 12 g/L
Cu.
The copper is extracted from the product solution 33 from the CCD
thickener overflow in two stages of extraction in the primary
solvent extraction stage 16 to produce a raffinate 37 with about 20
g/L free acid and about 0.3 to 1 g/L Cu. Most of this raffinate 37
is recycled to the atmospheric leaching stage 14 but about 25 to
30% is surplus to the acid requirements of the atmospheric leaching
stage 14 and must be neutralized. This surplus 121 is split off as
shown at 36 and neutralized.
The neutralization is effected in two stages to maximize copper
recovery and to prevent possible environmental problems with the
neutralization residue due to copper content, i.e. the unrecovered
copper from the raffinate 37 will precipitate upon neutralization
and can then re-dissolve later, in a tailing pond, for example.
The first stage neutralization takes place at pH 2 to pH 3, as
shown at 38, using limerock, which is very economical as a reagent,
compared with lime. The neutralization product is filtered at 40
and the resultant solids are washed with water from the external
source 45. The solids, which are mainly gypsum and iron hydroxides,
are discarded, as shown at 41.
The filtrate 39 is sent to the secondary solvent extraction stage
18 for the recovery of residual copper values. The secondary
solvent extraction 18 benefits from the primary neutralization 38
and results in a very low copper concentration in the secondary
raffinate 43, typically from about 0.03 to 0.06 g/L Cu.
As indicated by the broken lines in FIG. 1, the secondary solvent
extraction stage 18 uses the same organic extractant as the primary
solvent extraction circuit 16. This is also tied in with the
solvent extraction 27 of the pressure oxidation filtrate bleed 26.
The organic extractant which is washed at 42 with wash water 122
from an external source 45, and stripped at 44 is recycled to the
secondary solvent extraction stage 18 and then passes to the
primary extraction stage 16. The stripped organic 125 is split to
pass a portion thereof to the solvent extraction 27. The raffinate
from the solvent extraction 27 is added to the loaded organic 123
from the solvent extraction 16 prior to the wash 42. The wash water
47 from the wash 42 is passed to the pressure oxidation filter 24,
to serve as a feed wash water onto the filter 24. The resultant
wash filtrate is added to the pressure oxidation filtrate 29, thus
recovering the copper and chloride content from the solvent
extraction wash water (47).
The raffinate 43 from the secondary solvent extraction stage 18 is
neutralized again in a secondary neutralization stage 46, this time
at pH 10 and filtered at 48 to remove all dissolved heavy metals,
producing a solution 51 which is used as wash water in the CCD
circuit 34 for washing the final leach residue 35. The solid
residue from the filtration 48 is discarded, as shown at 53.
Stripping of the loaded and washed organic at 44 is effected by
means of spent acid or electrolyte 55 from the electrowinning stage
20 to obtain a pure copper sulphate solution or pregnant
electrolyte 57 which is then passed to the electrowinning stage 20
for electrowinning in the usual way.
It can be seen that all solution streams in the process are thus
recycled and there are no solution effluents from the process. Only
solid residues are discarded from the process.
Process Mode B
FIG. 2 is a flow diagram of Mode B. The same reference numerals are
used to indicate stages or steps in the process which correspond
with those in the previous embodiment of FIG. 1. For example, the
pressure oxidation stage is again indicated by 12, the atmospheric
leach stage by 14, the electrowinning stage by 20, the flash
tank(s) by 22, the pressure oxidation filtration by 24, the bleed
treatment of the pressure oxidation filtrate 29 by reference
numeral 28, the grinding stage by reference numerals 30 and the CCD
wash circuit by reference numeral 34.
In this mode of the process, the pressure oxidation 12 is carried
out both to oxidize and to leach into solution most of the copper
contained in the feed concentrate. Typically about 85-90% of the
copper is leached into the solution, with only about 10-15% being
left in the residue as the basic copper sulphate.
The conditions of the pressure oxidation stage 12 in the autoclave
are similar to those in Process Mode A except that the percentage
solids is lower, i.e. 150-225 g/L.
In this mode of the process, .DELTA.[Cu.sup.2+ ] is typically 30 to
40 g/L Cu, i.e. the copper concentration is greater in the product
solution 21 from the pressure oxidation stage 12. The feed solution
25 to the pressure oxidation stage 12 typically contains 10-15 g/L
Cu and 12 g/L Cl, together with about 20 to 30 g/L sulphuric
acid.
In this mode, no sulphuric acid is added to the pressure oxidation
stage 12 from an external source, as is the case with the FIG. 1
embodiment. In this mode, the acid is obtained from recycle in the
process, i.e. by the recycle of the pressure oxidation filtrate 29.
The product solution 21 from the pressure oxidation stage 12
contains about 40 to 50 g/L Cu and 11 to 12 g/L Cl at about pH 2 to
2.5.
The copper leached into the product liquor 21 from pressure
oxidation stage 12 must be controlled so as to obtain the desired
distribution of copper between liquor (85 to 90%) and residue (10
to 15%). This distribution results in a small but important amount
of basic copper sulphate solids in the leach residue. The pH is a
convenient indicator of the presence of basic copper sulphate,
since it is a buffering agent. With strong copper sulphate
concentration in solution, a pH range of 2 to 2.5 indicates basic
copper sulphate. Below pH 2 almost all the basic copper sulphate
will be dissolved, whereas above pH 2.5, too much basic copper
sulphate is formed and insufficient copper is likely to be found in
the solution 21.
The primary method of control is the amount of acid in the feed
liquor 25 to the pressure oxidation stage 12. The acid level in
turn is controlled by the degree of neutralization of the raffinate
from solvent extraction of the pressure oxidation filtrate 29
raffinate described below. Usually, about 25 to 50% of the acid
must be neutralized, depending on the amount of acid that is
required.
The acid generated during the pressure oxidation stage 12 varies
from one concentrate to another and according to conditions
employed. If the concentrate produces a large amount of acid during
the pressure oxidation stage 12, then the feed solution 25 will
need less acid to achieve the desired result. The minimum copper
(from concentrate feed) that should go to liquor 21 is about 10%.
Below 10%, the pH drops sufficiently low so that iron
concentrations in the pressure oxidation filtrate 29 increase
rapidly. Normally, iron is about 10 to 50 ppm, but if pH is below 2
and basic copper sulphate in residue disappears, then iron can
increase to above 1 g/L fairly quickly. This is undesirable because
there are several impurity elements such as As and Sb which are
only removed from solution simultaneously with iron hydrolysis.
Therefore, absence of iron in solution is a good guarantee of low
impurity content in the pressure oxidation filtrate 29. Iron is
also an impurity itself that must be avoided in the electrowinning
circuit 20 as far as possible.
There is another factor, however, which places a maximum on Cu in
solution. It has been found surprisingly that certain concentrates
actually leach more completely if the copper concentration is
lower. This is believed to be due to either formation of secondary
CuS, as described above, or to some other phenomenon related to
poor oxidation characteristics of the primary mineral,
chalcopyrite, in high copper concentration solutions. It is found
that elemental 10 sulphur, produced during the reaction in the
pressure oxidation stage 12, can coat or actually encapsulate
unreacted chalcopyrite particles and hinder the access of reagents.
This results in poor copper recovery. The phenomenon is apparently
accentuated by high Cu levels in solution. It can be overcome or
mitigated by the use of surfactants, as described above. The
problem is more severe with some concentrates, particularly high
grade, than others. Therefore, for these concentrates it is
desirable to limit the copper concentration in the pressure
oxidation filtrate (i.e. greater than about 95%) over all. To do
this, it is necessary to have a substantial proportion of the
copper as basic copper sulphate, i.e. in solid residue from the
pressure oxidation stage 12 rather than the pressure oxidation
filtrate. Typically, 20-40% of copper may report to solids, if
necessary, to keep the copper concentration low enough to obtain
high copper recovery.
Higher grade concentrates exhibit the problem of low copper
recovery with high copper in solution. Therefore, an increasing
proportion of copper must report to solids as the grade increases.
Tests with three different concentrates illustrate this
relationship:
H*/Cu Cu Distribution % Total Conc. # % Cu Molar PO liquor PO
residue recovered 1 41 0.55 0 100 97.3 2 28 0.70 63 37 95.7 3 22
0.96 85 15 94.7
The H.sup.+ /Cu molar ratio refers to H.sup.+ in the feed acid and
Cu in the feed concentrate. The H.sup.+ in the feed acid is taken
to be all the protons available on complete dissociation of the
acid even if under existing conditions the acid is not completely
dissociated. The H.sup.+ shown in the table is optimum level found
by experiment to give the best results.
For concentrate #1, which was a high grade concentrate, the process
chosen is Mode A, where all of the copper reports to the leach
liquor and .DELTA.[Cu.sup.2+ ]=0. The H.sup.+ /Cu ratio is that
found which was necessary by experimentation to give the desired
result of .DELTA.[Cu.sup.2+ ]=0.
For concentrate #2, a medium grade concentrate, Mode B was chosen,
but with a substantial amount of the copper reporting to the solid
basic copper sulphate. This was achieved by keeping the H.sup.+ /Cu
ratio low enough so that not all of the copper dissolved into the
liquor.
For concentrate #3, a low grade concentrate, Mode B was also chosen
but in this case, the minimum amount of copper reported to the
residue, by adjusting the H.sup.+ /Cu ratio high enough.
The residue from the pressure oxidation stage 12 is leached 14 with
raffinate 37 returning from the solvent extraction 16 which is
dilute acid, at 3-10 g/L H.sub.2 SO.sub.4. Since most of the copper
from the pressure oxidation stage 12 reports to the pressure
oxidation filtrate 29 and only a small fraction of the pressure
oxidation residue, the resultant leach liquor 31 from the
atmospheric leach 14 is quite dilute in copper. In turn, this
produces a dilute raffinate 37 from the solvent extraction 16.
Typically, the atmospheric leach liquor 31 is 3-7 g/L Cu and 0.2 to
0.5 g/L Fe.
The slurry resulting from the atmospheric leaching stage 14 is
difficult to filter, as was the case with Mode A. Good liquid/solid
separation and washing, however, can be achieved as before using a
series of thickeners in a CCD arrangement 34. Wash water 51 is
provided by raffinate from the solvent extraction 16, which is
neutralized, as indicated at 46. This is similar as in Mode A. The
only major difference is the lower tenor of the solution 33 and the
reduced volume.
The solution 33 produced by the atmospheric leaching stage 14 is
subjected to the solvent extraction 16. The copper containing
solution 29 from the pressure oxidation stage 12, is subject to a
solvent extraction stage 50. There are, therefore, two solvent
extraction operations, i.e. 16 and 50, treating two different
streams of liquor 33 and 29, respectively. It is a feature of the
process according to the invention that the organic extractant used
for effecting the solvent extraction operations is common to both
solvent extractions 16 and 50.
As shown in FIG. 2, the stripped organic 125 coming from the common
stripping operation 44 is first introduced into the solvent
extraction circuit 16, which has the weakest copper concentration
in the aqueous feed stream 33 and therefore needs the organic
extractant to be as low as possible in loading to be effective.
The loaded organic 126 from solvent extraction 16 is then sent to
the solvent extraction 50 where it contacts the higher copper
concentration liquor 29. It is not necessary for the solvent
extraction 50 to achieve a high extraction ratio because the
raffinate 63 from this extraction is recycled to the pressure
oxidation stage 12, a shown. On the other hand, the raffinate 37
from the solvent extraction 16 is only partly recycled and part is
neutralized 46 to remove excess acid from the circuit. Therefore,
it is more important to achieve high copper recovery from the
solvent extraction 16.
The raffinate 37 from the solvent extraction 16 is split at 36 as
in Mode A, with about one-third 121 to the neutralization 46 and
two-thirds 120 recycled to the atmospheric leach stage 14. An
important difference from Mode A is that the raffinate 37 from
solvent extraction 16 is sufficiently low in copper, i.e. below 100
ppm, so that it is not necessary to have a secondary solvent
extraction stage before neutralization 46, as was the case in Mode
A. This is due to the lower copper concentration and solution
volume, allowing the solvent extraction 16 to be more
efficient.
The loaded organic 65 produced by the two solvent extraction
operations 16, 50 in series, is washed in two stages in counter
current fashion with dilute acidic aqueous solution 122, as shown
at 42. This is primarily to remove entrained aqueous solution from
the loaded organic 65 and in particular to reduce the chloride
content before the organic goes to stripping at 44. The amount of
wash water required is about 1-3% of the organic volume. The
resultant wash liquor 47 produced is recycled to the pressure
oxidation stage 12.
The washed organic 69 is stripped at 44 with spent electrolyte 55
from the electrowinning stage 20 to provide a pure copper solution
or pregnant electrolyte 57 for electrowinning in the usual way.
The raffinate 63 is split at 70 in two portions 72, 74 as
determined by the required molar ratio of H.sup.+ /Cu. The portion
72 is recycled to the pressure oxidation stage 12. The portion 74
is neutralized at pH 2 with limerock at 76 and filtered 78. The
solid residue is washed and discarded, as shown at 80. The filtrate
82 is recycled with the portion 72 to form the feed solution 25 to
the pressure oxidation stage 12.
A novel feature of the process, therefore, is the use of a common
organic to extract copper from two separate aqueous feed liquors.
This provides considerable economies in lower capital and operating
costs in the solvent extraction circuits. Also, it allows for the
use of copious amounts of water in the atmospheric leaching CCD
circuit, so that good washing can be achieved on the final residue
and yet still recover copper from such a dilute liquor.
It has been found that the degree of sulphur oxidation that occurs
in the pressure oxidation stage 12 is highly dependent on the type
of concentrate, such as grade and mineralogy of the concentrate
being treated, as well as the conditions of the pressure oxidation
stage 12. Certain concentrates exhibit considerably higher sulphur
oxidation, i.e. oxidation of the sulphur in the concentrate to
sulphate, and the effect is particularly marked with the low grade
concentrates with less than about 28% Cu by weight. The inventor
has found that the significance of this variation is not so much
the copper grade itself but the copper/sulphur ratio in the
concentrate. The main impurity elements in a copper concentrate are
iron and sulphur due to the fact that copper ores are generally
composed of chalcopyrite together with other minerals, particularly
pyrite FeS.sub.2 or pyrrholite FeS.
Process Mode B deals with the problem of excess sulphur oxidation
in the pressure oxidation stage 12 when lower grade concentrates
are used by deliberately dissolving 90% of the copper and
minimizing the formation of basic copper sulphate. The reaction for
chalcopyrite is:
The filtrate 29 from the pressure oxidation stage 12 thus contains
high levels of copper sulphate and copper chloride and this is
treated in the solvent extraction stage 50 to produce a pure copper
sulphate solution which goes to the electrowinning stage 20.
With reference to FIG. 3, a hydro-metallurgical process for
extraction of zinc in addition to copper is shown. The same
reference numerals are used to indicate stages or steps in the
process which correspond with those in the previous
embodiments.
The concentrate is re-ground 30 as in the case of the previous
embodiments.
The pressure oxidation of a mixed zinc-copper concentrate is
carried out in similar fashion as for the concentrate containing
only copper as in FIG. 2.
Zinc oxidizes as readily or more readily than copper does and is
more likely to report to the leach liquor 29 as opposed to the
pressure oxidation residue. This is because zinc hydrolyzes less
readily as basic zinc sulphate than copper does, i.e. at higher
pH.
Recovery of copper or zinc is not hampered by high solution tenors
apparently as was found for high grade copper concentrations.
Therefore, it is possible to have most of the copper and zinc
report to the pressure oxidation filtrate 29, i.e. as in Process
Mode B. Sulphur oxidation is low, so that the amount of acid
generated within the pressure oxidation stage 12 is low. Hence, to
obtain a high H.sup.+ /Cu ratio, it is necessary to recycle
virtually all of the acid from the solvent extraction stage 12 with
minimal neutralization. The feed acid may be as high as 75 g/L
H.sub.2 SO.sub.4 with about 10 g/L Cu, 5 g/L Zn and 12 g/L Cl.
The pressure oxidation filtrate 29 will contain both zinc and
copper in substantial concentrations dependent on the feed
concentrate composition. For a concentrate with 20% Cu and 5% Zn,
the pressure oxidation filtrate 29 may contain about 50 g/L Cu, 15
g/L Zn and 12 g/L Cl.
The pressure oxidation residue is leached 14 in the same way using
raffinate 37 from the solvent extraction 16 as shown, producing a
mixed Cu-Zn solution for feed to the solvent extraction circuits.
Zinc is extracted first and then copper.
There are two aqueous streams to be treated by solvent extraction
as in Process Mode B for copper concentrates. The pressure
oxidation filtrate 29 contains high tenors of Cu and Zn, whereas
the atmospheric leach solution 33 is weak in both elements.
The novel arrangement outlined for the solvent extraction circuit
as for the embodiments described above, is continued for the zinc
solvent extraction, that is, the weak liquor is contacted first
with organic extractant followed by the strong aqueous liquor. In
this case, there are two circuits, one for zinc and one for
copper.
It is possible to extract copper first followed by zinc, depending
on the choice of organic extractant and its relative affinity for
the two elements. The applicant has found that satisfactory results
can be obtained by using DEHPA (diethyl-hexyl phosphoric acid) as
the first extractant, which is selective towards zinc over copper.
Therefore, two DEHPA extractions 100 and 102 are done, the first
extraction 100 is on the weak liquor 33 and the second extraction
102 is on the stronger liquor 29 from the pressure oxidation stage
12, to recover zinc and leave the bulk of the copper in
solution.
The zinc extraction by DEHPA is hampered by poor extraction
characteristics in the presence of high acid concentrations. In
practice, this means that the extraction effectively stops at about
pH 1.4 or about 7-10 g/L H.sub.2 SO.sub.4. To deal with this
problem, an interstage neutralization 104 at pH 2 is included for
the zinc solvent extraction. Thus, the zinc solvent extraction
occurs in two stages, i.e. the stage 102 and a second stage 103
with the neutralization 104 in between. Each stage 102, 103 will
extract only 5-7 g/L zinc before being stopped by the resultant
acid concentration in the raffinate. By using interstage
neutralization 104, the total zinc extraction can be increased to
10 g/L Zn or more. The raffinate 97 from the first extraction stage
102 is neutralized to about pH 2 to 2.5 at 104 with inexpensive
limerock (CaCO.sub.3) to produce gypsum solids which are filtered
off at 98 and discarded. The filtrate 99 is then fed to the second
solvent extraction stage 103. The feed to the second stage is
typically 10 g/L Zn and 50 g/L Cu at a pH of 2 to 2.5. After
extraction, the second raffinate 124 is typically 5 g/L Zn, 50 g/L
Cu and 8 g/L acid.
For the solvent extraction circuit 16, zinc concentrations are low
enough so that this does not present a problem.
The optimum zinc content of the pressure oxidation filtrate 29 is
determined largely by the ability of the zinc solvent extraction
circuit to extract the zinc. Due to the fact that zinc is extracted
quite weakly by the available extractants (e.g. DEHPA), there is a
maximum of about 5-7 g/L Zn that can be extracted before the
reaction stops due to acid buildup in the raffinate. Further
extraction requires neutralization of the acid. With interstage
neutralization it is possible to extract much higher levels of Zn,
however, the interstage neutralization removes sulphate from the
circuit which must be replaced either by sulphur oxidation or
adding fresh acid to the pressure oxidation circuit 23.
One inter-neutralization stage is likely to be compatible with
sulphate balance, therefore it is preferable to keep the
.DELTA.[Zn.sup.2+ ], which is the zinc concentration in the
pressure oxidation filtrate 29 minus the zinc concentration in the
recycled raffinate 72, to about 10 g/L. Thus, if the feed acid to
pressure oxidation recycled as raffinate 72 from solvent extraction
contains 5 g/L Zn, then the product filtrate 29 from pressure
oxidation should contain about 15 g/L Zn. This restriction on
.DELTA.[Zn] distinguishes the process for Zn compared to Cu. The
greater extraction ability of Cu solvent extraction means that good
extraction of Cu can be achieved with much higher acid levels, up
to 75 g/L H.sub.2 SO.sub.4 in raffinate compared to only about 7-10
g/L for Zn. Hence Cu can be extracted from 50 g/L Cu feed
streams.
After extraction, the loaded organic 106 from the Zn (DEHPA)
circuit contains some Cu, as a result of imperfect selectivity of
the DEHPA extractant towards Zn, and simple entrainment of the
strong Cu liquor. Typically the Zn/Cu ratio in the loaded organic
106 from Zn solvent extraction is about 150 to 300:1. If not
removed, all of the Cu will be stripped along with the Zn during
solvent stripping 114, and thus will be stripped into the Zn
pregnant electrolyte 120 which is fed to Zn electrowinning 118. Zn
electrowinning requires a very pure pregnant electrolyte if it is
to produce satisfactory (pure) Zn cathode at reasonable current
efficiency. The Zn/Cu ratio must be about 100,000:1 in pregnant
electrolyte. Therefore it is essential to remove almost all of the
Cu either from the loaded organic 106 or later from the pregnant
electrolyte before electrowinning. It is much easier to purify the
loaded organic 106.
To remove this copper, several washing or treatment stages 106,
e.g. 3 to 10, typically 5, are needed. Washing is done with dilute
acidified zinc sulphate aqueous solution. The wash stages are
arranged in series, i.e. the treated organic exiting from the first
wash stage enters the second wash stage and so through all the
other stages until the organic exits the last stage. Some zinc is
washed out with the copper, therefore, it is necessary to minimize
the amount of wash water added and make use of several wash stages
arranged in counter current fashion instead.
The resultant wash liquor 110 produced is recycled to the
atmospheric leach circuit for recovery of copper and zinc
values.
After washing, the organic stream 112 from the DEHPA extraction is
ready for stripping 114 with spent electrolyte 116 from a zinc
electrowinning circuit 118. This produces a pregnant electrolyte
120 for electrowinning zinc at high current efficiency.
After the stripping 114 the extraction solvent is further stripped
131 for removal or iron prior to recycling of the extractant to the
solvent extraction 100. The stripping 131 is effected with HCl
makeup solution 133 which is introduced into the pressure oxidation
stage.
The raffinates 122, 124 from the zinc extractions with DEHPA are
each extracted with a selective copper extractant, such as LIX.TM.,
in solvent extractions 16 and 50, respectively.
The design of these two circuit 16, 50 is similar as in Process
Mode B with a common organic used first in the solvent extraction
16 and then in the solvent extraction 50. The loaded organic is
then washed and stripped as before as shown at 42 and 44,
respectively.
Neutralization requirements in the solvent extraction 50 circuit
are found to be low because of the prior neutralization in the zinc
circuit.
The raffinates from the LIX.TM. extractions are recycled as before
back to the pressure oxidation stage 12 and the atmospheric leach
stage 14, respectively.
With reference to FIG. 4, a hydrometallurgical extraction process
for recovery of nickel in addition to copper is shown.
The same reference numerals are used to indicate stages or steps in
the process which correspond with those in the previous
embodiments.
For nickel-copper concentrates, the process is very similar as for
zinc, except that the available solvent extraction agents are all
less selective toward nickel than copper. Therefore, the nickel
solvent extraction circuits 130, 132 both are positioned after the
respective copper solvent extraction circuits, 16, 50,
respectively.
The loaded nickel extractant 135 from the solvent extraction 132 is
subject to a wash 137 and then stripped 139 before being recycled
to the solvent extraction 130. The stripping 139 is effected with
spent electrolyte from the nickel electrowinning 140.
In addition, nickel extraction is sufficiently weak that in situ
neutralization with ammonia, for example, is required, as indicated
at 134 and 136, respectively. The ammonia must be recovered from
the respective raffinates by a lime boil process 138, for example,
and recycled.
It has been found that there is a limit to the amount of sulphur
oxidation that can be accommodated by the process Mode B. If the
sulphur oxidation is high enough and sufficient acid is generated
during pressure oxidation, there will be a surplus of acid left
over after pressure oxidation, even if no acid is added to the
feed, such as in the form of acidic raffinate. In this situation,
not only will all the copper in the concentrate be converted to
dissolved copper sulphate, but also some of the iron in the
concentrate will be solubilized by the surplus acid, e.g. as ferric
sulphate.
It is desirable that iron in the concentrate report to the pressure
oxidation residue as stable hematite, Fe.sub.2 O.sub.3, and not to
the solution, where it must be separated from the copper. Typical
concentrates have an Fe:Cu ratio of at least 1:1, and therefore the
efficient and complete elimination of Fe at an early stage is an
important aspect of the process. Other impurities such as arsenic,
antimony, etc., are also removed with iron by co-adsorption or
precipitation mechanisms.
It has been found that some concentrates, however, exhibit so much
sulphur oxidation (acid generation) that the acid-consuming
capacity of pressure oxidation is exceeded, and some iron is
leached into solution, even under process Mode B conditions. It is
a target of the process to produce a low-iron liquor, typically
with 0.05 g/L Fe. Some concentrates which have been tested have
produced pressure oxidation liquors with 1.0 to 12.0 g/L Fe.
Similarly the pH of the pressure oxidation liquor is normally
targeted to be in the range 2.0 to 3.5, corresponding to less than
1 g/L free acid, but concentrates tested have produced pressure
oxidation liquors with pH in the range 1.2-2.0, corresponding to 1
to 15 g/L free acid.
Accordingly, a further embodiment of the process has been
developed, termed "process Mode C" for the treatment of the above
concentrates, termed "Mode C" concentrates. Process Mode C will now
be described below.
Process Mode C
The Mode C concentrates that exhibit a strong tendency towards
sulphur oxidation and hence acid generation are those with a high
S:Cu ratio, or more generally S:M ratio, where M=base metals, such
as Cu, Zn, Ni, Co, Pb, etc., but not including Fe, which does not
consume acid.
Nickel or nickel/copper concentrates may often by Mode C, because
they are frequently low-grade, with S:M ratio often about 2:1 or
higher. Some copper or copper/gold concentrates are also Mode C, if
they are low grade because of high pyrite content. Some copper/zinc
concentrates have also been found to be high in pyrite and hence of
Mode C type as well.
In general there is a correlation between pyrite (FeS.sub.2)
content and the tendency toward Mode C type behaviour. However,
there are also exceptions to this trend, as not all pyrites react
in the same way. Some pyrites oxidize sulphur more readily than
others. In contrast, pyrrhotite (Fe.sub.7 S.sub.8) or the related
iron-zinc mineral sphalerite, (Zn, Fe)S, appear to result in much
less sulphur oxidation, and thus exhibit Process Mode A
behaviour.
Process Mode C is essentially a special case of Process Mode B,
with two key features.
First, all the raffinate 63 (FIG. 2) is neutralized, before
returning this stream to the pressure 12 oxidation, (i.e. none is
bypassed).
Secondly, the pressure oxidation slurry (before filtration of leach
residue) is subjected to an extra neutralization, the pressure
oxidation neutralization, to neutralize excess acid and precipitate
any Fe in solution at this time. The pressure oxidation
neutralization is done as hot as practical, once the slurry has
been discharged from the autoclave. The most convenient opportunity
is in the conditioning tank after flash let-down to atmospheric
pressure, when the slurry is at or near the boiling point of the
solution, i.e. about 100.degree. C.
Limerock is used for this purpose, to neutralize any surplus acid
in the pressure oxidation slurry and thus bring the pH up to about
3. Simultaneously, any dissolved Fe present in the Fe.sup.3+ state
will be precipitated, along with any As or Sb that may be
present.
The principal products of these reactions are precipitated gypsum
and iron hydroxides or basic salts. Since the pressure oxidation
neutralization is done before filtration, these solids are mixed in
with the leach residues already present in the pressure oxidation
slurry, containing mostly elemental sulphur, hematite, unreacted
sulphides (pyrite), and any gangue minerals (quartz, feldspars,
etc., which are largely unchanged by pressure oxidation). This
mixing is advantageous as no additional filtration step is
required, and the other solids aid in the filtration of the
pressure oxidation neutralization products, which might otherwise
tend to filter poorly.
The resultant slurry, now at pH 3 is filtered and the filter cake
carefully washed, as always, to remove entrained liquor (Cu, Cl) as
much as practical. The filter cake proceeds to atmospheric leaching
where any precipitated copper is leached as usual at about pH
1.5-1.8, and the resultant washed thoroughly in a CCD circuit. The
filtrate 29 from the pressure oxidation filtration is treated as in
Process Mode B for Cu removal by the solvent extraction stage 50,
producing a raffinate 63 that then goes to neutralization 76 as
before, and then recycled back to the pressure oxidation 12, but
without the raffinate split 70, as indicated above. Thus the
pressure oxidation cycle is completed.
The important aspects of the process according to the invention can
be summarized as follows:
(i) oxidize completely all base metals contained in sulphide
concentrates, e.g. copper, nickel, zinc and cobalt, as well as
iron; and
(ii) minimize the oxidation of sulphur to sulphate and maximize the
production of elemental sulphur; and
(iii) precipitate the metals oxidized during pressure oxidation as
the basic salt, e.g. basic copper sulphate; or
(iv) solubilize the metals oxidized during pressure oxidation, as
the sulphate compound, e.g. zinc sulphate or nickel sulphate.
Although the pressure oxidation is chloride catalyzed, it does not
use a strong chloride solution, e.g., only about 12 g/L is needed
which will support about 11 g/L Cu or Zn as the respective chloride
salt. If a higher concentration of metals is needed or produced, it
is as the sulphate salt. Thus, the pressure oxidation solutions are
generally mixtures of the sulphate and chloride salts, not pure
chlorides.
The process according to the invention can be used to process
concentrates containing nickel alone or in combination with copper
or cobalt. Similarly, copperzinc concentrates can be processed.
This is achieved by the correct use of sulphate or sulphuric acid
during pressure oxidation in the presence of a halogen, such as
chloride. Insufficient acid or sulphate increases sulphur
oxidation, which is undesirable, as well as reduces metal oxidation
and hence metal recovery. Excess acid solubilizes iron from the
pressure oxidation slurry and causes unnecessary expense in cost of
acid and neutralizing agent.
Copper-Nickel Concentrates
The process flowsheet is shown as FIG. 5. It is intended for
concentrates containing 3-25% Cu and 3-10% Ni, with Cu predominant.
Generally cobalt is present at a Ni:Co ratio of between 10:1 and
30:1, which corresponds to about 0.1 to 0.8% Co in concentrate.
The process essentially is variation of Process Mode B above, where
Cu reports primarily to the liquor during pressure oxidation,
rather than to the solid product. Acid must be supplied to pressure
oxidation to enable both the Ni and Cu to solubilize primarily as
sulphate. Typically about 20-30 g/L acid as H.sub.2 SO.sub.4 is
added to pressure oxidation feed solution. Chloride addition to
pressure oxidation is sufficient to maintain 12 g/L Cl, same as for
Cu concentrates. Conditions of temperature, pressure, etc., are
also similar as for Cu concentrates. Co solubilizes along with
Ni.
The pressure oxidation liquor is treated first for Cu solvent
extraction, to remove essentially of the Cu, and then Ni is
precipitated as Basic Nickel sulphate, after first reheating to 85
to 90.degree. C., using limestone. Co is precipitated along with Ni
as a basic cobalt salt.
The precipitated basic nickel/cobalt sulphate is then leached in an
ammoniacal solution recycled from Ni solvent extraction. The
resultant Ni/Co leach liquor is then treated first for Co removal
by solvent extraction using a reagent specific for Co such as
Cyanex 272, a proprietary phosphinic acid from Cyanamid Inc. The Co
raffinate is then treated for Ni recovery by another solvent
extraction reagent, LIX 84, a proprietary hydroxy-oxime from Henkel
Corp.
Finally, the Ni raffinate is recycled to the Ni/Co leach. There is
a bleed of this raffinate which is treated to recover Ammonium
Sulphate which otherwise would build up in the circuit. This is due
to the introduction of sulphate ions in the basic nickel sulphate
filter cake. Ammonia must be added to make up for the loss of
ammonia in the ammonium sulphate.
Nickel-Copper Concentrates
Nickel-copper concentrates have Ni as the predominant element and
will contain about 8-25% Ni and 3-10% Cu. The process flowsheet is
shown in FIG. 6. Conditions in pressure oxidation are essentially
the same as for Copper-nickel concentrates. The difference from
FIG. 3 lies in the treatment of the pressure oxidation slurry.
These concentrates generally behave as in Process Mode A, in which
Cu reports primarily to the solid phase after pressure oxidation.
This is accomplished by addition of limerock to the pressure
oxidation slurry to raise the pH to about pH4, before the slurry is
filtered. This has the effect of neutralizing excess acid in
pressure oxidation liquor; precipitating any Fe; and precipitating
any Cu.
The neutralized slurry is filtered and the filter cake sent to
atmospheric leach, labelled a "copper leach" which in turn produces
a leach liquor for extraction by Cu solvent extraction.
The neutralized solution is treated for Ni/Co recovery by
precipitation and solvent extraction as for copper-nickel
concentrates.
Nickel Laterite Ores
Nickel laterites do not concentrate by flotation as sulphides do
and therefore have to be treated as a whole ore. Typically they
contain 1.5-3.0% Ni, and 0.1-0.3% Co, with negligible Cu. An
important feature is Mg content which can be up to 20% Mg, as well
as substantial Fe content. The flowsheet is shown in FIG. 7.
The process is similar to that used for Nickel-copper sulphide
concentrates, except that the absence of Cu means that the leach
residue, after neutralization and filtration can be discarded as it
has negligible metal values in Cu. There are also important
differences in the conditions used in pressure oxidation:
Temperature and pressure are much higher at 225.degree. C./450 psig
O.sub.2, and much higher acidity at 100 to 200 g/L free acid in
feed liquor. Chloride content stays the same at about 12 g/L Cl.
Chloride in leach liquor may be supplied as MgCl.sub.2 or HCl.
The other main difference is the need for a Mg removal step. Mg
leaches almost quantitatively into solution during pressure
oxidation, resulting in typically 40 g/L Mg per pass. This can be
removed evaporation/crystallization for example as MgSO.sub.4.
Copper-Zinc Concentrates
Copper-Zinc concentrates with 20-25% Cu and 1-10% Zn are treated by
Process Mode B type flowsheet, as shown in FIG. 8.
It has been found that excellent extraction of Zn in pressure
oxidation can be achieved so long as enough acid is added to the
feed solution that the final pH of the slurry is below about pH 2.
Otherwise, the conditions are similar as for Cu concentrates, i.e.,
150.degree. C. 200 psig O.sub.2, 12 g/L Cl.
In Process Mode B flowsheets, the Cu is primarily solubilized
during PO, and must be extracted by solvent extraction 50 This
solvent extraction is operated in conjunction with Cu solvent
extraction that extracts Cu from the leach liquor coming from
atmospheric leach ("Cu Leach").
Zinc, having been solubilized during pressure oxidation along with
Cu, is precipitated from the Cu raffinate as Basic Zinc sulphate by
limestone at pH 6 and at 85.degree.-90.degree. C. The Zn ppt is
then leached by acid raffinate returning from Zn solvent extraction
circuit. The Zn leach liquor is then extracted by Zn solvent
extraction producing a loaded Zn organic (DEHPA). This organic
stream must be carefully purified of Cu, Co, Cd, Cl, etc., before
stripping with spent acid from electrowinning. The purification is
done by scrubbing the loaded organic using ZnSO.sub.4 aqueous
solution.
While only preferred embodiments of the invention have been
described herein in detail, the invention is not limited thereby
and modifications can be made within the scope of the attached
claims.
* * * * *