CA2697187C - Method for upgrading copper concentrate - Google Patents
Method for upgrading copper concentrate Download PDFInfo
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- CA2697187C CA2697187C CA2697187A CA2697187A CA2697187C CA 2697187 C CA2697187 C CA 2697187C CA 2697187 A CA2697187 A CA 2697187A CA 2697187 A CA2697187 A CA 2697187A CA 2697187 C CA2697187 C CA 2697187C
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- Canada
- Prior art keywords
- copper
- concentrate
- zinc
- leaching
- leach
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- 239000012141 concentrate Substances 0.000 title claims abstract description 62
- 239000010949 copper Substances 0.000 title claims abstract description 50
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 47
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 47
- 238000000034 method Methods 0.000 title claims abstract description 33
- 239000011701 zinc Substances 0.000 claims abstract description 38
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims abstract description 32
- 229910052725 zinc Inorganic materials 0.000 claims abstract description 32
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 14
- 235000011149 sulphuric acid Nutrition 0.000 claims abstract description 14
- 239000001117 sulphuric acid Substances 0.000 claims abstract description 10
- 238000002386 leaching Methods 0.000 claims description 39
- 239000003638 chemical reducing agent Substances 0.000 claims description 18
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 16
- BWFPGXWASODCHM-UHFFFAOYSA-N copper monosulfide Chemical compound [Cu]=S BWFPGXWASODCHM-UHFFFAOYSA-N 0.000 claims description 13
- 239000007789 gas Substances 0.000 claims description 13
- 239000007800 oxidant agent Substances 0.000 claims description 11
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 8
- 229910052742 iron Inorganic materials 0.000 claims description 8
- 239000001301 oxygen Substances 0.000 claims description 8
- 229910052760 oxygen Inorganic materials 0.000 claims description 8
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 claims description 6
- RWSOTUBLDIXVET-UHFFFAOYSA-N Dihydrogen sulfide Chemical compound S RWSOTUBLDIXVET-UHFFFAOYSA-N 0.000 claims description 5
- NUJOXMJBOLGQSY-UHFFFAOYSA-N manganese dioxide Chemical compound O=[Mn]=O NUJOXMJBOLGQSY-UHFFFAOYSA-N 0.000 claims description 4
- 230000001590 oxidative effect Effects 0.000 claims description 4
- 238000011084 recovery Methods 0.000 claims description 4
- HYHCSLBZRBJJCH-UHFFFAOYSA-M sodium hydrosulfide Chemical group [Na+].[SH-] HYHCSLBZRBJJCH-UHFFFAOYSA-M 0.000 claims description 4
- GRVFOGOEDUUMBP-UHFFFAOYSA-N sodium sulfide (anhydrous) Chemical compound [Na+].[Na+].[S-2] GRVFOGOEDUUMBP-UHFFFAOYSA-N 0.000 claims description 4
- 239000012286 potassium permanganate Substances 0.000 claims description 3
- 230000001105 regulatory effect Effects 0.000 abstract description 5
- 239000012535 impurity Substances 0.000 abstract description 3
- 238000002360 preparation method Methods 0.000 abstract description 2
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 description 7
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 description 4
- 238000004090 dissolution Methods 0.000 description 4
- 230000003647 oxidation Effects 0.000 description 4
- 238000007254 oxidation reaction Methods 0.000 description 4
- 239000002253 acid Substances 0.000 description 3
- 229910052951 chalcopyrite Inorganic materials 0.000 description 3
- DVRDHUBQLOKMHZ-UHFFFAOYSA-N chalcopyrite Chemical compound [S-2].[S-2].[Fe+2].[Cu+2] DVRDHUBQLOKMHZ-UHFFFAOYSA-N 0.000 description 3
- 239000000203 mixture Substances 0.000 description 3
- 229910017518 Cu Zn Inorganic materials 0.000 description 2
- PXHVJJICTQNCMI-UHFFFAOYSA-N Nickel Chemical compound [Ni] PXHVJJICTQNCMI-UHFFFAOYSA-N 0.000 description 2
- RAHZWNYVWXNFOC-UHFFFAOYSA-N Sulphur dioxide Chemical compound O=S=O RAHZWNYVWXNFOC-UHFFFAOYSA-N 0.000 description 2
- 238000006243 chemical reaction Methods 0.000 description 2
- 239000003153 chemical reaction reagent Substances 0.000 description 2
- 238000005188 flotation Methods 0.000 description 2
- AMWRITDGCCNYAT-UHFFFAOYSA-L hydroxy(oxo)manganese;manganese Chemical compound [Mn].O[Mn]=O.O[Mn]=O AMWRITDGCCNYAT-UHFFFAOYSA-L 0.000 description 2
- 229910000360 iron(III) sulfate Inorganic materials 0.000 description 2
- 229910052751 metal Inorganic materials 0.000 description 2
- 239000002184 metal Substances 0.000 description 2
- 239000007787 solid Substances 0.000 description 2
- 229910052950 sphalerite Inorganic materials 0.000 description 2
- 239000011787 zinc oxide Substances 0.000 description 2
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 description 2
- 239000011686 zinc sulphate Substances 0.000 description 2
- 235000009529 zinc sulphate Nutrition 0.000 description 2
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 description 1
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 description 1
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 239000003795 chemical substances by application Substances 0.000 description 1
- 150000001805 chlorine compounds Chemical class 0.000 description 1
- 229910017052 cobalt Inorganic materials 0.000 description 1
- 239000010941 cobalt Substances 0.000 description 1
- GUTLYIVDDKVIGB-UHFFFAOYSA-N cobalt atom Chemical compound [Co] GUTLYIVDDKVIGB-UHFFFAOYSA-N 0.000 description 1
- 239000013065 commercial product Substances 0.000 description 1
- 238000010276 construction Methods 0.000 description 1
- 239000000428 dust Substances 0.000 description 1
- -1 ferrous metals Chemical class 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- 239000007788 liquid Substances 0.000 description 1
- 229910052960 marcasite Inorganic materials 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 235000010755 mineral Nutrition 0.000 description 1
- 229910052759 nickel Inorganic materials 0.000 description 1
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 description 1
- 229910052683 pyrite Inorganic materials 0.000 description 1
- 238000009853 pyrometallurgy Methods 0.000 description 1
- 238000000926 separation method Methods 0.000 description 1
- 229910052979 sodium sulfide Inorganic materials 0.000 description 1
- 150000004763 sulfides Chemical group 0.000 description 1
- 229910021653 sulphate ion Inorganic materials 0.000 description 1
- 235000010269 sulphur dioxide Nutrition 0.000 description 1
- 239000004291 sulphur dioxide Substances 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 1
- 229910000368 zinc sulfate Inorganic materials 0.000 description 1
- 229910052984 zinc sulfide Inorganic materials 0.000 description 1
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0002—Preliminary treatment
- C22B15/0004—Preliminary treatment without modification of the copper constituent
- C22B15/0008—Preliminary treatment without modification of the copper constituent by wet processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Landscapes
- Chemical & Material Sciences (AREA)
- Engineering & Computer Science (AREA)
- Organic Chemistry (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Manufacturing & Machinery (AREA)
- Inorganic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geochemistry & Mineralogy (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to a method for the preparation of a copper concentrate suitable for pyrometallurgical treatment from a concentrate in which zinc is present as an impurity. In accordance with the method, zinc is leached from the concentrate with a solution containing sulphuric acid in oxidising conditions at atmospheric temperature and pressure. At the end of the leach, the conditions are regulated to be such that any copper that may have dissolved will be precipitated.
Description
METHOD FOR UPGRADING COPPER CONCENTRATE
FIELD OF THE INVENTION
The invention relates to a method for the preparation of a copper concentrate suitable for pyrometallurgical treatment from a concentrate in which zinc is present as an impurity. In accordance with the method, zinc is leached from the concentrate with a solution containing sulphuric acid in oxidising conditions at atmospheric temperature and pressure. At the end of the leach, the conditions are regulated to be such that any copper that may have io dissolved will be precipitated.
BACKGROUND OF THE INVENTION
Complex copper sulphide concentrates contain other non-ferrous metals besides copper, such as zinc. When sulphide concentrate intended for ts pyrometallurgical recovery of copper contains zinc, for example, it complicates the process. In a pyrometallurgical process, the zinc contained in the concentrate evaporates and enters the fine dust as zinc oxide, from which it should be recovered separately. Current smelters charge a penalty for impurities in concentrate, for instance the zinc content of copper 20 concentrate should be below 3%.
Many ways of separating zinc from copper concentrate are known in the prior art. Separation using selective flotation is described in many patent publications such as e.g. US 4,279,867, JP 59 160558 and RU 2135298.
25 However, from time to time the mineralogy of the ore is such that good results cannot be achieved even with selective flotation.
Another method is to treat the concentrate first at such a temperature that the zinc vaporises and is recovered as zinc oxide, after which the copper 30 concentrate can be processed further in the desired way. This kind of method is described in e.g. patent CN 1288966. If the further processing of the concentrate takes place pyrometallurgically, the method requires the construction of two furnace units. Additionally, one disadvantage is the formation of sulphur dioxide, which necessitates its own processing equipment.
US patent publication 4,260,588 describes a way to handle the upgrading of a complex copper sulphide concentrate. There the concentrate is leached with a copper-containing chloride solution at an elevated temperature and pressure at a pH value of about 3. As a result of leaching, a sulphidic concentrate containing copper and iron is obtained and the other non-ferrous io metals contained in the complex concentrate such as zinc, nickel and cobalt are recovered from the leach. The method appears fairly complicated and costly. Another fact that could be regarded as a drawback is that the chlorides need to be removed from the concentrate meticulously, so that they do not cause problems in further processing of the concentrate.
PURPOSE OF THE INVENTION
The purpose of the invention is to achieve a method for upgrading copper sulphide concentrate containing zinc by removing the zinc from the concentrate, so that the concentrate is suitable for the pyrometallurgical 2o recovery of copper.
SUMMARY OF THE INVENTION
The invention relates to a method for upgrading copper sulphide concentrate containing zinc by removing the zinc from the concentrate, so that the concentrate is suitable for the pyrometallurgical recovery of copper. It is typical of the method that zinc is removed from the concentrate by leaching the concentrate with a solution containing sulphuric acid at atmospheric pressure and temperature, firstly in oxidising conditions and that before the end of the leach, a reducing agent is fed into the leaching stage to precipitate out the dissolved copper.
FIELD OF THE INVENTION
The invention relates to a method for the preparation of a copper concentrate suitable for pyrometallurgical treatment from a concentrate in which zinc is present as an impurity. In accordance with the method, zinc is leached from the concentrate with a solution containing sulphuric acid in oxidising conditions at atmospheric temperature and pressure. At the end of the leach, the conditions are regulated to be such that any copper that may have io dissolved will be precipitated.
BACKGROUND OF THE INVENTION
Complex copper sulphide concentrates contain other non-ferrous metals besides copper, such as zinc. When sulphide concentrate intended for ts pyrometallurgical recovery of copper contains zinc, for example, it complicates the process. In a pyrometallurgical process, the zinc contained in the concentrate evaporates and enters the fine dust as zinc oxide, from which it should be recovered separately. Current smelters charge a penalty for impurities in concentrate, for instance the zinc content of copper 20 concentrate should be below 3%.
Many ways of separating zinc from copper concentrate are known in the prior art. Separation using selective flotation is described in many patent publications such as e.g. US 4,279,867, JP 59 160558 and RU 2135298.
25 However, from time to time the mineralogy of the ore is such that good results cannot be achieved even with selective flotation.
Another method is to treat the concentrate first at such a temperature that the zinc vaporises and is recovered as zinc oxide, after which the copper 30 concentrate can be processed further in the desired way. This kind of method is described in e.g. patent CN 1288966. If the further processing of the concentrate takes place pyrometallurgically, the method requires the construction of two furnace units. Additionally, one disadvantage is the formation of sulphur dioxide, which necessitates its own processing equipment.
US patent publication 4,260,588 describes a way to handle the upgrading of a complex copper sulphide concentrate. There the concentrate is leached with a copper-containing chloride solution at an elevated temperature and pressure at a pH value of about 3. As a result of leaching, a sulphidic concentrate containing copper and iron is obtained and the other non-ferrous io metals contained in the complex concentrate such as zinc, nickel and cobalt are recovered from the leach. The method appears fairly complicated and costly. Another fact that could be regarded as a drawback is that the chlorides need to be removed from the concentrate meticulously, so that they do not cause problems in further processing of the concentrate.
PURPOSE OF THE INVENTION
The purpose of the invention is to achieve a method for upgrading copper sulphide concentrate containing zinc by removing the zinc from the concentrate, so that the concentrate is suitable for the pyrometallurgical 2o recovery of copper.
SUMMARY OF THE INVENTION
The invention relates to a method for upgrading copper sulphide concentrate containing zinc by removing the zinc from the concentrate, so that the concentrate is suitable for the pyrometallurgical recovery of copper. It is typical of the method that zinc is removed from the concentrate by leaching the concentrate with a solution containing sulphuric acid at atmospheric pressure and temperature, firstly in oxidising conditions and that before the end of the leach, a reducing agent is fed into the leaching stage to precipitate out the dissolved copper.
It is typical of the method accordant with the invention that the sulphuric acid concentration at the end of the leaching stage is regulated to the region of 0-100 g/I, preferably 5 - 25 g/I.
It is also typical of the method accordant with the invention that an oxidising agent is fed into the leaching stage.
According to one embodiment of the invention, the oxidising agent used is a gas, which is at least one of the following: air, oxygen-enriched air and io oxygen.
According to another embodiment of the invention, the oxidising agent used is at least one of the following: hydrogen peroxide, manganese dioxide and potassium permanganate.
It is also typical of the method accordant with the invention that the redox potential is adjusted so that there is a minimum of 10 mg/I of copper at the start of the leach and that at the end of the leach there is no copper in the solution.
According to one embodiment of the invention, the reducing agent to be fed into the end of the leaching stage is a reducing gas. The gas is typically hydrogen sulphide.
According to another embodiment of the invention, the reducing agent to be fed into the end of the leaching stage is sodium hydrosulphide or sodium sulphide.
According to yet another embodiment of the invention, the reducing agent to 3o be fed into the end of the leaching stage is copper sulphide concentrate.
It is also typical of the method accordant with the invention that an oxidising agent is fed into the leaching stage.
According to one embodiment of the invention, the oxidising agent used is a gas, which is at least one of the following: air, oxygen-enriched air and io oxygen.
According to another embodiment of the invention, the oxidising agent used is at least one of the following: hydrogen peroxide, manganese dioxide and potassium permanganate.
It is also typical of the method accordant with the invention that the redox potential is adjusted so that there is a minimum of 10 mg/I of copper at the start of the leach and that at the end of the leach there is no copper in the solution.
According to one embodiment of the invention, the reducing agent to be fed into the end of the leaching stage is a reducing gas. The gas is typically hydrogen sulphide.
According to another embodiment of the invention, the reducing agent to be fed into the end of the leaching stage is sodium hydrosulphide or sodium sulphide.
According to yet another embodiment of the invention, the reducing agent to 3o be fed into the end of the leaching stage is copper sulphide concentrate.
It is typical of the method accordant with the invention that there is a minimum of 0.1 g/I of iron and at least 10 mg/1 of copper in the solution at the start of the leach.
The essential features of the invention will be made apparent in the attached claims.
DETAILED DESCRIPTION OF THE INVENTION
In the method accordant with the invention, a copper sulphide concentrate to containing zinc is routed to leaching, with the aim of leaching such an amount of zinc from the concentrate that the concentrate is suitable for pyrometallurgical processing. Leaching is sulphate-based and occurs in atmospheric conditions. Atmospheric conditions means leaching that takes place in unpressurised reactors and at a temperature that is at the most 1s 105 C. The copper sulphide concentrate is mainly a primary copper sulphide such as chalcopyrite and does not substantially contain secondary sulphides.
Chalcopyrite does not dissolve substantially in conditions where zinc is leached from the concentrate.
20 Zinc-containing copper sulphide concentrate is fed into a solution containing sulphuric acid, in which the acid concentration is regulated to be in the range of 0 - 100 g/I at the end of the leach and preferably 5 - 25 g/l. At the beginning of the leach an oxidising agent is fed into the solution, said agent being for example an oxidising gas such as air, oxygen-enriched air or 25 oxygen. The oxidising agent may also be something other than gas, such as hydrogen peroxide, potassium permanganate or manganese oxide.
It is also advantageous for leaching that the acid-containing solution contains a small amount of dissolved iron and copper. A sufficient amount of iron 3o allows the zinc to dissolve at a sufficient rate i.e. it improves the dissolution kinetics. Preferably,there is at least 0.1 g/I of iron, typically over 5 g/l.
There is at least 10 mg/I of copper in the solution. The copper in the solution catalyses the oxidation of the iron. The oxidised iron participates in the dissolution of zinc according to reaction (2) below:
2FeSO4 + H2SO4 + 0.502 4 Fe2(SO4)3 + H20 (1) 5 ZnS + Fe2(SO4)3 4 ZnSO4 + 2FeSO4 + S (2) It has been found in the tests carried out that during the time that the zinc dissolves, the dissolution of copper is, however, fairly minor. Nevertheless, copper dissolution is not a problem, because it is precipitated back at the end io of the leach.
The redox potential of the leach is adjusted so that at the beginning of the leach there is at least 10 mg/i of copper in the solution and at the end of the leach there is no copper in the solution.
It is characteristic of the selective leaching method accordant with the invention that the progress of the leach is monitored so that the acid concentration of the solution is regulated to be within the desired region at the end of the leach. At the same time the amount of oxidising agent required to dissolve the zinc is determined, and is fed into the leaching reactor at the start of the leach. The leaching reactor is preferably an agitated reactor.
When the required amount of oxidising agent has been fed, the reactions continue further, even though the feed of oxidant has finished. When the zinc content of the concentrate has fallen to the desired level, part of the copper may also have dissolved. Dissolved copper is precipitated at the end of the leach by means of some reducing agent. The reducing agent may be solid, liquid or gaseous such as hydrogen sulphide. Other reducing agents are for instance sodium hydrosulphide NaHS, sodium sulphide Na2S or copper sulphide concentrate itself, which acts as a slow reductant. The result obtained from leaching is copper concentrate, with a zinc content that is suitably low for the pyrometallurgical processing of concentrate. The zinc sulphate solution is processed in the desired way in order to fabricate a commercial product.
The aim is to keep the total time and costs of leaching in process conditions at the most economically viable level. Thus the long leaching time demanded by the slowly-reducing reductant and on the other hand the possibly higher cost of a quick-acting reductant and the reagent consumption are taken into account (i.e. time vs. reagent consumption).
io EXAMPLES
Example 1 Copper sulphide concentrate was leached in a reactor that had a volume of L. The concentrate was composed of the following minerals: CuFeS2, FeS2, ZnS. The grain size of the concentrate was do.5 = 10 pm and do.9 = 35 pm. The composition of the concentrate was as follows:
Table I
Cu Zn Fe Pb S Znaq, Soi.
% % % % % %
18.4 7.3 28.4 1.5 40.0 2.3 Leaching was performed at atmospheric pressure and a temperature of 95 C
2o and the leaching time was 9 h. The solution contained 9.6 g/l of ferrous iron and 1 g/l of copper. The solids content of the solution was 200 g/l and at the beginning of the leach the H2SO4 concentration of the solution was 50.8 g/I.
The oxygen-containing gas used was oxygen, which was fed into the reactor for 3 h after which time the feed was stopped. Fresh concentrate was used for reduction, and 1.1 kg was fed into the reactor when 6 - 7 hours had passed since the start of the test. As shown in Table 2 below, the zinc content of the concentrate had fallen to a value of 1.4 % and the copper content had risen to 21.4%. According to the balance calculation, the amount of copper lost in leaching was below 10%.
The essential features of the invention will be made apparent in the attached claims.
DETAILED DESCRIPTION OF THE INVENTION
In the method accordant with the invention, a copper sulphide concentrate to containing zinc is routed to leaching, with the aim of leaching such an amount of zinc from the concentrate that the concentrate is suitable for pyrometallurgical processing. Leaching is sulphate-based and occurs in atmospheric conditions. Atmospheric conditions means leaching that takes place in unpressurised reactors and at a temperature that is at the most 1s 105 C. The copper sulphide concentrate is mainly a primary copper sulphide such as chalcopyrite and does not substantially contain secondary sulphides.
Chalcopyrite does not dissolve substantially in conditions where zinc is leached from the concentrate.
20 Zinc-containing copper sulphide concentrate is fed into a solution containing sulphuric acid, in which the acid concentration is regulated to be in the range of 0 - 100 g/I at the end of the leach and preferably 5 - 25 g/l. At the beginning of the leach an oxidising agent is fed into the solution, said agent being for example an oxidising gas such as air, oxygen-enriched air or 25 oxygen. The oxidising agent may also be something other than gas, such as hydrogen peroxide, potassium permanganate or manganese oxide.
It is also advantageous for leaching that the acid-containing solution contains a small amount of dissolved iron and copper. A sufficient amount of iron 3o allows the zinc to dissolve at a sufficient rate i.e. it improves the dissolution kinetics. Preferably,there is at least 0.1 g/I of iron, typically over 5 g/l.
There is at least 10 mg/I of copper in the solution. The copper in the solution catalyses the oxidation of the iron. The oxidised iron participates in the dissolution of zinc according to reaction (2) below:
2FeSO4 + H2SO4 + 0.502 4 Fe2(SO4)3 + H20 (1) 5 ZnS + Fe2(SO4)3 4 ZnSO4 + 2FeSO4 + S (2) It has been found in the tests carried out that during the time that the zinc dissolves, the dissolution of copper is, however, fairly minor. Nevertheless, copper dissolution is not a problem, because it is precipitated back at the end io of the leach.
The redox potential of the leach is adjusted so that at the beginning of the leach there is at least 10 mg/i of copper in the solution and at the end of the leach there is no copper in the solution.
It is characteristic of the selective leaching method accordant with the invention that the progress of the leach is monitored so that the acid concentration of the solution is regulated to be within the desired region at the end of the leach. At the same time the amount of oxidising agent required to dissolve the zinc is determined, and is fed into the leaching reactor at the start of the leach. The leaching reactor is preferably an agitated reactor.
When the required amount of oxidising agent has been fed, the reactions continue further, even though the feed of oxidant has finished. When the zinc content of the concentrate has fallen to the desired level, part of the copper may also have dissolved. Dissolved copper is precipitated at the end of the leach by means of some reducing agent. The reducing agent may be solid, liquid or gaseous such as hydrogen sulphide. Other reducing agents are for instance sodium hydrosulphide NaHS, sodium sulphide Na2S or copper sulphide concentrate itself, which acts as a slow reductant. The result obtained from leaching is copper concentrate, with a zinc content that is suitably low for the pyrometallurgical processing of concentrate. The zinc sulphate solution is processed in the desired way in order to fabricate a commercial product.
The aim is to keep the total time and costs of leaching in process conditions at the most economically viable level. Thus the long leaching time demanded by the slowly-reducing reductant and on the other hand the possibly higher cost of a quick-acting reductant and the reagent consumption are taken into account (i.e. time vs. reagent consumption).
io EXAMPLES
Example 1 Copper sulphide concentrate was leached in a reactor that had a volume of L. The concentrate was composed of the following minerals: CuFeS2, FeS2, ZnS. The grain size of the concentrate was do.5 = 10 pm and do.9 = 35 pm. The composition of the concentrate was as follows:
Table I
Cu Zn Fe Pb S Znaq, Soi.
% % % % % %
18.4 7.3 28.4 1.5 40.0 2.3 Leaching was performed at atmospheric pressure and a temperature of 95 C
2o and the leaching time was 9 h. The solution contained 9.6 g/l of ferrous iron and 1 g/l of copper. The solids content of the solution was 200 g/l and at the beginning of the leach the H2SO4 concentration of the solution was 50.8 g/I.
The oxygen-containing gas used was oxygen, which was fed into the reactor for 3 h after which time the feed was stopped. Fresh concentrate was used for reduction, and 1.1 kg was fed into the reactor when 6 - 7 hours had passed since the start of the test. As shown in Table 2 below, the zinc content of the concentrate had fallen to a value of 1.4 % and the copper content had risen to 21.4%. According to the balance calculation, the amount of copper lost in leaching was below 10%.
The results show that the concentrate acts as a slow reductant, and if leaching had continued, the copper would have gradually been precipitated back. Concentrate is without doubt the most cost-efficient reducing agent, if it is possible to perform reduction for a long period of time. The redox potential was measured with Pt electrodes vs. Ag/AgCI electrode.
Table 2 Solution Concentrate Time Fe3+ Cu Fe2+ Zn H2SO4 Redox Cu Fe Pb Zn S
h /I /I /I g/I /I mV % % % % %
0 0.53 1.0 9.6 0.0 50.8 412 18.4 28.4 1.5 7.3 40.0 1 0.90 4.9 13.6 9.4 29.3 424 20.0 31.6 2.0 3.3 44.6 3 2.58 8.4 16.5 12.1 14.2 433 18.3 31.3 2.6 1.5 46.8 5.5 1.19 7.8 17.2 13.4 15.8 401 19.4 32.3 1.7 1.0 46.0 9 0.82 4.3 18.5 19 32 396 21.4 31.9 1.9 1.4 46.1 Example 2 In this test the composition of the leached concentrate was the same as in example 1, likewise the leaching reactor, leaching pressure and temperature.
Leaching time was 11 h. The solution contained 52.5 g/1 of sulphuric acid, 9.8 is g/l of ferrous iron and 1.1 g/I of copper. Oxidation of the solution was stopped after lh and hydrogen sulphide was fed as the reducing gas between 10 -11 h. The zinc content of the concentrate had fallen to a value of 1.0% and the copper content had risen to a value of 20.6%. The reduction precipitated all the dissolved copper, so no copper losses were generated. The progress of leaching is shown in Table 3.
Table 3 shows that, when oxidation ended, the redox potential fell quickly from a level of 406 mV to a level of 390 mV. Before the feed of reductant the value was 330 mV and fell quickly once the reductant feed began to a value of 86 mV. The final value of 207 mV indicates the probable oxygen introduced with the make-up water. The table also indicates that a smaller feed of reduction gas would have been sufficient, because the dissolved copper was precipitated as soon as the feed of reduction gas was begun.
Table 3 Solution Concentrate Time Fe3+ Cu Fe2+ Zn H2S04 Redox Cu Fe Pb Zn S
h g/I /I /I /I g/I mV % % % % %
0 0.4 1.1 9.8 0.0 52.5 510 18.4 28.4 1.5 7.3 40.0 1.0 0.6 4.2 14.1 11.0 48.5 406 19.6 30.3 1.6 2.3 46.3 2.0 0.3 3.4 15.0 12.2 42.0 390 20.4 30.2 1.7 2.1 46.2 3.0 0.3 2.9 15.5 12.9 44.0 386 20.4 30.0 1.7 1.7 44.6 4.0 0.3 2.4 15.6 13.2 47.0 383 5.0 0.3 1.8 16.1 13.9 44.3 384 6.0 0.3 1.5 16.7 14.5 44.8 383 21.3 29.3 1.3 1.2 46.1 8.0 0.4 0.4 16.7 14.5 46.0 370 21.8 29.3 1.4 1.0 46.6 10.0 0.1 0.03 16.4 14.3 43.7 330 22.1 29.5 1.5 1.0 44.7 10.2 0.03 0.001 16.5 14.4 86 10.5 0.03 0.001 16.6 14.3 97 11.0 0.03 0.001 14.9 12.7 40.7 207 20.6 30.1 1.5 1.0 47.4 Example 3 The composition of the concentrate leached in the test was as given in Table 4.
Table 4 Cu Zn Fe Pb S Znaq, W.
% % % % % %
18.5 10.0 27.1 1.4 40.0 0.2 The leaching reactor, leaching pressure and temperature were the same as in examples 1 and 2. Leaching time was 12 h. At the start the solution contained 51.4 g/l of sulphuric acid, 9.1 g/l of ferrous iron and 0.9 g/l of copper. The oxidation of the solution was stopped after 4h and hydrogen sulphide was fed as the reduction gas between 11.5 - 12 h. The zinc content of the concentrate had fallen to a level of 1.4% and the copper content had risen to a level of 21.7%. Reduction precipitated all of the dissolved copper so no copper losses were generated. The progress of leaching is shown in Table 5:
Table 5 Solution Concentrate Time Fe3+ Cu Fe2+ Zn H2SO4 Redox Cu Fe Pb Zn S
h /I g/l /I /I g/l mV % % % % %
0 0.4 0.9 9.5 0.0 51.4 361 18.5 27.1 1.4 10.0 40.0 1 0.3 1.3 10.3 6.7 36.7 409 19.6 29.0 1.4 7.3 43.0 2 0.4 2.0 10.6 10.4 27.8 412 19.8 29.9 1.5 5.3 44.5 3 0.6 2.9 11.1 12.9 21.0 414 19.7 29.7 1.7 3.9 43.7 6 0.6 2.9 12.4 16.1 17.2 394 20.2 30.0 1.3 2.1 49.6 8 18.2 387 9 0.5 1.0 13.4 18.0 382 21.2 29.5 1.3 1.3 45.2 19,4 377 11.5 0.4 0.2 13.7 18.7 366 21.7 29.4 1.4 1.1 46.1 11.7 0.1 0.001 13.0 18.3 177 21.9 30.0 1.3 1.1 46.7 12 0.1 0.0004 13.8 18.3 21.8 279 21.7 29.0 1.5 1.4 46.2
Table 2 Solution Concentrate Time Fe3+ Cu Fe2+ Zn H2SO4 Redox Cu Fe Pb Zn S
h /I /I /I g/I /I mV % % % % %
0 0.53 1.0 9.6 0.0 50.8 412 18.4 28.4 1.5 7.3 40.0 1 0.90 4.9 13.6 9.4 29.3 424 20.0 31.6 2.0 3.3 44.6 3 2.58 8.4 16.5 12.1 14.2 433 18.3 31.3 2.6 1.5 46.8 5.5 1.19 7.8 17.2 13.4 15.8 401 19.4 32.3 1.7 1.0 46.0 9 0.82 4.3 18.5 19 32 396 21.4 31.9 1.9 1.4 46.1 Example 2 In this test the composition of the leached concentrate was the same as in example 1, likewise the leaching reactor, leaching pressure and temperature.
Leaching time was 11 h. The solution contained 52.5 g/1 of sulphuric acid, 9.8 is g/l of ferrous iron and 1.1 g/I of copper. Oxidation of the solution was stopped after lh and hydrogen sulphide was fed as the reducing gas between 10 -11 h. The zinc content of the concentrate had fallen to a value of 1.0% and the copper content had risen to a value of 20.6%. The reduction precipitated all the dissolved copper, so no copper losses were generated. The progress of leaching is shown in Table 3.
Table 3 shows that, when oxidation ended, the redox potential fell quickly from a level of 406 mV to a level of 390 mV. Before the feed of reductant the value was 330 mV and fell quickly once the reductant feed began to a value of 86 mV. The final value of 207 mV indicates the probable oxygen introduced with the make-up water. The table also indicates that a smaller feed of reduction gas would have been sufficient, because the dissolved copper was precipitated as soon as the feed of reduction gas was begun.
Table 3 Solution Concentrate Time Fe3+ Cu Fe2+ Zn H2S04 Redox Cu Fe Pb Zn S
h g/I /I /I /I g/I mV % % % % %
0 0.4 1.1 9.8 0.0 52.5 510 18.4 28.4 1.5 7.3 40.0 1.0 0.6 4.2 14.1 11.0 48.5 406 19.6 30.3 1.6 2.3 46.3 2.0 0.3 3.4 15.0 12.2 42.0 390 20.4 30.2 1.7 2.1 46.2 3.0 0.3 2.9 15.5 12.9 44.0 386 20.4 30.0 1.7 1.7 44.6 4.0 0.3 2.4 15.6 13.2 47.0 383 5.0 0.3 1.8 16.1 13.9 44.3 384 6.0 0.3 1.5 16.7 14.5 44.8 383 21.3 29.3 1.3 1.2 46.1 8.0 0.4 0.4 16.7 14.5 46.0 370 21.8 29.3 1.4 1.0 46.6 10.0 0.1 0.03 16.4 14.3 43.7 330 22.1 29.5 1.5 1.0 44.7 10.2 0.03 0.001 16.5 14.4 86 10.5 0.03 0.001 16.6 14.3 97 11.0 0.03 0.001 14.9 12.7 40.7 207 20.6 30.1 1.5 1.0 47.4 Example 3 The composition of the concentrate leached in the test was as given in Table 4.
Table 4 Cu Zn Fe Pb S Znaq, W.
% % % % % %
18.5 10.0 27.1 1.4 40.0 0.2 The leaching reactor, leaching pressure and temperature were the same as in examples 1 and 2. Leaching time was 12 h. At the start the solution contained 51.4 g/l of sulphuric acid, 9.1 g/l of ferrous iron and 0.9 g/l of copper. The oxidation of the solution was stopped after 4h and hydrogen sulphide was fed as the reduction gas between 11.5 - 12 h. The zinc content of the concentrate had fallen to a level of 1.4% and the copper content had risen to a level of 21.7%. Reduction precipitated all of the dissolved copper so no copper losses were generated. The progress of leaching is shown in Table 5:
Table 5 Solution Concentrate Time Fe3+ Cu Fe2+ Zn H2SO4 Redox Cu Fe Pb Zn S
h /I g/l /I /I g/l mV % % % % %
0 0.4 0.9 9.5 0.0 51.4 361 18.5 27.1 1.4 10.0 40.0 1 0.3 1.3 10.3 6.7 36.7 409 19.6 29.0 1.4 7.3 43.0 2 0.4 2.0 10.6 10.4 27.8 412 19.8 29.9 1.5 5.3 44.5 3 0.6 2.9 11.1 12.9 21.0 414 19.7 29.7 1.7 3.9 43.7 6 0.6 2.9 12.4 16.1 17.2 394 20.2 30.0 1.3 2.1 49.6 8 18.2 387 9 0.5 1.0 13.4 18.0 382 21.2 29.5 1.3 1.3 45.2 19,4 377 11.5 0.4 0.2 13.7 18.7 366 21.7 29.4 1.4 1.1 46.1 11.7 0.1 0.001 13.0 18.3 177 21.9 30.0 1.3 1.1 46.7 12 0.1 0.0004 13.8 18.3 21.8 279 21.7 29.0 1.5 1.4 46.2
Claims (12)
1. A method for upgrading copper sulphide concentrate containing zinc by removing the zinc from the concentrate for the pyrometallurgical recovery of copper, wherein zinc is removed from the concentrate by leaching the concentrate with a solution containing sulphuric acid at atmospheric pressure and temperature firstly in oxidising conditions; before the end of the leach a reducing agent is fed into the leaching stage in order to precipitate the dissolved copper.
2. A method according to claim 1, wherein at the end of the leaching stage the sulphuric acid concentration is adjusted to the range 0-100 g/l.
3. A method according to claim 2, wherein at the end of the leaching stage the sulphuric acid concentration is adjusted to the range 5 - 25 g/l.
4. A method according to any one of claims 1 to 3, wherein an oxidant is fed into the leaching stage.
5. A method according to claim 4, wherein a gas is used as the oxidant, which is at least one of the following: air, oxygen- enriched air and oxygen.
6. A method according to claim 4, wherein at least one of the following is used as oxidant: hydrogen peroxide, manganese dioxide and potassium permanganate.
7. A method according to any one of claims 1 to 6, wherein the redox potential is adjusted so that at the start of the leach the solution contains at least 10 mg/l of copper and at the end of the leach there is no copper in the solution.
8. A method according to any one of claims 1 to 7, wherein the reducing agent fed into the end of the leaching stage is a reducing gas.
9. A method according to claim 8, wherein the gas is hydrogen sulphide.
10. A method according to any one of claims 1 to 7, wherein the reducing agent fed into the end of the leaching stage is sodium hydrosulphide or sodium sulphide.
11. A method according to any one of claims 1 to 7, wherein the reducing agent fed into the end of the leaching stage is copper sulphide concentrate.
12. A method according to any one of claims 1 to 11, wherein at the beginning of the leach the solution contains at least 0.1 g/I of iron and at least 10 mg/I of copper.
Applications Claiming Priority (3)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| FI20070684A FI119819B (en) | 2007-09-07 | 2007-09-07 | Method for improving the quality of copper concentrate |
| FI20070684 | 2007-09-07 | ||
| PCT/FI2008/050483 WO2009030811A1 (en) | 2007-09-07 | 2008-09-03 | Method for upgrading copper concentrate |
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| CA2697187A1 CA2697187A1 (en) | 2009-03-12 |
| CA2697187C true CA2697187C (en) | 2014-04-15 |
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| CN (2) | CN101796204B (en) |
| CA (1) | CA2697187C (en) |
| EA (1) | EA017095B1 (en) |
| FI (1) | FI119819B (en) |
| PE (1) | PE20090848A1 (en) |
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| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US1435699A (en) * | 1919-09-30 | 1922-11-14 | Niels C Christensen | Process of treating sulphide ores of zinc |
| US2609272A (en) * | 1946-12-09 | 1952-09-02 | Guaranty Invest Corp Ltd | Process for the treatment of matte to recover metallic salts |
| DE2840424A1 (en) * | 1978-09-16 | 1980-04-24 | Duisburger Kupferhuette | METHOD FOR PRODUCING SULFIDIC COPPER CONCENTRATES |
| CA1130934A (en) * | 1980-02-08 | 1982-08-31 | Donald R. Weir | Process for the recovery of copper and zinc values from sulphidic ore |
| US5344479A (en) * | 1992-03-13 | 1994-09-06 | Sherritt Gordon Limited | Upgrading copper sulphide residues containing nickel and arsenic |
| AU698137B2 (en) * | 1994-08-15 | 1998-10-22 | R & O Mining Processing Ltd. | Hydrometallurgical conversion of zinc sulfide to sulfate from zinc sulfide co ntaining ores and concentrates |
| AUPN191395A0 (en) * | 1995-03-22 | 1995-04-27 | M.I.M. Holdings Limited | Atmospheric mineral leaching process |
| CN1277939C (en) * | 2004-01-12 | 2006-10-04 | 张在海 | Method of leaching copper in copper containing pyrite using bacteria |
| FI117389B (en) * | 2004-12-28 | 2006-09-29 | Outokumpu Oy | A process for hydrometallurgical treatment of a sulphide concentrate containing several precious metals |
| FI118473B (en) * | 2006-02-17 | 2007-11-30 | Outotec Oyj | Process for extracting copper from copper sulphide ore |
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2007
- 2007-09-07 FI FI20070684A patent/FI119819B/en active IP Right Grant
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- 2008-08-08 PE PE2008001323A patent/PE20090848A1/en active IP Right Grant
- 2008-09-03 CN CN200880105783.2A patent/CN101796204B/en active Active
- 2008-09-03 WO PCT/FI2008/050483 patent/WO2009030811A1/en active Application Filing
- 2008-09-03 CN CN201210296043XA patent/CN102787245A/en active Pending
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| FI20070684A0 (en) | 2007-09-07 |
| CN101796204A (en) | 2010-08-04 |
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| PE20090848A1 (en) | 2009-07-25 |
| CN101796204B (en) | 2014-06-18 |
| FI119819B (en) | 2009-03-31 |
| CA2697187A1 (en) | 2009-03-12 |
| CN102787245A (en) | 2012-11-21 |
| WO2009030811A1 (en) | 2009-03-12 |
| EA017095B1 (en) | 2012-09-28 |
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