CA2099333A1 - Chloride assisted hydrometallurgical copper extraction - Google Patents
Chloride assisted hydrometallurgical copper extractionInfo
- Publication number
- CA2099333A1 CA2099333A1 CA002099333A CA2099333A CA2099333A1 CA 2099333 A1 CA2099333 A1 CA 2099333A1 CA 002099333 A CA002099333 A CA 002099333A CA 2099333 A CA2099333 A CA 2099333A CA 2099333 A1 CA2099333 A1 CA 2099333A1
- Authority
- CA
- Canada
- Prior art keywords
- copper
- sulphate
- solution
- leaching
- ions
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Abandoned
Links
- 239000010949 copper Substances 0.000 title claims abstract description 125
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 119
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 117
- 238000000605 extraction Methods 0.000 title claims abstract description 35
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 title claims abstract description 32
- 239000000243 solution Substances 0.000 claims abstract description 79
- 239000012141 concentrate Substances 0.000 claims abstract description 72
- 238000002386 leaching Methods 0.000 claims abstract description 70
- 229910021653 sulphate ion Inorganic materials 0.000 claims abstract description 64
- 238000000034 method Methods 0.000 claims abstract description 58
- 230000008569 process Effects 0.000 claims abstract description 54
- 238000000638 solvent extraction Methods 0.000 claims abstract description 44
- -1 sulphate ions Chemical class 0.000 claims abstract description 38
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 37
- 235000011149 sulphuric acid Nutrition 0.000 claims abstract description 37
- 238000000909 electrodialysis Methods 0.000 claims abstract description 35
- 239000001117 sulphuric acid Substances 0.000 claims abstract description 35
- ARUVKPQLZAKDPS-UHFFFAOYSA-L copper(II) sulfate Chemical compound [Cu+2].[O-][S+2]([O-])([O-])[O-] ARUVKPQLZAKDPS-UHFFFAOYSA-L 0.000 claims abstract description 30
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 claims abstract description 26
- 150000001879 copper Chemical class 0.000 claims abstract description 25
- 230000002378 acidificating effect Effects 0.000 claims abstract description 15
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 claims abstract description 14
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims abstract description 12
- 229910052760 oxygen Inorganic materials 0.000 claims abstract description 12
- 239000001301 oxygen Substances 0.000 claims abstract description 12
- 239000003929 acidic solution Substances 0.000 claims abstract description 6
- 239000002184 metal Substances 0.000 claims abstract description 4
- 229910052751 metal Inorganic materials 0.000 claims abstract description 4
- 238000005363 electrowinning Methods 0.000 claims description 16
- 238000004064 recycling Methods 0.000 claims description 11
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 claims description 9
- 229910001431 copper ion Inorganic materials 0.000 claims description 9
- 230000000694 effects Effects 0.000 claims description 8
- ORTQZVOHEJQUHG-UHFFFAOYSA-L copper(II) chloride Chemical compound Cl[Cu]Cl ORTQZVOHEJQUHG-UHFFFAOYSA-L 0.000 claims description 6
- 239000000203 mixture Substances 0.000 claims description 6
- 239000002253 acid Substances 0.000 description 62
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 40
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 38
- 239000005864 Sulphur Substances 0.000 description 38
- 230000003647 oxidation Effects 0.000 description 23
- 238000007254 oxidation reaction Methods 0.000 description 23
- 229910052742 iron Inorganic materials 0.000 description 18
- 238000006243 chemical reaction Methods 0.000 description 17
- 239000000706 filtrate Substances 0.000 description 15
- 238000012360 testing method Methods 0.000 description 14
- 239000007787 solid Substances 0.000 description 10
- DGAQECJNVWCQMB-PUAWFVPOSA-M Ilexoside XXIX Chemical compound C[C@@H]1CC[C@@]2(CC[C@@]3(C(=CC[C@H]4[C@]3(CC[C@@H]5[C@@]4(CC[C@@H](C5(C)C)OS(=O)(=O)[O-])C)C)[C@@H]2[C@]1(C)O)C)C(=O)O[C@H]6[C@@H]([C@H]([C@@H]([C@H](O6)CO)O)O)O.[Na+] DGAQECJNVWCQMB-PUAWFVPOSA-M 0.000 description 8
- 229910052708 sodium Inorganic materials 0.000 description 8
- 239000011734 sodium Substances 0.000 description 8
- 238000010977 unit operation Methods 0.000 description 8
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 8
- 229910052951 chalcopyrite Inorganic materials 0.000 description 7
- DVRDHUBQLOKMHZ-UHFFFAOYSA-N chalcopyrite Chemical compound [S-2].[S-2].[Fe+2].[Cu+2] DVRDHUBQLOKMHZ-UHFFFAOYSA-N 0.000 description 6
- 229910052595 hematite Inorganic materials 0.000 description 6
- 239000011019 hematite Substances 0.000 description 6
- LIKBJVNGSGBSGK-UHFFFAOYSA-N iron(3+);oxygen(2-) Chemical compound [O-2].[O-2].[O-2].[Fe+3].[Fe+3] LIKBJVNGSGBSGK-UHFFFAOYSA-N 0.000 description 6
- 239000002002 slurry Substances 0.000 description 6
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 description 5
- 238000001914 filtration Methods 0.000 description 5
- 239000002904 solvent Substances 0.000 description 5
- 230000008901 benefit Effects 0.000 description 4
- BWFPGXWASODCHM-UHFFFAOYSA-N copper monosulfide Chemical compound [Cu]=S BWFPGXWASODCHM-UHFFFAOYSA-N 0.000 description 4
- 239000003792 electrolyte Substances 0.000 description 4
- 239000012535 impurity Substances 0.000 description 4
- 239000012528 membrane Substances 0.000 description 4
- 238000011084 recovery Methods 0.000 description 4
- 239000012065 filter cake Substances 0.000 description 3
- 230000006872 improvement Effects 0.000 description 3
- 150000002500 ions Chemical class 0.000 description 3
- 238000006386 neutralization reaction Methods 0.000 description 3
- UWGTVLYQSJNUFP-CAPFRKAQSA-N 4-dodecyl-2-[(E)-hydroxyiminomethyl]phenol Chemical compound [H]\C(=N/O)C1=CC(CCCCCCCCCCCC)=CC=C1O UWGTVLYQSJNUFP-CAPFRKAQSA-N 0.000 description 2
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N Iron oxide Chemical compound [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 description 2
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 2
- 230000009286 beneficial effect Effects 0.000 description 2
- 229910052948 bornite Inorganic materials 0.000 description 2
- 229910000366 copper(II) sulfate Inorganic materials 0.000 description 2
- 229910052971 enargite Inorganic materials 0.000 description 2
- 238000001704 evaporation Methods 0.000 description 2
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 2
- 229910052737 gold Inorganic materials 0.000 description 2
- 239000010931 gold Substances 0.000 description 2
- JEIPFZHSYJVQDO-UHFFFAOYSA-N iron(III) oxide Inorganic materials O=[Fe]O[Fe]=O JEIPFZHSYJVQDO-UHFFFAOYSA-N 0.000 description 2
- 239000003350 kerosene Substances 0.000 description 2
- 239000012633 leachable Substances 0.000 description 2
- 230000014759 maintenance of location Effects 0.000 description 2
- 229910052709 silver Inorganic materials 0.000 description 2
- 239000004332 silver Substances 0.000 description 2
- 239000002699 waste material Substances 0.000 description 2
- 229910052725 zinc Inorganic materials 0.000 description 2
- 239000011701 zinc Substances 0.000 description 2
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 1
- 235000011941 Tilia x europaea Nutrition 0.000 description 1
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 1
- 238000004458 analytical method Methods 0.000 description 1
- 230000003190 augmentative effect Effects 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 229910052799 carbon Inorganic materials 0.000 description 1
- 229910052947 chalcocite Inorganic materials 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 239000000470 constituent Substances 0.000 description 1
- 238000001816 cooling Methods 0.000 description 1
- BUGICWZUDIWQRQ-UHFFFAOYSA-N copper iron sulfane Chemical compound S.[Fe].[Cu] BUGICWZUDIWQRQ-UHFFFAOYSA-N 0.000 description 1
- JJLJMEJHUUYSSY-UHFFFAOYSA-L copper(II) hydroxide Inorganic materials [OH-].[OH-].[Cu+2] JJLJMEJHUUYSSY-UHFFFAOYSA-L 0.000 description 1
- AEJIMXVJZFYIHN-UHFFFAOYSA-N copper;dihydrate Chemical compound O.O.[Cu] AEJIMXVJZFYIHN-UHFFFAOYSA-N 0.000 description 1
- 230000003247 decreasing effect Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000006073 displacement reaction Methods 0.000 description 1
- 238000004090 dissolution Methods 0.000 description 1
- 239000008151 electrolyte solution Substances 0.000 description 1
- 230000008020 evaporation Effects 0.000 description 1
- 238000002474 experimental method Methods 0.000 description 1
- 239000012527 feed solution Substances 0.000 description 1
- 238000005188 flotation Methods 0.000 description 1
- 239000003292 glue Substances 0.000 description 1
- 238000011065 in-situ storage Methods 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- 229910052935 jarosite Inorganic materials 0.000 description 1
- 239000004571 lime Substances 0.000 description 1
- 239000007788 liquid Substances 0.000 description 1
- 238000011068 loading method Methods 0.000 description 1
- 235000010755 mineral Nutrition 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 238000005065 mining Methods 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 239000013110 organic ligand Substances 0.000 description 1
- 230000001590 oxidative effect Effects 0.000 description 1
- 239000002245 particle Substances 0.000 description 1
- 238000001556 precipitation Methods 0.000 description 1
- 238000002360 preparation method Methods 0.000 description 1
- 230000009467 reduction Effects 0.000 description 1
- 238000009877 rendering Methods 0.000 description 1
- 230000000630 rising effect Effects 0.000 description 1
- 238000012216 screening Methods 0.000 description 1
- 239000011780 sodium chloride Substances 0.000 description 1
- 239000011343 solid material Substances 0.000 description 1
- 229910052969 tetrahedrite Inorganic materials 0.000 description 1
- 238000012546 transfer Methods 0.000 description 1
- 238000013022 venting Methods 0.000 description 1
- 238000005406 washing Methods 0.000 description 1
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 description 1
- 239000011686 zinc sulphate Substances 0.000 description 1
- 235000009529 zinc sulphate Nutrition 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
- C22B15/0063—Hydrometallurgy
- C22B15/0065—Leaching or slurrying
- C22B15/0067—Leaching or slurrying with acids or salts thereof
- C22B15/0071—Leaching or slurrying with acids or salts thereof containing sulfur
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B01—PHYSICAL OR CHEMICAL PROCESSES OR APPARATUS IN GENERAL
- B01D—SEPARATION
- B01D61/00—Processes of separation using semi-permeable membranes, e.g. dialysis, osmosis or ultrafiltration; Apparatus, accessories or auxiliary operations specially adapted therefor
- B01D61/42—Electrodialysis; Electro-osmosis ; Electro-ultrafiltration; Membrane capacitive deionization
- B01D61/44—Ion-selective electrodialysis
-
- 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
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
-
- 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
-
- 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
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
-
- 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
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
-
- 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)
- Metallurgy (AREA)
- Mechanical Engineering (AREA)
- Materials Engineering (AREA)
- Manufacturing & Machinery (AREA)
- Geology (AREA)
- Life Sciences & Earth Sciences (AREA)
- Geochemistry & Mineralogy (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- Water Supply & Treatment (AREA)
- Health & Medical Sciences (AREA)
- Urology & Nephrology (AREA)
- Chemical Kinetics & Catalysis (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
CHLORIDE ASSISTED HYDROMETALLURGICAL COPPER EXTRACTION
ABSTRACT OF THE DISCLOSURE
A process for the extraction of copper from a sulphide copper ore or concentrate comprises subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant which comprises an acidic solution of chloride and sulphate ions to produce an insoluble basic copper sulphate. The basic copper sulphate so produced is leached in a second leaching preferably at atmospheric pressure with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution. The resulting leach liquor is subjected to solvent extraction to produce a copper concentrate solution and a raffinate comprising protons and sulphate ions in solution. Protons and sulphate ions are extracted from the raffinate to produce a sulphuric acid solution which sulphuric acid solution is then recycled to the first leaching at elevated pressure and temperature to serve as a source of sulphate ions in the lixiviant. According to a preferred embodiment, the extraction of the protons and sulphate ions from the raffinate is effected by means of an electrodialysis process. In alternative embodiments copper sulphate is also extracted from the raffinate and recycled to the first leaching step or alternatively a source of sulphate ions, such as a hydrolyzable metal sulphate is introduced from another source.
ABSTRACT OF THE DISCLOSURE
A process for the extraction of copper from a sulphide copper ore or concentrate comprises subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant which comprises an acidic solution of chloride and sulphate ions to produce an insoluble basic copper sulphate. The basic copper sulphate so produced is leached in a second leaching preferably at atmospheric pressure with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution. The resulting leach liquor is subjected to solvent extraction to produce a copper concentrate solution and a raffinate comprising protons and sulphate ions in solution. Protons and sulphate ions are extracted from the raffinate to produce a sulphuric acid solution which sulphuric acid solution is then recycled to the first leaching at elevated pressure and temperature to serve as a source of sulphate ions in the lixiviant. According to a preferred embodiment, the extraction of the protons and sulphate ions from the raffinate is effected by means of an electrodialysis process. In alternative embodiments copper sulphate is also extracted from the raffinate and recycled to the first leaching step or alternatively a source of sulphate ions, such as a hydrolyzable metal sulphate is introduced from another source.
Description
9 3 ~, 3 3 CHLORIDE ASSISTED HYDROMETALL~RGICAL COPPER EXTRACTION
FIELD OF THE INVENTION
This invention relates to a hydrometallurgical ~ treatment of copper sulphide ores or concentrates in the 3 presence of chloride ions.
BACKGROUND OF THE INVENTION
Effective hydrometallurgical treatment of copper sulphide ores, such as chalcopyrite (CuFeS2) has been a long standing goal in the copper mining industry which has thus far eluded success. The problem lies in the fact that the severe conditions required for the effective leaching of copper from these ores results in I oxidation of the sulphide in the ore or concentrate to I sulphate, resulting in the generation of acid which requires expensive neutralization, rendering the process impractical and uneconomical. 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 I to sulphate. These attempts include the pretreatment of ¦ 25 the concentrate prior to the pressure leaching 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. Patent 4,039,406.
In this process, the copper values in the concentrate are transformed into a solid basic copper sulphate from which th~ copper values must then be subsequently recoveréd, as described in U.S. Patent 4,338,168. In the process described in patent 4,039,406 a significant amount (20-25%) of sulphide in the ore or concentrate is still oxidized to sulphate, resulting in greater oxygen demand t;' ~
FIELD OF THE INVENTION
This invention relates to a hydrometallurgical ~ treatment of copper sulphide ores or concentrates in the 3 presence of chloride ions.
BACKGROUND OF THE INVENTION
Effective hydrometallurgical treatment of copper sulphide ores, such as chalcopyrite (CuFeS2) has been a long standing goal in the copper mining industry which has thus far eluded success. The problem lies in the fact that the severe conditions required for the effective leaching of copper from these ores results in I oxidation of the sulphide in the ore or concentrate to I sulphate, resulting in the generation of acid which requires expensive neutralization, rendering the process impractical and uneconomical. 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 I to sulphate. These attempts include the pretreatment of ¦ 25 the concentrate prior to the pressure leaching 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. Patent 4,039,406.
In this process, the copper values in the concentrate are transformed into a solid basic copper sulphate from which th~ copper values must then be subsequently recoveréd, as described in U.S. Patent 4,338,168. In the process described in patent 4,039,406 a significant amount (20-25%) of sulphide in the ore or concentrate is still oxidized to sulphate, resulting in greater oxygen demand t;' ~
- 2 - 2 ~ ~ 9 3 3 ~
during the pressure leach and the generation of sulphuric acid.
It is accordingly an object of the present invention to provide a hydrometallurgical copper extraction process wherein the oxidation of sulphide in the ore or concentrate to sulphate is reduced.
SUMMARY OF THE INVENTION
According to the invention, there is provided a process for the extraction of copper from a sulphide, copper ore or concentrate, comprising the steps of subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt; leaching the basic copper salt produced by said Pirst leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution; subjecting said leach liquor to a solvent extraction process to produce a copper concentrate solution and a raffinate comprising protons and sulphate ions in solution; extracting protons and sulphate ions from said raffinate to produce a sulphuric acid solution;
and recycling said sulphuric acid solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant.
According to a preferred embodiment, the raffinate is subjected to electrodialysis to effect said extraction of protons and sulphate ions therefrom.
Also according to the invention, there is provided a process for the extraction of copper from a .
~" ,,, ,, ", , ' " " ~ ", ~ " . ' , , ' ~ ", ', ~
,: ~ . ~- ~ , , , .." , , :, , - : : . ,: ;---` 2{~9~333~ - 3 -, , sulphide copper ore or concentrate, comprising the steps of subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt; leaching the basic copper salt produced by said first leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in : 10 solution; subjecting said leach liquor to a solvent extraction process to produce a first copper concentrate solution and a raffinate comprising protons, copper ions and sulphate ions in solution; extracting copper ions and sulphate ions from said raffinate to produce a second copper concentrate solution, recycling said second copper concentrate solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant; and subjecting said first ::
copper concentrate solution to electrowinning to recover copper values therefrom.
' .
Further according to the invention, there is provided a process for the extraction of copper from a ! sulphide copper ore or concentrate, comprising the steps of subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a solution of chloride ions and at an acidic pH
to produce an insoluble basic copper salt; adding to said solution, during said first leaching, a source of sulphate ions selected from the group consisting of sulphuric acid, copper sulphate and a metal sulphate which hydrolyzes at said acidic pH and providing sufficient of said source of sulphate ions to react said source of sulphate ions with said ore or concentrate to form said basic copper salt;
and leaching said basic copper salt in a second leaching with an acidic sulphate solution to dissolve the basic ;: :", 2~99~33 copper salt to produce a leach liquor containing copper I sulphate in solution.
In the present specification, elevated temperature means a temperature above room temperature (25C), preferably from about 125C to about 175C, and elevated pressure means an oxygen partial pressure above atmospheric pressure, preferably from about 50 psig to about 250 psig (1380 kPa).
Further objects and advantages of the invention will become apparent from the description of a preferred embodiment of the invention below.
BRIEF DESCRIPTION OF THE DRAWING
, .
The single drawing is a flow diagram of a hydrometallurgical copper extraction process according to the invention.
DETAILED DESCRIPTION OF PREFERRED EMBODIMENT
The stages of the process comprise a pressure leaching stage 38 in an autoclave, an atmospheric leaching stage 40, a solvent extraction and stripping stages 42 and 56, an electrodialysis stage 44 and an electrowinning stage 46. -After each of the leaching stages 38 and 40, filtration is carried out as indicated at 48 and 50, respectively, to separate the liquids and solids.
The process will now be described in greater detail by way of a specific example in which a chalcopyrite concentrate is treated.
.~ . . . : . ~ : : ~ , , ~ 9~3~
The copper concentrate is first ground in a ball mill, during a feed preparation step 52, to reduce the size of the particles to about 95% minus 325 'ryler mesh or smaller. Although satisfactory results are obtainable without regrinding, it has been found that there is a small but significant improvement with regrinding~
The concentrate is leached in the pressure leaching stage 38 at an elevated pressure and temperature with a lixiviant containing from about 3-15 grams per litre copper, 6-18 grams per litre chloride and about 15-35 grams per litre sulphuric acid. About 80% of the lixiviant is leach liquor which is recycled after the atmospheric leaching stage 40 and filtration 50. The remaining 20~ comprises concentrated sulphuric acid which is recycled from the electrodialysis stage 44, as will be described in more detail below.
The temperature of the leach 38 is about 150C
and the pressure about 200 psig (1380 kPa). This is total pressure comprising oxygen pressure on top of the 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.
The solids content is maintained at about 15-20%, i.e. 170-250 grams per litre solids as determined by the heat balance. A higher percentage solids would require some form of heat removal to prevent the temperature from rising above the desired limit of about 150C.
As referred to above, the lixiviant used in the pressure leach 38 is made up partly of recycled lixiviant from a previous pressure leach (Stream 8) but augmented by an acid concentrate (Stream 60) which is recycled from the ~.. . . .
"
. . -. . ~
. .. ...
~g33~
.
; electrodialysis stage 44. The immediate effect of adding the acid concentrate to the lixiviant is to increase the acidity of the lixiviant which is fed to the autoclave for the pressure leaching stage 38, but the most important effect, surprisingly, has been found to be that the addition of the acid, or ~ore specifically the sulphate ions, actually suppresses the oxidation of sulphur emanating from the concentrate during the pressure , leaching stage 38.
Typically the oxidation of sulphur that is experienced if no acid recycle is used is about 25% of the feed sulphur in the concentrate, as is the case with the process described in U.S. Patent ~,039,406. However, if acid recycle is used, it has been found that the sulphur oxidation to sulphate is reduced to about 5-10%. This improvement has substantial beneficial effects OJI the hydrometallurgical extraction process. The oxidation of sulphur to sulphate creates additional costs in several ways, such as additional oxygen required for the reaction, additional reagent required to neutralize the acid so I 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 38 takes place.
The chemistry of the reaction in the pressure leaching stage 38 is believed to be altered by the addition of the acid as follows:
No acid addition-CuFeS2 + 7/402 + 2/3H20 ~ [1/3CuSO4-2/3Cu(OH) 2 ] + 1/2Fe203 + 5/3S
~3~
_ 7 _ ~ 9 ~ 3 3 3 With acid addition:
CuFeS2 + 5/402 + 1/3H20 + l/3H2S04 ~ [l/3CuSO4-2/3Cu(OH)2]
+ 1/2Fe203 + 2S0 In both reactions, the copper is precipitated in the form of a basic copper salt, which has been found to comprise about 90% of basic copper sulphate, which contains a sulphate anion, as indicated in the reaction equations, but about 10~ of basic copper chloride is also formed. 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 recycle, thus obviating the need for the oxidation of sulphur to sulphate. Thus in the second reaction, there is a nett consumption of ~ulphate ions to form the basic copper salt.
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% is found. With acid addition, experiments indicated about 5-10~ 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 38 but are only an approximation.
In order to take advantage of the beneficial effect of the sulphuric acid recycle to inhibit the oxidation of sulphur, it is necessary to find efficient ways of adding acid into the autoclave, which has limited ability to absorb acid because the bulk of the leach ., ,. :,.
` ~9333~
liquor is recycled. However, there is some loss of leach liquor in the pressure leaching stage 38, due to venting (steam losses) and due to leach liquor carried off in the filter cake after the pressure leach stage 38 (Stream 7).
It has been found that about 20% of the volume of leach liquor is lost during each cycle in this fashion. The amount of sulphuric acid needed to suppress sulphur oxidation has been found experimentally to be about 25 grams per litre. Therefore, since this amount of acid must be contained in 20% of the volume, it must come in a concentrated form, i.e.
25 grams per litre = 150 grams per litre There is surplus acid produced in the solvent extraction stage 42 where the CUSO4 solution is changed into an H2S04 solution. However, the acid so produced is very dilute (Stream 17), only about 20-40 grams per litre, due to the nature of the solvent extraction chemistry.
The difficulty cannot effectively be overcome by simply evaporating the raffinate (Stream 17) coming from the solvent extraction stages 42 and 56 to produce a concentrated acidic solution because it is necessary to eliminate impurities in the raffinate, such as iron and zinc/ and evaporation followed by recycling would return the impurities to the pressure leaching stage 38.
This problem has been solved in the present invention by extracting sulphuric acid from the raffinate and in the particular embodiment described here, use is made of electrodialysis to effect this extraction. Thus/
by introducing the electrodialysis stage 44/ which will be described in more detail below, sulphuric acid is recovered from the raffinate in a concentrated form suitable for use in the pressure leaching stage 38. -~
~.,, ~ - , . . .
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. . .
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, g ., The slurry produced by the pressure leach 38 is cooled to below 100C and then filtered 48 to separate the residue from the leach liguor or lixiviant, which is recycled through a cooling tower 54 to the leaching stage 38 as noted above.
i , The residue contains the copper originally present in the concentrate as insoluble basic copper sulphate and basic copper chloride together with all the other solid materials, such as Fe2O3 ~hematite) and elemental sulphur.
There is a gain in the weight of the leach residue. Typically it has 30-40% more weight than the Peed concentrate. It has been found that the leach residue contains about 0.5-2% chloride, as well as the copper, iron oxide and sulphur, which is due to the presence of the basic copper chloride and the basic ~ 20 copper sulphate. The iron in the chalcopyrite concentrate 1 i8 converted almost completely to hematite, while sulphur is mostly converted to the elemental form with only a fraction (about 5-10%) being oxidized to sulphate, as noted above.
The leach liquor produced by the pressure leaching step 38 has much the same composition as the feed ~ lixiviant except that there is a drop in the chloride I concentration from about 12 grams per litre to about 7-10 grams per litre, depending on the conditions, due to the formation of the basic copper chloride.
The filter cake or leach residue is repulped in raffinate from the subsequent solvent extraction stage 42, which comprises an acidic sulphate solution containing ~ . .. .
~ . , ,,;;, 2 1~ 3 about 20-40 grams per litre H2S04 and a small amount of -copper, about 1-3 grams per litre.
The second leaching sta~e 40 takes place at atmospheric pressure and a temperature of about 40C for a retention time of about 15-120 minutes. The percentage solids is about 6-11% or about 70-140 grams per litre.
The final acidity of the slurry is about pH 1.5-2.0 or about 2-5 grams per litre H2SO4.
During the atmospheric leaching stage 40, the basic copper salts dissolve almost completely with very little of the iron going into solution.
Typically, the leach liquor produced after filtration 50 contains about 10-20 grams per litre copper with less than 1 grams per litre iron and about 0.3-1.0 grams per litre chloride.
The percentage solids is kept low during the atmospheric leaching stage 40 because higher copper concentrations cannot be treated satisfactorily by the subsequent solvent extraction circuit.
The copper extraction has been found to be about 97-98% based on the original feed to the pressure leaching stage 38. Iron extraction to solution has been found to be less than about 5%.
The main constituents of the solid residue after filtration 50 are hematite and elemental sulphur, as well as any gold or silver which may have been present in the original concentrate. The sulphur can be recovered by screening or flotation to separate it from the hematite into a high-grade sulphur concentrate, which can be further treated for recovery of sulphur (Stream 64). The - ~ .
~ ~ ' ' - : ' ' , , ~ r~ 20~3~3 .:
, gold and silver can be recovered by cyanidation after sulphllr is removed from the leach residue (Stream 12).
The copper leached in the atmospheric leaching stage 40 is extracted by means of solvent extraction 42 to produce a loaded copper electrolyte suitable for electrowinning 46. After the solvent extraction stage 42, the loaded organic extractant is subjected to washing and stripping 56. The high copper concentration of about 10-20 grams per litre derived from the atmospheric leaching stage 40 provides significant advantages over conventional solvent extraction/electrowinning plants because much higher loading of the organic is possible, thus reducing the size of the plant for a given tonnage of copper. Stripping of the loaded organic is effected by ! means of spent acid from the electrowinning stage 46 to obtain a pure copper sulphate solution which is then passed to the electrowinning stage 46.
Spent acid from the electrowinning stage 46 is recycled to the solvent extraction stage 42 to strip the copper from the loaded extractant.
The raffinate from the solvent extraction s'cage 42 is divided into two portions. A first portion (Stream 9) comprising two-thirds of the raffinate is recycled to the atmospheric leach stage 40. A second portion -i comprising one-third of the raffinate (Stream 18) is sent to the electrodialysis stage 44 to produce a diluate and a concentrate acid solution. The diluate solution is 0.5-3 1.0 grams per litre copper and 7-12 grams per litre sulphuric acid, and the concentrate acid solution is 5-10 grams per litre copper and 150-170 grams per litre H2S04.
Typically the concentrate stream (Stream 19) from the electrodialysis stage 44 will be about 10-20~ of Ji: ,. .. ,. . .
- 12 - ~ 9333 the feed flow, whereas the diluate stream (Stream 21) comprises the rest or 80-90% and contains the bulk of the water in the feed stream, as well as any ions that have been rejected by the membrane, such as Cu2' present in the ¦ 5 feed. It is desirable to recover such CUSO4 and this is effected in the process according to the present invention by subjecting the diluate stream to an auxiliary solvent extraction circuit 58. The circuit 58 comprises an extracting stage 64 and a stripping stage 66, for the extraction and stripping operations, respectively. The concentrate acid solution from the electrodialysis stage 44 is used as stripping acid in the stripping stage 66 to strip the copper from the loaded organic. The acid solution resulting from the stripping stage 66 is 140-160 grams per litre sulphuric acid which is recycled to the pressure leach stage 38. Typically the feed to this circuit (the diluate stream) will contain about 0.5-2.0 ; grams per litre copper and the raffinate or waste stream will contain about 0.05-0.1 grams per litre copper. This corresponds to about 0.3% of the original feed copper, considering that, for example, the feed liquor to the main solvent extraction stage 42 (Stream 11) contains about 10 grams per litre copper and is three times the flow of the waste stream from the auxiliary extracting stage 64. The overall extraction of copper in the process has been found to be as high as 99.7%. -As noted above, the residue from the pressure leaching stage 38 (Stream 7) comprises a mixture of elemental sulphur (S), hematite (Fe203) and mainly basic copper sulphate. This residue is fed to the atmospheric leaching stage 40 where the basic copper sulphate is dissolved in acid as far as possible, leaving the hematite and elemental sulphur components essentially untouched.
This produces a solution of copper sulphate (CuS04) (Stream 11~ which is fed to the solvent extraction stage .
2~9~333 42, where copper is exchanged with an organic ligand (R-H), producing acid in the aqueous stream, the raffinate (Stream 17). The reactions can be summarized as follows:
Atmospheric Leachina Stage:
CuSO4 2Cu(OH) 2 + 2HzSO4 ~ 3CuSO4 + 4H2O
Solvent Extraction Stage:
3CuSO4 + 6R-H , 3~2Cu + 3H2SO4 Thus the overall reaction can be represented as follows:
CuSO4-2Cu(OH) 2 + 6R-H ~ 3R2Cu + H2S04 + 4Hz0 There is thus one extra mole of acid produced for every three moles of Cu leached in the atmospheric leach. In order to make use of this extra mole of H2SO4, the raffinate stream from the solvent extraction stage 42 is split, as noted above, so that two-thirds thereof (Stream 9) are returned to the atmospheric leaching stage 40 and the remaining one-third (Stream 18) is fed to the electrodialysis stage 44 to produce the acid concentrate which is fed to the pressure leaching stage 38.
An additional benefit of the process according to the invention is that chloride ions lost from the pressure leach circuit into the pressure leach residue, either as insoluble basic copper chloride or as entrained solution losses in the filter cake, can be recycled along with the acid concentrate back to the pressure leach. Any chloride ions present in the pressure leach residue will report almost quantitatively to the atmospheric leach liquor and thence to the raffinate after solvent extraction. If not bled ~rom this circuit they would quickly build up to higher levels in the atmospheric leach liquor and eventually transfer to the electrowinning circuit where chloride is particularly undesirable. By - - 14 - ~a~9333 splitting the raffinate from the solvent extraction stage -,~ 42, as noted above, and treating one-third thereof through j the electrodialysis stage 44 and solvent extraction 58, this effectively recycles the chloride content back to the pressure leach thereby minimizing any chloride makeup requirements therein.
.~ .
With the process according to the present invention, relatively high copper recoveries, typically 97 to 98% at quite low pressures, such as 40C, have been obtained. Such low temperatures are known to suppress iron dissolution and test results have shown only about 200 ppm Fe with 10 grams per litre Cu in the atmospheric leach solution after filtration 50 (Stream 11). This is a marked improvement over prior art processes, such as described in U.S. Patent 4,338,168, which reports only about a 93% recovery at this temperature and requires higher temperature and/or acid levels to obtain ;~ satisfactory copper recovery values. Unfortunately, such more severe conditions also dissolves about 50% of the Fe in the feed to the atmospheric leaching stage 40 complicating the process by requiring the addition of a ¦ jarosite precipitation process to separate the Fe from the copper. In the present process the Fe is rejected in the atmospheric leach residue.
Due to the reduction in sulphur oxidation and the effective recycling of the sulphuric acid as de~cribed, the process according to the invention does not ; 30 require any special neutralization procedure. Since only about 5-10% of the sulphur is oxidized to sulphate only a '~ relatively small amount of acid is produced which can effectively be taken care of by a lime neutralization process (Stream 35) which is required in any event for the treatment 68 of the final effluent from the solvent !
r"~
~::, . : ~ , , .
- 15 -2~9~333 extraction cycle 58, which is the bleed of impurities such as Zn and Mg from the circuit.
The results of tests which were carried out for the various stages of the process will now be given in the following Examples. In Example 1 the feed to the pressure j leach did not contain acid. In Example 2, an acidic feed was charged to the leach.
10Example 1 The copper concentrate from a porphyry deposit in Highland Valley, British Columbia (Stream 1 on the flowsheet) is composed of 40.19% copper, 20.50% iron, and 29.24% sulphur. In both Examples 1 and 2, the concentrate was ground to 98~-400 mesh. ~n Example 1, the charge to the autoclave had a wet weight of 175.1 grams at 14.4%
moisture. The solution feed to the leach was a combination of 900 ml of recycled pressure leach filtrate (Stream 8) containing 1.5 grams per litre copper, less than 1 ppm iron and 11.47 grams per litre chloride and 100 ml of water. The makeup water (Stream 31 on the flowsheet) actually contained 2.8 grams of sodium chloride so that the total chloride concentration in the leach was 12.0 grams per litre. The concentrate was leached for one hour at 200 psi and 150C. Upon completion of the pressure leach, the slurry (Stream 51 on the flowsheet) was filtered. The 995 ml of filtrate, Stream 6, contained 1.0 grams per litre copper, less than 1 ppm iron, 8.3 grams per litre sodium, and 11.6 grams per litre chloride and had a pH of 3.9. The total wet weight of the residue from the pressure leach was 323.0 grams. A 91 gram sample was taken for analysis. This sample contained 32.4%
copper, 16.9% iron, 0.49% sodium and 10.4% elemental sulphur and had a moisture content of 37.7% moisture.
, - ,, . -... .
~: . ~ ', ' ' :
20~333 The residue from the pressure leach was subjected to an atmospheric leach for an hour at 40C and ; a pH 1.7. The charge to this leach consisted of 231.6 grams of solids at 37.7% moisture and 2120 ml of water with 30 ml of concentrated H2S04 (Stream 9 on the flowsheet). The slurry from this leach, Stream 10 was filtered to obtain lOg.4 grams of residue at 36.2%
moisture and a 2120 ml filtrate. The residue was washed , once by displacement and resulted in a 245 ml wash water containing 4.2 grams per litre copper and 217 ppm iron.
~ The filtrate, Stream 11, consisted of 23.2 grams per litre copper, 403 ppm iron, 420 ppm sodium, and 1.2 grams per litre chloride. The pH of the filtrate was 1.7 and had a free acid of 3.1 grams per litre. The residue from the leach contained 2.52% copper, 0.16% sodium and 28.9% iron.
The results of this example are given below.
Example 2 To reduce sulphur oxidation, sulphuric acid was added to the feed of the pressure leach. The sulphuric acid provided the sulphur needed to form Basic Copper ! Sulphate instead of oxidizing sulphur in the concentrate.
' The following equation defines th~ reaction which is 1 25 occurring:
CuFeS2 + 5/402 + 1/3H20 + 1/3H2S04 ~ [1/3CuS04-2/3Cu(OH)2]
+ 1/2Fe203 + 2S
This addition of acid to the feed of the leach reduces the sulphur oxidation from 28% to 9%. The charge to the present leach consisted of 183.0 grams of wet concentrate at 16% moisture. Since the actual liquor from a past pressure leach, Stream 8, and the concentrated acid from the electrodialysis/solvent extraction, Stream 60, were unavailable, these feeds were made synthetically.
, - . :
~, . .. : . ~ : , ., -: , . ..
- 17 ~ 2~333 The pressure leach was charged with 1000 ml of feed solution having a chloride concentration of 12.0 grams per litre and a free acid concentration of 27.0 grams per litre. The concentrate was leached for an hour at 200 psi and 150C. The slurry from this leach, Stream 51, was filtered. The 1025 ml filtrate, Stream 6, contained 6.0 grams per litre copper, 20 ppm iron, 6.0 grams per litre ~ sodium, and 10.4 grams per litre chloride. The pH of the j filtrate was 3.1. The residue weighed 331.5 grams wet ; 10 from which a 48 gram sample was taken. The residue had a moisture content of 32.6% and contained 25.4% copper, 13.6% iron, 0.23% sodium and 1~.41% elemental sulphur.
As in Example 1, the residue from the pressure leach was subjected to an atmospheric leach.
Approximately 284 grams of wet residue from the pressure ¦ leach was combined with 3500 ml of acidic water (Stream 9 on the flowsheet). The residue was leached for an hour at 40C at a pH of 1.5. The slurry from this leach was once again filtered to obtain 3205 ml of filtrate, Stream ~1, and 132.~ grams of wet residue. The residue was washed with 385 ml o~ water and produced a wash water with 3.38 grams per litre copper and 143 ppm iron. The filtrate contained 11.9 grams per litre copper, 580 ppm iron, 0.10 gram~ per litre chloride and 0.16 grams per litre sodium.
The residue had a moisture content of 28.1% and consisted of 1.65% copper, 0.03% sodium and 16.23% iron.
Comparison of ~xamples 1 and 2 To compare the effect of adding acid to the feed of the pressure leach on the leaching of copper, the following tables illustrate the copper extraction and the sulphur oxidation for both tests.
,,~,.. .
~,.... . .
~: .
3 3 3 : -Table 1: Copper extracti~n for tests Examples 1 and 2 r~ T~ --T---~ ---- ~~~~~~ ---~ -T---- ~
1 I Fesd Copper ¦ Residue Copper ¦ X
¦ Example ~ ------T~ ~ ~ ~ ~ Extraction ¦
¦ Dry ~t.(g~ ¦ X Cu ¦ g Cu ¦ Dry ~t. ~g) ¦ X Cu ¦ g Cu ¦
10 ¦ 1 ¦ 150.0 ¦ 40.19 ¦ 60.2 ¦ 96.2 ¦ 2.52 ¦ 2.42 ¦ 96.0 I~ - I ----t--- ~
2 ¦ 153.7 ¦ 40-19 ¦ 61.8 ¦ 110.9 ¦ 1-65 ¦ 1-83 ¦ 97.0 ~----- ~-~-- --__~1_ _ _ __L_ 1______1_____~
Note: Compensating for samples taken:
Example 1 = 91 g / 323 g ~ ie. 323/232 Example 2 = 48 g / 332 g ~ ie. 332/284 The wet residue weights were recorded. The dry weight can be calculated from the percentage moisture. Example: atmospheric leach feed in Example 1:
108.4 g wet * [(100-36.2)/lOOJ ~ 323/232 = 96.2 g dry With the addition of sulphuric acid to the feed of the pressure leach, the extraction of copper increased.
The per cent copper remaining behind in the residue of the atmospheric leach, Stream 12, decreased from 2.52% to 1.65%. The extraction of copper increased from 96% to 97%
in Example 2.
Table 2: % Sulphur oxidation ¦ ¦ X Sulphur Oxidation ¦
¦ Exflmple ¦ Sulphur Balance Method ¦ Acid Generation Method ¦ 1 ¦ 27 ¦ n.a ~- - - - - -- - - ~ ------ - - _ _ ___ _ ~ _ ______ _ _ _ _ ____ _ _~
I__ _ _ _ _____1_____ ~ ___ _ _ _ _ _ 1_ __ ~ ___________ _ __ _~
: ', .
- 19 ~ 99333 ExamDles 3 to 6 :`
In order to determine the acid addition to the feed of the pressure leach, four tests (Examples 3 to 6) were ran consecutively. The goal of these tests was to vary the acid concentration, Stream 60, so that the copper ¦ in the feed, Stream 8, and in the filtrate, Stream 6, after the pressure leach were in equilibrium. The I following table summarizes the results of this work.
!3 10 ~ Table 3: Results of Acid Addition Acid in feed ~g/L) ¦ Cu in Feed ~ Cu in Filt.~) ¦ Gain/Los~ ¦ pH
¦ ¦ Example ¦ SStream 60) ¦ ~stream 8~ ¦ tStream 6) ¦ In F~ltrate ¦ F11t. ¦
,~ ~ _ _ __+_ _ ____ _ _ ~ _ _ _+______ _ _ ____+__ _ _ _______ ~ _____~
3 ¦ 27.3 ¦ 4.1 1 9-7 ¦ 5.6 ¦ 3.2 1 t + ~ ---------- +______~_ _ ~
during the pressure leach and the generation of sulphuric acid.
It is accordingly an object of the present invention to provide a hydrometallurgical copper extraction process wherein the oxidation of sulphide in the ore or concentrate to sulphate is reduced.
SUMMARY OF THE INVENTION
According to the invention, there is provided a process for the extraction of copper from a sulphide, copper ore or concentrate, comprising the steps of subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt; leaching the basic copper salt produced by said Pirst leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution; subjecting said leach liquor to a solvent extraction process to produce a copper concentrate solution and a raffinate comprising protons and sulphate ions in solution; extracting protons and sulphate ions from said raffinate to produce a sulphuric acid solution;
and recycling said sulphuric acid solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant.
According to a preferred embodiment, the raffinate is subjected to electrodialysis to effect said extraction of protons and sulphate ions therefrom.
Also according to the invention, there is provided a process for the extraction of copper from a .
~" ,,, ,, ", , ' " " ~ ", ~ " . ' , , ' ~ ", ', ~
,: ~ . ~- ~ , , , .." , , :, , - : : . ,: ;---` 2{~9~333~ - 3 -, , sulphide copper ore or concentrate, comprising the steps of subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt; leaching the basic copper salt produced by said first leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in : 10 solution; subjecting said leach liquor to a solvent extraction process to produce a first copper concentrate solution and a raffinate comprising protons, copper ions and sulphate ions in solution; extracting copper ions and sulphate ions from said raffinate to produce a second copper concentrate solution, recycling said second copper concentrate solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant; and subjecting said first ::
copper concentrate solution to electrowinning to recover copper values therefrom.
' .
Further according to the invention, there is provided a process for the extraction of copper from a ! sulphide copper ore or concentrate, comprising the steps of subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a solution of chloride ions and at an acidic pH
to produce an insoluble basic copper salt; adding to said solution, during said first leaching, a source of sulphate ions selected from the group consisting of sulphuric acid, copper sulphate and a metal sulphate which hydrolyzes at said acidic pH and providing sufficient of said source of sulphate ions to react said source of sulphate ions with said ore or concentrate to form said basic copper salt;
and leaching said basic copper salt in a second leaching with an acidic sulphate solution to dissolve the basic ;: :", 2~99~33 copper salt to produce a leach liquor containing copper I sulphate in solution.
In the present specification, elevated temperature means a temperature above room temperature (25C), preferably from about 125C to about 175C, and elevated pressure means an oxygen partial pressure above atmospheric pressure, preferably from about 50 psig to about 250 psig (1380 kPa).
Further objects and advantages of the invention will become apparent from the description of a preferred embodiment of the invention below.
BRIEF DESCRIPTION OF THE DRAWING
, .
The single drawing is a flow diagram of a hydrometallurgical copper extraction process according to the invention.
DETAILED DESCRIPTION OF PREFERRED EMBODIMENT
The stages of the process comprise a pressure leaching stage 38 in an autoclave, an atmospheric leaching stage 40, a solvent extraction and stripping stages 42 and 56, an electrodialysis stage 44 and an electrowinning stage 46. -After each of the leaching stages 38 and 40, filtration is carried out as indicated at 48 and 50, respectively, to separate the liquids and solids.
The process will now be described in greater detail by way of a specific example in which a chalcopyrite concentrate is treated.
.~ . . . : . ~ : : ~ , , ~ 9~3~
The copper concentrate is first ground in a ball mill, during a feed preparation step 52, to reduce the size of the particles to about 95% minus 325 'ryler mesh or smaller. Although satisfactory results are obtainable without regrinding, it has been found that there is a small but significant improvement with regrinding~
The concentrate is leached in the pressure leaching stage 38 at an elevated pressure and temperature with a lixiviant containing from about 3-15 grams per litre copper, 6-18 grams per litre chloride and about 15-35 grams per litre sulphuric acid. About 80% of the lixiviant is leach liquor which is recycled after the atmospheric leaching stage 40 and filtration 50. The remaining 20~ comprises concentrated sulphuric acid which is recycled from the electrodialysis stage 44, as will be described in more detail below.
The temperature of the leach 38 is about 150C
and the pressure about 200 psig (1380 kPa). This is total pressure comprising oxygen pressure on top of the 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.
The solids content is maintained at about 15-20%, i.e. 170-250 grams per litre solids as determined by the heat balance. A higher percentage solids would require some form of heat removal to prevent the temperature from rising above the desired limit of about 150C.
As referred to above, the lixiviant used in the pressure leach 38 is made up partly of recycled lixiviant from a previous pressure leach (Stream 8) but augmented by an acid concentrate (Stream 60) which is recycled from the ~.. . . .
"
. . -. . ~
. .. ...
~g33~
.
; electrodialysis stage 44. The immediate effect of adding the acid concentrate to the lixiviant is to increase the acidity of the lixiviant which is fed to the autoclave for the pressure leaching stage 38, but the most important effect, surprisingly, has been found to be that the addition of the acid, or ~ore specifically the sulphate ions, actually suppresses the oxidation of sulphur emanating from the concentrate during the pressure , leaching stage 38.
Typically the oxidation of sulphur that is experienced if no acid recycle is used is about 25% of the feed sulphur in the concentrate, as is the case with the process described in U.S. Patent ~,039,406. However, if acid recycle is used, it has been found that the sulphur oxidation to sulphate is reduced to about 5-10%. This improvement has substantial beneficial effects OJI the hydrometallurgical extraction process. The oxidation of sulphur to sulphate creates additional costs in several ways, such as additional oxygen required for the reaction, additional reagent required to neutralize the acid so I 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 38 takes place.
The chemistry of the reaction in the pressure leaching stage 38 is believed to be altered by the addition of the acid as follows:
No acid addition-CuFeS2 + 7/402 + 2/3H20 ~ [1/3CuSO4-2/3Cu(OH) 2 ] + 1/2Fe203 + 5/3S
~3~
_ 7 _ ~ 9 ~ 3 3 3 With acid addition:
CuFeS2 + 5/402 + 1/3H20 + l/3H2S04 ~ [l/3CuSO4-2/3Cu(OH)2]
+ 1/2Fe203 + 2S0 In both reactions, the copper is precipitated in the form of a basic copper salt, which has been found to comprise about 90% of basic copper sulphate, which contains a sulphate anion, as indicated in the reaction equations, but about 10~ of basic copper chloride is also formed. 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 recycle, thus obviating the need for the oxidation of sulphur to sulphate. Thus in the second reaction, there is a nett consumption of ~ulphate ions to form the basic copper salt.
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% is found. With acid addition, experiments indicated about 5-10~ 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 38 but are only an approximation.
In order to take advantage of the beneficial effect of the sulphuric acid recycle to inhibit the oxidation of sulphur, it is necessary to find efficient ways of adding acid into the autoclave, which has limited ability to absorb acid because the bulk of the leach ., ,. :,.
` ~9333~
liquor is recycled. However, there is some loss of leach liquor in the pressure leaching stage 38, due to venting (steam losses) and due to leach liquor carried off in the filter cake after the pressure leach stage 38 (Stream 7).
It has been found that about 20% of the volume of leach liquor is lost during each cycle in this fashion. The amount of sulphuric acid needed to suppress sulphur oxidation has been found experimentally to be about 25 grams per litre. Therefore, since this amount of acid must be contained in 20% of the volume, it must come in a concentrated form, i.e.
25 grams per litre = 150 grams per litre There is surplus acid produced in the solvent extraction stage 42 where the CUSO4 solution is changed into an H2S04 solution. However, the acid so produced is very dilute (Stream 17), only about 20-40 grams per litre, due to the nature of the solvent extraction chemistry.
The difficulty cannot effectively be overcome by simply evaporating the raffinate (Stream 17) coming from the solvent extraction stages 42 and 56 to produce a concentrated acidic solution because it is necessary to eliminate impurities in the raffinate, such as iron and zinc/ and evaporation followed by recycling would return the impurities to the pressure leaching stage 38.
This problem has been solved in the present invention by extracting sulphuric acid from the raffinate and in the particular embodiment described here, use is made of electrodialysis to effect this extraction. Thus/
by introducing the electrodialysis stage 44/ which will be described in more detail below, sulphuric acid is recovered from the raffinate in a concentrated form suitable for use in the pressure leaching stage 38. -~
~.,, ~ - , . . .
~: "",~
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. . .
-~ ~V!~3~3~
, g ., The slurry produced by the pressure leach 38 is cooled to below 100C and then filtered 48 to separate the residue from the leach liguor or lixiviant, which is recycled through a cooling tower 54 to the leaching stage 38 as noted above.
i , The residue contains the copper originally present in the concentrate as insoluble basic copper sulphate and basic copper chloride together with all the other solid materials, such as Fe2O3 ~hematite) and elemental sulphur.
There is a gain in the weight of the leach residue. Typically it has 30-40% more weight than the Peed concentrate. It has been found that the leach residue contains about 0.5-2% chloride, as well as the copper, iron oxide and sulphur, which is due to the presence of the basic copper chloride and the basic ~ 20 copper sulphate. The iron in the chalcopyrite concentrate 1 i8 converted almost completely to hematite, while sulphur is mostly converted to the elemental form with only a fraction (about 5-10%) being oxidized to sulphate, as noted above.
The leach liquor produced by the pressure leaching step 38 has much the same composition as the feed ~ lixiviant except that there is a drop in the chloride I concentration from about 12 grams per litre to about 7-10 grams per litre, depending on the conditions, due to the formation of the basic copper chloride.
The filter cake or leach residue is repulped in raffinate from the subsequent solvent extraction stage 42, which comprises an acidic sulphate solution containing ~ . .. .
~ . , ,,;;, 2 1~ 3 about 20-40 grams per litre H2S04 and a small amount of -copper, about 1-3 grams per litre.
The second leaching sta~e 40 takes place at atmospheric pressure and a temperature of about 40C for a retention time of about 15-120 minutes. The percentage solids is about 6-11% or about 70-140 grams per litre.
The final acidity of the slurry is about pH 1.5-2.0 or about 2-5 grams per litre H2SO4.
During the atmospheric leaching stage 40, the basic copper salts dissolve almost completely with very little of the iron going into solution.
Typically, the leach liquor produced after filtration 50 contains about 10-20 grams per litre copper with less than 1 grams per litre iron and about 0.3-1.0 grams per litre chloride.
The percentage solids is kept low during the atmospheric leaching stage 40 because higher copper concentrations cannot be treated satisfactorily by the subsequent solvent extraction circuit.
The copper extraction has been found to be about 97-98% based on the original feed to the pressure leaching stage 38. Iron extraction to solution has been found to be less than about 5%.
The main constituents of the solid residue after filtration 50 are hematite and elemental sulphur, as well as any gold or silver which may have been present in the original concentrate. The sulphur can be recovered by screening or flotation to separate it from the hematite into a high-grade sulphur concentrate, which can be further treated for recovery of sulphur (Stream 64). The - ~ .
~ ~ ' ' - : ' ' , , ~ r~ 20~3~3 .:
, gold and silver can be recovered by cyanidation after sulphllr is removed from the leach residue (Stream 12).
The copper leached in the atmospheric leaching stage 40 is extracted by means of solvent extraction 42 to produce a loaded copper electrolyte suitable for electrowinning 46. After the solvent extraction stage 42, the loaded organic extractant is subjected to washing and stripping 56. The high copper concentration of about 10-20 grams per litre derived from the atmospheric leaching stage 40 provides significant advantages over conventional solvent extraction/electrowinning plants because much higher loading of the organic is possible, thus reducing the size of the plant for a given tonnage of copper. Stripping of the loaded organic is effected by ! means of spent acid from the electrowinning stage 46 to obtain a pure copper sulphate solution which is then passed to the electrowinning stage 46.
Spent acid from the electrowinning stage 46 is recycled to the solvent extraction stage 42 to strip the copper from the loaded extractant.
The raffinate from the solvent extraction s'cage 42 is divided into two portions. A first portion (Stream 9) comprising two-thirds of the raffinate is recycled to the atmospheric leach stage 40. A second portion -i comprising one-third of the raffinate (Stream 18) is sent to the electrodialysis stage 44 to produce a diluate and a concentrate acid solution. The diluate solution is 0.5-3 1.0 grams per litre copper and 7-12 grams per litre sulphuric acid, and the concentrate acid solution is 5-10 grams per litre copper and 150-170 grams per litre H2S04.
Typically the concentrate stream (Stream 19) from the electrodialysis stage 44 will be about 10-20~ of Ji: ,. .. ,. . .
- 12 - ~ 9333 the feed flow, whereas the diluate stream (Stream 21) comprises the rest or 80-90% and contains the bulk of the water in the feed stream, as well as any ions that have been rejected by the membrane, such as Cu2' present in the ¦ 5 feed. It is desirable to recover such CUSO4 and this is effected in the process according to the present invention by subjecting the diluate stream to an auxiliary solvent extraction circuit 58. The circuit 58 comprises an extracting stage 64 and a stripping stage 66, for the extraction and stripping operations, respectively. The concentrate acid solution from the electrodialysis stage 44 is used as stripping acid in the stripping stage 66 to strip the copper from the loaded organic. The acid solution resulting from the stripping stage 66 is 140-160 grams per litre sulphuric acid which is recycled to the pressure leach stage 38. Typically the feed to this circuit (the diluate stream) will contain about 0.5-2.0 ; grams per litre copper and the raffinate or waste stream will contain about 0.05-0.1 grams per litre copper. This corresponds to about 0.3% of the original feed copper, considering that, for example, the feed liquor to the main solvent extraction stage 42 (Stream 11) contains about 10 grams per litre copper and is three times the flow of the waste stream from the auxiliary extracting stage 64. The overall extraction of copper in the process has been found to be as high as 99.7%. -As noted above, the residue from the pressure leaching stage 38 (Stream 7) comprises a mixture of elemental sulphur (S), hematite (Fe203) and mainly basic copper sulphate. This residue is fed to the atmospheric leaching stage 40 where the basic copper sulphate is dissolved in acid as far as possible, leaving the hematite and elemental sulphur components essentially untouched.
This produces a solution of copper sulphate (CuS04) (Stream 11~ which is fed to the solvent extraction stage .
2~9~333 42, where copper is exchanged with an organic ligand (R-H), producing acid in the aqueous stream, the raffinate (Stream 17). The reactions can be summarized as follows:
Atmospheric Leachina Stage:
CuSO4 2Cu(OH) 2 + 2HzSO4 ~ 3CuSO4 + 4H2O
Solvent Extraction Stage:
3CuSO4 + 6R-H , 3~2Cu + 3H2SO4 Thus the overall reaction can be represented as follows:
CuSO4-2Cu(OH) 2 + 6R-H ~ 3R2Cu + H2S04 + 4Hz0 There is thus one extra mole of acid produced for every three moles of Cu leached in the atmospheric leach. In order to make use of this extra mole of H2SO4, the raffinate stream from the solvent extraction stage 42 is split, as noted above, so that two-thirds thereof (Stream 9) are returned to the atmospheric leaching stage 40 and the remaining one-third (Stream 18) is fed to the electrodialysis stage 44 to produce the acid concentrate which is fed to the pressure leaching stage 38.
An additional benefit of the process according to the invention is that chloride ions lost from the pressure leach circuit into the pressure leach residue, either as insoluble basic copper chloride or as entrained solution losses in the filter cake, can be recycled along with the acid concentrate back to the pressure leach. Any chloride ions present in the pressure leach residue will report almost quantitatively to the atmospheric leach liquor and thence to the raffinate after solvent extraction. If not bled ~rom this circuit they would quickly build up to higher levels in the atmospheric leach liquor and eventually transfer to the electrowinning circuit where chloride is particularly undesirable. By - - 14 - ~a~9333 splitting the raffinate from the solvent extraction stage -,~ 42, as noted above, and treating one-third thereof through j the electrodialysis stage 44 and solvent extraction 58, this effectively recycles the chloride content back to the pressure leach thereby minimizing any chloride makeup requirements therein.
.~ .
With the process according to the present invention, relatively high copper recoveries, typically 97 to 98% at quite low pressures, such as 40C, have been obtained. Such low temperatures are known to suppress iron dissolution and test results have shown only about 200 ppm Fe with 10 grams per litre Cu in the atmospheric leach solution after filtration 50 (Stream 11). This is a marked improvement over prior art processes, such as described in U.S. Patent 4,338,168, which reports only about a 93% recovery at this temperature and requires higher temperature and/or acid levels to obtain ;~ satisfactory copper recovery values. Unfortunately, such more severe conditions also dissolves about 50% of the Fe in the feed to the atmospheric leaching stage 40 complicating the process by requiring the addition of a ¦ jarosite precipitation process to separate the Fe from the copper. In the present process the Fe is rejected in the atmospheric leach residue.
Due to the reduction in sulphur oxidation and the effective recycling of the sulphuric acid as de~cribed, the process according to the invention does not ; 30 require any special neutralization procedure. Since only about 5-10% of the sulphur is oxidized to sulphate only a '~ relatively small amount of acid is produced which can effectively be taken care of by a lime neutralization process (Stream 35) which is required in any event for the treatment 68 of the final effluent from the solvent !
r"~
~::, . : ~ , , .
- 15 -2~9~333 extraction cycle 58, which is the bleed of impurities such as Zn and Mg from the circuit.
The results of tests which were carried out for the various stages of the process will now be given in the following Examples. In Example 1 the feed to the pressure j leach did not contain acid. In Example 2, an acidic feed was charged to the leach.
10Example 1 The copper concentrate from a porphyry deposit in Highland Valley, British Columbia (Stream 1 on the flowsheet) is composed of 40.19% copper, 20.50% iron, and 29.24% sulphur. In both Examples 1 and 2, the concentrate was ground to 98~-400 mesh. ~n Example 1, the charge to the autoclave had a wet weight of 175.1 grams at 14.4%
moisture. The solution feed to the leach was a combination of 900 ml of recycled pressure leach filtrate (Stream 8) containing 1.5 grams per litre copper, less than 1 ppm iron and 11.47 grams per litre chloride and 100 ml of water. The makeup water (Stream 31 on the flowsheet) actually contained 2.8 grams of sodium chloride so that the total chloride concentration in the leach was 12.0 grams per litre. The concentrate was leached for one hour at 200 psi and 150C. Upon completion of the pressure leach, the slurry (Stream 51 on the flowsheet) was filtered. The 995 ml of filtrate, Stream 6, contained 1.0 grams per litre copper, less than 1 ppm iron, 8.3 grams per litre sodium, and 11.6 grams per litre chloride and had a pH of 3.9. The total wet weight of the residue from the pressure leach was 323.0 grams. A 91 gram sample was taken for analysis. This sample contained 32.4%
copper, 16.9% iron, 0.49% sodium and 10.4% elemental sulphur and had a moisture content of 37.7% moisture.
, - ,, . -... .
~: . ~ ', ' ' :
20~333 The residue from the pressure leach was subjected to an atmospheric leach for an hour at 40C and ; a pH 1.7. The charge to this leach consisted of 231.6 grams of solids at 37.7% moisture and 2120 ml of water with 30 ml of concentrated H2S04 (Stream 9 on the flowsheet). The slurry from this leach, Stream 10 was filtered to obtain lOg.4 grams of residue at 36.2%
moisture and a 2120 ml filtrate. The residue was washed , once by displacement and resulted in a 245 ml wash water containing 4.2 grams per litre copper and 217 ppm iron.
~ The filtrate, Stream 11, consisted of 23.2 grams per litre copper, 403 ppm iron, 420 ppm sodium, and 1.2 grams per litre chloride. The pH of the filtrate was 1.7 and had a free acid of 3.1 grams per litre. The residue from the leach contained 2.52% copper, 0.16% sodium and 28.9% iron.
The results of this example are given below.
Example 2 To reduce sulphur oxidation, sulphuric acid was added to the feed of the pressure leach. The sulphuric acid provided the sulphur needed to form Basic Copper ! Sulphate instead of oxidizing sulphur in the concentrate.
' The following equation defines th~ reaction which is 1 25 occurring:
CuFeS2 + 5/402 + 1/3H20 + 1/3H2S04 ~ [1/3CuS04-2/3Cu(OH)2]
+ 1/2Fe203 + 2S
This addition of acid to the feed of the leach reduces the sulphur oxidation from 28% to 9%. The charge to the present leach consisted of 183.0 grams of wet concentrate at 16% moisture. Since the actual liquor from a past pressure leach, Stream 8, and the concentrated acid from the electrodialysis/solvent extraction, Stream 60, were unavailable, these feeds were made synthetically.
, - . :
~, . .. : . ~ : , ., -: , . ..
- 17 ~ 2~333 The pressure leach was charged with 1000 ml of feed solution having a chloride concentration of 12.0 grams per litre and a free acid concentration of 27.0 grams per litre. The concentrate was leached for an hour at 200 psi and 150C. The slurry from this leach, Stream 51, was filtered. The 1025 ml filtrate, Stream 6, contained 6.0 grams per litre copper, 20 ppm iron, 6.0 grams per litre ~ sodium, and 10.4 grams per litre chloride. The pH of the j filtrate was 3.1. The residue weighed 331.5 grams wet ; 10 from which a 48 gram sample was taken. The residue had a moisture content of 32.6% and contained 25.4% copper, 13.6% iron, 0.23% sodium and 1~.41% elemental sulphur.
As in Example 1, the residue from the pressure leach was subjected to an atmospheric leach.
Approximately 284 grams of wet residue from the pressure ¦ leach was combined with 3500 ml of acidic water (Stream 9 on the flowsheet). The residue was leached for an hour at 40C at a pH of 1.5. The slurry from this leach was once again filtered to obtain 3205 ml of filtrate, Stream ~1, and 132.~ grams of wet residue. The residue was washed with 385 ml o~ water and produced a wash water with 3.38 grams per litre copper and 143 ppm iron. The filtrate contained 11.9 grams per litre copper, 580 ppm iron, 0.10 gram~ per litre chloride and 0.16 grams per litre sodium.
The residue had a moisture content of 28.1% and consisted of 1.65% copper, 0.03% sodium and 16.23% iron.
Comparison of ~xamples 1 and 2 To compare the effect of adding acid to the feed of the pressure leach on the leaching of copper, the following tables illustrate the copper extraction and the sulphur oxidation for both tests.
,,~,.. .
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~: .
3 3 3 : -Table 1: Copper extracti~n for tests Examples 1 and 2 r~ T~ --T---~ ---- ~~~~~~ ---~ -T---- ~
1 I Fesd Copper ¦ Residue Copper ¦ X
¦ Example ~ ------T~ ~ ~ ~ ~ Extraction ¦
¦ Dry ~t.(g~ ¦ X Cu ¦ g Cu ¦ Dry ~t. ~g) ¦ X Cu ¦ g Cu ¦
10 ¦ 1 ¦ 150.0 ¦ 40.19 ¦ 60.2 ¦ 96.2 ¦ 2.52 ¦ 2.42 ¦ 96.0 I~ - I ----t--- ~
2 ¦ 153.7 ¦ 40-19 ¦ 61.8 ¦ 110.9 ¦ 1-65 ¦ 1-83 ¦ 97.0 ~----- ~-~-- --__~1_ _ _ __L_ 1______1_____~
Note: Compensating for samples taken:
Example 1 = 91 g / 323 g ~ ie. 323/232 Example 2 = 48 g / 332 g ~ ie. 332/284 The wet residue weights were recorded. The dry weight can be calculated from the percentage moisture. Example: atmospheric leach feed in Example 1:
108.4 g wet * [(100-36.2)/lOOJ ~ 323/232 = 96.2 g dry With the addition of sulphuric acid to the feed of the pressure leach, the extraction of copper increased.
The per cent copper remaining behind in the residue of the atmospheric leach, Stream 12, decreased from 2.52% to 1.65%. The extraction of copper increased from 96% to 97%
in Example 2.
Table 2: % Sulphur oxidation ¦ ¦ X Sulphur Oxidation ¦
¦ Exflmple ¦ Sulphur Balance Method ¦ Acid Generation Method ¦ 1 ¦ 27 ¦ n.a ~- - - - - -- - - ~ ------ - - _ _ ___ _ ~ _ ______ _ _ _ _ ____ _ _~
I__ _ _ _ _____1_____ ~ ___ _ _ _ _ _ 1_ __ ~ ___________ _ __ _~
: ', .
- 19 ~ 99333 ExamDles 3 to 6 :`
In order to determine the acid addition to the feed of the pressure leach, four tests (Examples 3 to 6) were ran consecutively. The goal of these tests was to vary the acid concentration, Stream 60, so that the copper ¦ in the feed, Stream 8, and in the filtrate, Stream 6, after the pressure leach were in equilibrium. The I following table summarizes the results of this work.
!3 10 ~ Table 3: Results of Acid Addition Acid in feed ~g/L) ¦ Cu in Feed ~ Cu in Filt.~) ¦ Gain/Los~ ¦ pH
¦ ¦ Example ¦ SStream 60) ¦ ~stream 8~ ¦ tStream 6) ¦ In F~ltrate ¦ F11t. ¦
,~ ~ _ _ __+_ _ ____ _ _ ~ _ _ _+______ _ _ ____+__ _ _ _______ ~ _____~
3 ¦ 27.3 ¦ 4.1 1 9-7 ¦ 5.6 ¦ 3.2 1 t + ~ ---------- +______~_ _ ~
4 1 22.5 1 7.0 1 10-38 1 3.3 1 3.7 1 20 ~------~------------_____ _~__ __+___________~________~____~
., 1 5 1 24.0 1 8.1 1 7.1 1 -1.0 1 3.3 t ~ + ~ +
1 6 1 23.0 1 6.7 1 4.0 1 -2.7 1 3.6 1 . ~ J
Note: Pressure Leach Conditions - 60 min., 150C, 300 psi, 225 grams per litre concentrate At 27.3 grams per litre acid in the feed, the filtrate gained 5.6 grams copper indicating that too much acid had been added to the feed. At 23.0 grams per litre acid in the feed, the filtrate loss copper indicating that too little acid had been added to the feed. Based on these results, the equilibrium acid concentration in the feed to the pressure leach was set at 25.0 grams per ~ 35 litre.
.1 .
The pH of the feeds to the pressure leach were acidic; however, the filtrates had pH's of 3.2 and higher.
This difference in the pH's of the feed and filtrates indicates that the acid was consumed during the leach and is an indication that the Basic Copper Sulphate is ., - .
- ~ .
2~99333 forming from the acid in the feed and not from the oxidation of the sulphide in the feed.
To further illustrate that 97% of the copper is leached in this process, the following table summarizes the copper extractions for the tests. The extractions of copper were above 97% in all the tests except Example 4.
The sulphur oxidations were all below 7.4%
Table 4: Copper extractions for Examples 3-6 r ~ -------T~ ~ --T-----~ ~~~~~~~~~l % Cu l X Sulphur Example ¦Extraction ¦ Oxidation 3 1 97-2 1 6.73 t --~---- ___ ___~
4 1 95-4 1 7.39 ~------~--__~ _~________~ .
., 1 5 1 24.0 1 8.1 1 7.1 1 -1.0 1 3.3 t ~ + ~ +
1 6 1 23.0 1 6.7 1 4.0 1 -2.7 1 3.6 1 . ~ J
Note: Pressure Leach Conditions - 60 min., 150C, 300 psi, 225 grams per litre concentrate At 27.3 grams per litre acid in the feed, the filtrate gained 5.6 grams copper indicating that too much acid had been added to the feed. At 23.0 grams per litre acid in the feed, the filtrate loss copper indicating that too little acid had been added to the feed. Based on these results, the equilibrium acid concentration in the feed to the pressure leach was set at 25.0 grams per ~ 35 litre.
.1 .
The pH of the feeds to the pressure leach were acidic; however, the filtrates had pH's of 3.2 and higher.
This difference in the pH's of the feed and filtrates indicates that the acid was consumed during the leach and is an indication that the Basic Copper Sulphate is ., - .
- ~ .
2~99333 forming from the acid in the feed and not from the oxidation of the sulphide in the feed.
To further illustrate that 97% of the copper is leached in this process, the following table summarizes the copper extractions for the tests. The extractions of copper were above 97% in all the tests except Example 4.
The sulphur oxidations were all below 7.4%
Table 4: Copper extractions for Examples 3-6 r ~ -------T~ ~ --T-----~ ~~~~~~~~~l % Cu l X Sulphur Example ¦Extraction ¦ Oxidation 3 1 97-2 1 6.73 t --~---- ___ ___~
4 1 95-4 1 7.39 ~------~--__~ _~________~ .
5 1 97.7 1 5.64 ~_ _ __ ~ __ _ _ ~ __________~
6 1 98.0 1 7.36 1 :
L_ ~L __ _______ I
:
Example 7: Solvent ~xtraction and Stripping Process A solvent extraction and stripping test was performed and used as an example to demonstrate the solvent extraction and solvent stripping unit operations of the process as shown in the flowsheet.
A copper sulphate filtrate solution from the atmospheric leach corresponding to Stream 11 on the flowsheet, and containing 9.5 grams per litre copper and 3.75 grams per litre sulphuric acid at pH 1.6 was subjected to solvent extraction and solvent stripping in the solvent extraction unit. A 70% to 30% ratio mixture of LIX 84 to LIX 860 at 30~ volume ratio with kerosene was used for the solvent extraction unit operation. A 2 stage extraction and 2 stage stripping were used. A 2:1 organic ""~
20~9333 to aqueous ratio was used in the two solvent extraction stages and a 3:1 organic to aqueous ratio was used in the two stripping stages. Spent acid (Stream 27 on flowsheet), containing 32.9 grams per litre copper and 180 grams per litre acid from the solvent extraction and the electrowinning were mixed and used for stripping the loaded organic (Stream 15 on flowsheet) in the strippin~
circuit.
The copper sulphate pregnant electrode solution produced from the stripping circuit (Stream 25 on flowsheet) was found to contain 43.5 grams per litre copper and 146 grams per litre free acid, and was sent for electrowinning. The raffinate produced (Stream 17 on flowsheet) was found to contain 0.73 grams per litre copper and 19.4 grams per litre sulphuric acid. About 1/3 of this raffinate (Stream 18 on flowsheet) will be further treated in the electrodialysis unit operation and 2/3 of raffinate (Stream 9 on flowsheet) to be recycled back to the atmospheric leach unit operation.
The loaded organic (Stream 15 on flowsheet) was found to contain 7.9 grams per litre of copper, and the stripped organic (Stream 20 on flowsheet) was found to contain 3.39 grams per litre copper.
Example 8: Electrodialysis An electrodialysis test was performed and used as an example to illustrate the electrodialysis unit operation of the process shown in the flowsheet.
The raffinate from the solvent extraction circuit (Stream 18 on flowsheet) was subjected to electrodialysis in the electrodialysis unit. The raffinate solutions from the solvent extraction were mixed '`;' ' " ' . ' . ' '. ~ ~, ' ' ' . ' ' .~ ' ' ' ~ - 22 -2 ~ ., 3 A~
and used as feed to the electrodialysis. The raffinate solution was found to contain 700 mg/l copper and 20 grams per litre sulphuric acid. SELEMIONTM CMV catonic and AAV
anionic membranes were used. These membranes are designed to pass both monovalent and divalent ions into the concentrate. The raffinate solution was passed through an activated carbon column and a polish filter to remove any organic and suspended solids. The total organic carbon in the raffinate was reduced from 7.84 ppm to 1.53 ppm after passing through the activated carbon column and the polish filter. The solution was then fed to the electrodialysis unit at a rate of 9.5 l/hr~m2. A current was applied between the electrodes to give a current density of 1000 A/m2 The temperature was controlled at below 40C. The electrode compartments were rinsed with a rinse solution containing 1 molar sulphuric acid.
The final diluate solution (Stream 21 on flowsheet) was analyzed, and was found to contain 261 mg/l copper and 9.5 grams per litre sulphuric acid. The final concentrate acid solution (Stream 19 on flowsheet) was found to contain 7.32 grams per litre copper and 168 grams per litre sulphuric acid~ Both the diluate solution and the concentrate acid solution was treated in the auxiliary solvent extraction and stripping process to produce a final raffinate solution (Stream 63 on the flowsheet) containing less than 100 mg/l copper which is sent to the effluent treatment stage 68.
' 30 Example 9: Auxiliary Solvent Extraction A solvent extraction was performed and used as ! an example to illustrate the auxiliary solvent extraction (E3) and stripping (S3) unit operations of the process shown in the flowsheet.
;
r~ ' ' . :
,', . . ` ' ~ ' - 23 -2~99~33 The diluate solution from the electrodialysis unit (Stream 21 on flowsheet) was subjected to third stage (E3) solvent extraction 64. Diluate solutions from electrodialysis were mixed and used as feed to this test.
The diluate was found to contain 264 mg/l copper and 9.0 grams per litre sulphuric acid at pH 1.02. A 70% to 30%
volume ratio mixture of LIX 84 to LIX 860 at 30% volume ratio with kerosene was used for the E3-S3 solvent extraction and stripping operations. A one stage extraction and one stage stripping was used at a 1:1 organic to aqueous ratio in the extraction stage and a 10:1 organic to aqueous ratio in the stripping stage.
Concentrate acid solution (Stream 19 on flowsheet), containing 5.36 grams per litre copper and 180 grams per litre acid was used for stripping the loaded organic (5tream 61) in the stripping circuit.
The copper sulphate solution produced (Stream 60) from the stripping circuit was found to contain 8.8 grams per litre copper and 161.3 grams per litre free acid, and was recycled back to pressure leaching. The final raffinate produced (Stream 63) was found to contain 66 mg/l copper and 12.3 grams per litre sulphuric acid, and the raffinate is sent to effluent treatment operation.
The loaded organic (Stream 61) was found to contain 1.7 grams per litre of copper, and the stripped organic (Stream 62) was found to contain 1.1 grams per litre copper.
Example 10: Copper Electrowinnina A copper electrowinning test was performed and used as an example to demonstrate the copper - 24 - 2 ag93 3 ~
electrowinning unit operations of the process shown in the flowsheet.
A copper sulphate electrolyte produced from the solvent extraction unit operation (Stream 25 on the flowsheet) was subjected to electrowinning in the electrowinning unit. Copper electrolyte from the solvent ! extraction stage was used in this test and the copper electrolyte was 43.5 grams per litre copper and 146 grams per litre sulphuric acid. About 10 mg/l of animal glue ~I was added to the electrolyte solution to provide cathode deposit control and to counteract the negative effect of trace impurities on cathode deposit. A current was applied between the electrodes to give a current density of 300 A/m2. A voltage drop of 2 volts per cell from anode to cathode was used. The temperature was maintained at 35C and the unit was operated for 8.4 hours.
A high quality copper was produced at the cathodes (Stream 28 on flowsheet) at a current efficiency of 97.8%. The spent acid from electrowinning (Stream 27 on flowsheet) was found to contain 28.5 grams per litre copper and 177 grams per litre sulphuric acid, and was recycled and used as stripping acid for the solvent extraction unit.
While the process has been described with specific reference to a chalcopyrite concentrate, it will be appreciated that the process can also be applied to concentrates of other copper sulphide minerals, such as bornite (Cu5FeS4), covellite (CuS), chalcocite (Cu2S), enargite (Cu3AsS4), tetrahedrite (Cu3SbS3), and the like, or mixtures thereof.
,....................................................... .
,, ~. ,: , ~
- 25 - ~ ~99~33 In the above description the electrodialysis stage 44 is selective, i.e. copper ions are in the diluate stream (Stream 21).
As an alternative to subjecting the raffinate from the solvent extraction 42 to the electrodialysis step 44, to produce a concentrated sulphuri~ acid solution for recycling to the first leaching step 38, it is also possible to produce a concentrated copper sulphate solution by subjecting the raffinate to the electrodialysis step 44, by the choice of a suitable membrane and recycling the concentrated copper sulphate as produced to the first leach 38 to serve as a source of sulphate ions required for the leach 38. In this case the electrodialysis is non-selective, i.e. the copper ions are allowed into the concentrate stream (Stream 19) with the acid. The reaction taking place in the pressure leach stage 38 can be represented as follows:
CuFeS~ + 5/402 + H20 + l/2CuS04 ~ [l/2CuS04 Cu(OH)2] +
l/2Fe2~3 + 2S
Electrodialysis processes are described in U.S.
Patents 5,064,538, 5,084,180 and 5,110,432, the entire contents of which is incorporated herein by reference. A
description and example of the selective process is given in U.S. Patent 5,084,180, column 5, line 22 to line 19, column 6 and Example 5 of U.S. Patent 5,064,538, respectively. This relates to a zinc sulphate solution, but is also applicable to a copper sulphate solution. An example of a non-selective process is given in Example 3 (column 11) of U.S. Patent 5,110,432.
In yet a further alternative method, the sulphate ions may be provided from another source, which may be external to the rest of the process, such as by the , ... . . . .
3, 3 3 addition of a metal sulphate, which will hydrolyze under the leaching conditions and thus produce acid in situ, such as Fe2(so3)3~ to the pressure leach stage 38, in which case the reaction taking place can be represented as follows:
CuFeS2 + 5/402 + 2/3H20 + 1/9Fe2(S04) [l/3CuSO4-2/3Cu(OH)2] + 11/18Fe2O3 + 2S0 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.
: , ,: :
,j , . . .
" ~ -
L_ ~L __ _______ I
:
Example 7: Solvent ~xtraction and Stripping Process A solvent extraction and stripping test was performed and used as an example to demonstrate the solvent extraction and solvent stripping unit operations of the process as shown in the flowsheet.
A copper sulphate filtrate solution from the atmospheric leach corresponding to Stream 11 on the flowsheet, and containing 9.5 grams per litre copper and 3.75 grams per litre sulphuric acid at pH 1.6 was subjected to solvent extraction and solvent stripping in the solvent extraction unit. A 70% to 30% ratio mixture of LIX 84 to LIX 860 at 30~ volume ratio with kerosene was used for the solvent extraction unit operation. A 2 stage extraction and 2 stage stripping were used. A 2:1 organic ""~
20~9333 to aqueous ratio was used in the two solvent extraction stages and a 3:1 organic to aqueous ratio was used in the two stripping stages. Spent acid (Stream 27 on flowsheet), containing 32.9 grams per litre copper and 180 grams per litre acid from the solvent extraction and the electrowinning were mixed and used for stripping the loaded organic (Stream 15 on flowsheet) in the strippin~
circuit.
The copper sulphate pregnant electrode solution produced from the stripping circuit (Stream 25 on flowsheet) was found to contain 43.5 grams per litre copper and 146 grams per litre free acid, and was sent for electrowinning. The raffinate produced (Stream 17 on flowsheet) was found to contain 0.73 grams per litre copper and 19.4 grams per litre sulphuric acid. About 1/3 of this raffinate (Stream 18 on flowsheet) will be further treated in the electrodialysis unit operation and 2/3 of raffinate (Stream 9 on flowsheet) to be recycled back to the atmospheric leach unit operation.
The loaded organic (Stream 15 on flowsheet) was found to contain 7.9 grams per litre of copper, and the stripped organic (Stream 20 on flowsheet) was found to contain 3.39 grams per litre copper.
Example 8: Electrodialysis An electrodialysis test was performed and used as an example to illustrate the electrodialysis unit operation of the process shown in the flowsheet.
The raffinate from the solvent extraction circuit (Stream 18 on flowsheet) was subjected to electrodialysis in the electrodialysis unit. The raffinate solutions from the solvent extraction were mixed '`;' ' " ' . ' . ' '. ~ ~, ' ' ' . ' ' .~ ' ' ' ~ - 22 -2 ~ ., 3 A~
and used as feed to the electrodialysis. The raffinate solution was found to contain 700 mg/l copper and 20 grams per litre sulphuric acid. SELEMIONTM CMV catonic and AAV
anionic membranes were used. These membranes are designed to pass both monovalent and divalent ions into the concentrate. The raffinate solution was passed through an activated carbon column and a polish filter to remove any organic and suspended solids. The total organic carbon in the raffinate was reduced from 7.84 ppm to 1.53 ppm after passing through the activated carbon column and the polish filter. The solution was then fed to the electrodialysis unit at a rate of 9.5 l/hr~m2. A current was applied between the electrodes to give a current density of 1000 A/m2 The temperature was controlled at below 40C. The electrode compartments were rinsed with a rinse solution containing 1 molar sulphuric acid.
The final diluate solution (Stream 21 on flowsheet) was analyzed, and was found to contain 261 mg/l copper and 9.5 grams per litre sulphuric acid. The final concentrate acid solution (Stream 19 on flowsheet) was found to contain 7.32 grams per litre copper and 168 grams per litre sulphuric acid~ Both the diluate solution and the concentrate acid solution was treated in the auxiliary solvent extraction and stripping process to produce a final raffinate solution (Stream 63 on the flowsheet) containing less than 100 mg/l copper which is sent to the effluent treatment stage 68.
' 30 Example 9: Auxiliary Solvent Extraction A solvent extraction was performed and used as ! an example to illustrate the auxiliary solvent extraction (E3) and stripping (S3) unit operations of the process shown in the flowsheet.
;
r~ ' ' . :
,', . . ` ' ~ ' - 23 -2~99~33 The diluate solution from the electrodialysis unit (Stream 21 on flowsheet) was subjected to third stage (E3) solvent extraction 64. Diluate solutions from electrodialysis were mixed and used as feed to this test.
The diluate was found to contain 264 mg/l copper and 9.0 grams per litre sulphuric acid at pH 1.02. A 70% to 30%
volume ratio mixture of LIX 84 to LIX 860 at 30% volume ratio with kerosene was used for the E3-S3 solvent extraction and stripping operations. A one stage extraction and one stage stripping was used at a 1:1 organic to aqueous ratio in the extraction stage and a 10:1 organic to aqueous ratio in the stripping stage.
Concentrate acid solution (Stream 19 on flowsheet), containing 5.36 grams per litre copper and 180 grams per litre acid was used for stripping the loaded organic (5tream 61) in the stripping circuit.
The copper sulphate solution produced (Stream 60) from the stripping circuit was found to contain 8.8 grams per litre copper and 161.3 grams per litre free acid, and was recycled back to pressure leaching. The final raffinate produced (Stream 63) was found to contain 66 mg/l copper and 12.3 grams per litre sulphuric acid, and the raffinate is sent to effluent treatment operation.
The loaded organic (Stream 61) was found to contain 1.7 grams per litre of copper, and the stripped organic (Stream 62) was found to contain 1.1 grams per litre copper.
Example 10: Copper Electrowinnina A copper electrowinning test was performed and used as an example to demonstrate the copper - 24 - 2 ag93 3 ~
electrowinning unit operations of the process shown in the flowsheet.
A copper sulphate electrolyte produced from the solvent extraction unit operation (Stream 25 on the flowsheet) was subjected to electrowinning in the electrowinning unit. Copper electrolyte from the solvent ! extraction stage was used in this test and the copper electrolyte was 43.5 grams per litre copper and 146 grams per litre sulphuric acid. About 10 mg/l of animal glue ~I was added to the electrolyte solution to provide cathode deposit control and to counteract the negative effect of trace impurities on cathode deposit. A current was applied between the electrodes to give a current density of 300 A/m2. A voltage drop of 2 volts per cell from anode to cathode was used. The temperature was maintained at 35C and the unit was operated for 8.4 hours.
A high quality copper was produced at the cathodes (Stream 28 on flowsheet) at a current efficiency of 97.8%. The spent acid from electrowinning (Stream 27 on flowsheet) was found to contain 28.5 grams per litre copper and 177 grams per litre sulphuric acid, and was recycled and used as stripping acid for the solvent extraction unit.
While the process has been described with specific reference to a chalcopyrite concentrate, it will be appreciated that the process can also be applied to concentrates of other copper sulphide minerals, such as bornite (Cu5FeS4), covellite (CuS), chalcocite (Cu2S), enargite (Cu3AsS4), tetrahedrite (Cu3SbS3), and the like, or mixtures thereof.
,....................................................... .
,, ~. ,: , ~
- 25 - ~ ~99~33 In the above description the electrodialysis stage 44 is selective, i.e. copper ions are in the diluate stream (Stream 21).
As an alternative to subjecting the raffinate from the solvent extraction 42 to the electrodialysis step 44, to produce a concentrated sulphuri~ acid solution for recycling to the first leaching step 38, it is also possible to produce a concentrated copper sulphate solution by subjecting the raffinate to the electrodialysis step 44, by the choice of a suitable membrane and recycling the concentrated copper sulphate as produced to the first leach 38 to serve as a source of sulphate ions required for the leach 38. In this case the electrodialysis is non-selective, i.e. the copper ions are allowed into the concentrate stream (Stream 19) with the acid. The reaction taking place in the pressure leach stage 38 can be represented as follows:
CuFeS~ + 5/402 + H20 + l/2CuS04 ~ [l/2CuS04 Cu(OH)2] +
l/2Fe2~3 + 2S
Electrodialysis processes are described in U.S.
Patents 5,064,538, 5,084,180 and 5,110,432, the entire contents of which is incorporated herein by reference. A
description and example of the selective process is given in U.S. Patent 5,084,180, column 5, line 22 to line 19, column 6 and Example 5 of U.S. Patent 5,064,538, respectively. This relates to a zinc sulphate solution, but is also applicable to a copper sulphate solution. An example of a non-selective process is given in Example 3 (column 11) of U.S. Patent 5,110,432.
In yet a further alternative method, the sulphate ions may be provided from another source, which may be external to the rest of the process, such as by the , ... . . . .
3, 3 3 addition of a metal sulphate, which will hydrolyze under the leaching conditions and thus produce acid in situ, such as Fe2(so3)3~ to the pressure leach stage 38, in which case the reaction taking place can be represented as follows:
CuFeS2 + 5/402 + 2/3H20 + 1/9Fe2(S04) [l/3CuSO4-2/3Cu(OH)2] + 11/18Fe2O3 + 2S0 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.
: , ,: :
,j , . . .
" ~ -
Claims (14)
1. A process for the extraction of copper from a sulphide copper ore or concentrate, comprising the steps of:
subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt;
leaching the basic copper salt produced by said first leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution;
subjecting said leach liquor to a solvent extraction process to produce a copper concentrate solution and a raffinate comprising protons and sulphate ions in solution;
extracting protons and sulphate ions from said raffinate to produce a sulphuric acid solution; and recycling said sulphuric acid solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant.
subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt;
leaching the basic copper salt produced by said first leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution;
subjecting said leach liquor to a solvent extraction process to produce a copper concentrate solution and a raffinate comprising protons and sulphate ions in solution;
extracting protons and sulphate ions from said raffinate to produce a sulphuric acid solution; and recycling said sulphuric acid solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant.
2. The process according to claim 1, wherein said raffinate is subjected to electrodialysis to effect said extraction of protons and sulphate ions therefrom.
3. The process according to claim 2, wherein said raffinate is split into a first portion comprising about two-thirds of said raffinate and a second portion comprising about one-third of said raffinate and wherein said first portion is recycled to said second leaching at atmospheric pressure and wherein said second portion comprises said raffinate which is subjected to said electrodialysis.
4. The process according to claim 3, wherein said electrodialysis produces a concentrate stream containing said sulphuric acid solution and a diluate stream containing copper ions in solution.
5. The process according to claim 4, further comprising the steps of subjecting said diluate stream to an auxiliary solvent extraction step to extract said copper ions therefrom and combining said extracted copper ions with said concentrate stream containing said sulphuric acid, prior to recycling said sulphuric acid to said first leaching at elevated temperature and pressure.
6. The process according to claim 3, further comprising the steps of extracting chloride ions from said second raffinate portion during said electrodialysis and recycling said chloride ions with said sulphuric acid solution to said first leaching at elevated temperature and pressure.
7. The process according to claim 1, wherein said elevated temperature is from about 125°C - 175°C
8. The process according to claim 1, wherein said elevated pressure comprises an oxygen partial pressure from about 50 psig to about 250 psig.
9. The process according to claim 1, wherein said insoluble basic copper salt comprises a mixture of basic copper sulphate and basic copper chloride.
10. A process for the extraction of copper from a sulphide copper ore or concentrate, comprising the steps of:
subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt;
leaching the basic copper salt produced by said first leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution;
subjecting said leach liquor to a solvent extraction process to produce a first copper concentrate solution and a raffinate comprising protons, copper ions and sulphate ions in solution;
extracting copper ions and sulphate ions from said raffinate to produce a second copper concentrate solution;
recycling said second copper concentrate solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant; and subjecting said first copper concentrate solution to electrowinning to recover copper values therefrom.
subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a lixiviant comprising an acidic solution of chloride and sulphate ions to produce an insoluble basic copper salt;
leaching the basic copper salt produced by said first leaching step in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution;
subjecting said leach liquor to a solvent extraction process to produce a first copper concentrate solution and a raffinate comprising protons, copper ions and sulphate ions in solution;
extracting copper ions and sulphate ions from said raffinate to produce a second copper concentrate solution;
recycling said second copper concentrate solution to said first leaching at elevated temperature and pressure to serve as a source of said sulphate ions in said lixiviant; and subjecting said first copper concentrate solution to electrowinning to recover copper values therefrom.
11. The process according to claim 10, wherein said raffinate is subjected to electrodialysis to effect said extraction of copper and sulphate ions therefrom.
12. The process according to claim 11 wherein said raffinate is split into a first portion comprising about two-thirds of said raffinate and a second portion comprising about one-third of said raffinate and wherein said first portion is recycled to said second leaching at atmospheric pressure and wherein said second portion comprises said raffinate which is subjected to said electrodialysis.
13. The process according to claim 12, further comprising the steps of extracting chloride ions from said second raffinate portion during said electrodialysis and recycling said chloride ions with said second copper concentrate solution to said first leaching at elevated temperature and pressure.
14. A process for the extraction of copper from a sulphide copper ore or concentrate, comprising the steps of:
subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a solution of chloride ions and at an acidic pH to produce an insoluble basic copper salt;
adding to said solution, during said first leaching, a source of sulphate ions selected from the group consisting of sulphuric acid, copper sulphate and a metal sulphate which hydrolyzes at said acidic pH and providing sufficient of said source of sulphate ions to react said source of sulphate ions with said ore or concentrate so that there is a nett consumption of sulphate ions from said source of sulphate ions to form said basic copper salt; and leaching said basic copper salt in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution.
subjecting the ore or concentrate to a first leaching at an elevated temperature and pressure in the presence of oxygen and a solution of chloride ions and at an acidic pH to produce an insoluble basic copper salt;
adding to said solution, during said first leaching, a source of sulphate ions selected from the group consisting of sulphuric acid, copper sulphate and a metal sulphate which hydrolyzes at said acidic pH and providing sufficient of said source of sulphate ions to react said source of sulphate ions with said ore or concentrate so that there is a nett consumption of sulphate ions from said source of sulphate ions to form said basic copper salt; and leaching said basic copper salt in a second leaching with an acidic sulphate solution to dissolve the basic copper salt to produce a leach liquor containing copper sulphate in solution.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA002099333A CA2099333A1 (en) | 1993-06-28 | 1993-06-28 | Chloride assisted hydrometallurgical copper extraction |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA002099333A CA2099333A1 (en) | 1993-06-28 | 1993-06-28 | Chloride assisted hydrometallurgical copper extraction |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA2099333A1 true CA2099333A1 (en) | 1994-12-29 |
Family
ID=4151851
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA002099333A Abandoned CA2099333A1 (en) | 1993-06-28 | 1993-06-28 | Chloride assisted hydrometallurgical copper extraction |
Country Status (1)
| Country | Link |
|---|---|
| CA (1) | CA2099333A1 (en) |
Cited By (4)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| RU2179192C2 (en) * | 1995-06-07 | 2002-02-10 | Коминко Енджиниэринг Сэвисиз Элтиди | Method of extraction of metal |
| WO2002050321A3 (en) * | 2000-11-29 | 2003-01-03 | Walter Curlook | Acid leaching of nickel laterite ores for the extraction of their nickel and cobalt values |
| CN110563021A (en) * | 2019-10-16 | 2019-12-13 | 大冶有色金属有限责任公司 | method and device for harmlessly treating and recycling basic copper chloride |
| CN116272424A (en) * | 2023-04-20 | 2023-06-23 | 中国长江三峡集团有限公司 | CuFeS 2 Modified catalytic ceramic membrane and preparation method and application thereof |
-
1993
- 1993-06-28 CA CA002099333A patent/CA2099333A1/en not_active Abandoned
Cited By (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| RU2179192C2 (en) * | 1995-06-07 | 2002-02-10 | Коминко Енджиниэринг Сэвисиз Элтиди | Method of extraction of metal |
| WO2002050321A3 (en) * | 2000-11-29 | 2003-01-03 | Walter Curlook | Acid leaching of nickel laterite ores for the extraction of their nickel and cobalt values |
| CN110563021A (en) * | 2019-10-16 | 2019-12-13 | 大冶有色金属有限责任公司 | method and device for harmlessly treating and recycling basic copper chloride |
| CN110563021B (en) * | 2019-10-16 | 2023-04-07 | 大冶有色金属有限责任公司 | Method and device for harmless treatment and recovery of basic copper chloride |
| CN116272424A (en) * | 2023-04-20 | 2023-06-23 | 中国长江三峡集团有限公司 | CuFeS 2 Modified catalytic ceramic membrane and preparation method and application thereof |
| CN116272424B (en) * | 2023-04-20 | 2024-04-26 | 中国长江三峡集团有限公司 | CuFeS2Modified catalytic ceramic membrane and preparation method and application thereof |
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