CN1328398C - Method for extracting platinum-palladium and base metal from platinum metal sulphide ore - Google Patents
Method for extracting platinum-palladium and base metal from platinum metal sulphide ore Download PDFInfo
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- CN1328398C CN1328398C CNB2004100401018A CN200410040101A CN1328398C CN 1328398 C CN1328398 C CN 1328398C CN B2004100401018 A CNB2004100401018 A CN B2004100401018A CN 200410040101 A CN200410040101 A CN 200410040101A CN 1328398 C CN1328398 C CN 1328398C
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- BASFCYQUMIYNBI-UHFFFAOYSA-N platinum Chemical compound [Pt] BASFCYQUMIYNBI-UHFFFAOYSA-N 0.000 title claims abstract description 110
- KDLHZDBZIXYQEI-UHFFFAOYSA-N palladium Substances [Pd] KDLHZDBZIXYQEI-UHFFFAOYSA-N 0.000 title claims abstract description 68
- 238000000034 method Methods 0.000 title claims abstract description 50
- 229910052763 palladium Inorganic materials 0.000 title claims abstract description 35
- 229910052976 metal sulfide Inorganic materials 0.000 title description 6
- 239000010953 base metal Substances 0.000 title description 3
- PXHVJJICTQNCMI-UHFFFAOYSA-N Nickel Chemical compound [Ni] PXHVJJICTQNCMI-UHFFFAOYSA-N 0.000 claims abstract description 66
- 229910052697 platinum Inorganic materials 0.000 claims abstract description 54
- 229910052751 metal Inorganic materials 0.000 claims abstract description 48
- 239000002184 metal Substances 0.000 claims abstract description 48
- 229910052802 copper Inorganic materials 0.000 claims abstract description 36
- 239000010949 copper Substances 0.000 claims abstract description 36
- 229910052759 nickel Inorganic materials 0.000 claims abstract description 34
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 32
- 229910017052 cobalt Inorganic materials 0.000 claims abstract description 29
- 239000010941 cobalt Substances 0.000 claims abstract description 29
- GUTLYIVDDKVIGB-UHFFFAOYSA-N cobalt atom Chemical compound [Co] GUTLYIVDDKVIGB-UHFFFAOYSA-N 0.000 claims abstract description 29
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 claims abstract description 21
- 239000002893 slag Substances 0.000 claims abstract description 13
- 238000007670 refining Methods 0.000 claims abstract description 6
- 239000007788 liquid Substances 0.000 claims description 61
- 230000029087 digestion Effects 0.000 claims description 48
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 24
- 239000007787 solid Substances 0.000 claims description 19
- 229910052500 inorganic mineral Inorganic materials 0.000 claims description 18
- 239000011707 mineral Substances 0.000 claims description 18
- 239000007789 gas Substances 0.000 claims description 17
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 15
- 229910052760 oxygen Inorganic materials 0.000 claims description 15
- 239000001301 oxygen Substances 0.000 claims description 15
- SUKJFIGYRHOWBL-UHFFFAOYSA-N sodium hypochlorite Chemical compound [Na+].Cl[O-] SUKJFIGYRHOWBL-UHFFFAOYSA-N 0.000 claims description 15
- QWPPOHNGKGFGJK-UHFFFAOYSA-N hypochlorous acid Chemical compound ClO QWPPOHNGKGFGJK-UHFFFAOYSA-N 0.000 claims description 13
- 239000000203 mixture Substances 0.000 claims description 12
- 235000008733 Citrus aurantifolia Nutrition 0.000 claims description 7
- 235000011941 Tilia x europaea Nutrition 0.000 claims description 7
- 239000004571 lime Substances 0.000 claims description 7
- 238000000605 extraction Methods 0.000 claims description 6
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 3
- MYMOFIZGZYHOMD-UHFFFAOYSA-N Dioxygen Chemical compound O=O MYMOFIZGZYHOMD-UHFFFAOYSA-N 0.000 claims description 2
- 229910001882 dioxygen Inorganic materials 0.000 claims description 2
- 238000002386 leaching Methods 0.000 abstract description 35
- -1 platinum group metals Chemical class 0.000 abstract description 21
- 238000005516 engineering process Methods 0.000 abstract description 17
- 239000002253 acid Substances 0.000 abstract description 15
- 239000000463 material Substances 0.000 abstract description 5
- 150000002739 metals Chemical class 0.000 abstract description 5
- 230000003647 oxidation Effects 0.000 abstract description 5
- 238000007254 oxidation reaction Methods 0.000 abstract description 5
- 238000011084 recovery Methods 0.000 abstract description 5
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 abstract description 4
- 229910052742 iron Inorganic materials 0.000 abstract description 3
- 229910000510 noble metal Inorganic materials 0.000 abstract 2
- 239000007800 oxidant agent Substances 0.000 abstract 1
- 231100000331 toxic Toxicity 0.000 abstract 1
- 230000002588 toxic effect Effects 0.000 abstract 1
- 239000007864 aqueous solution Substances 0.000 description 24
- 238000006243 chemical reaction Methods 0.000 description 20
- 239000012141 concentrate Substances 0.000 description 16
- 238000005188 flotation Methods 0.000 description 15
- 239000010970 precious metal Substances 0.000 description 13
- 238000001556 precipitation Methods 0.000 description 9
- 238000000746 purification Methods 0.000 description 9
- 239000000243 solution Substances 0.000 description 9
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 description 8
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 8
- 238000010438 heat treatment Methods 0.000 description 8
- 238000002156 mixing Methods 0.000 description 8
- 238000003723 Smelting Methods 0.000 description 6
- RWSOTUBLDIXVET-UHFFFAOYSA-O sulfonium Chemical compound [SH3+] RWSOTUBLDIXVET-UHFFFAOYSA-O 0.000 description 5
- XFXPMWWXUTWYJX-UHFFFAOYSA-N Cyanide Chemical compound N#[C-] XFXPMWWXUTWYJX-UHFFFAOYSA-N 0.000 description 4
- 239000000284 extract Substances 0.000 description 4
- 230000035484 reaction time Effects 0.000 description 4
- 230000008901 benefit Effects 0.000 description 3
- 239000003153 chemical reaction reagent Substances 0.000 description 3
- 238000000638 solvent extraction Methods 0.000 description 3
- 239000000126 substance Substances 0.000 description 3
- 229910018879 Pt—Pd Inorganic materials 0.000 description 2
- 229910004298 SiO 2 Inorganic materials 0.000 description 2
- 238000010521 absorption reaction Methods 0.000 description 2
- 150000001875 compounds Chemical class 0.000 description 2
- 238000003912 environmental pollution Methods 0.000 description 2
- 238000005342 ion exchange Methods 0.000 description 2
- 238000002203 pretreatment Methods 0.000 description 2
- 239000002904 solvent Substances 0.000 description 2
- 238000006467 substitution reaction Methods 0.000 description 2
- 229910000570 Cupronickel Inorganic materials 0.000 description 1
- 229910017709 Ni Co Inorganic materials 0.000 description 1
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical group [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 102100040653 Tryptophan 2,3-dioxygenase Human genes 0.000 description 1
- 101710136122 Tryptophan 2,3-dioxygenase Proteins 0.000 description 1
- 229910052782 aluminium Inorganic materials 0.000 description 1
- 238000007664 blowing Methods 0.000 description 1
- 238000010276 construction Methods 0.000 description 1
- BWFPGXWASODCHM-UHFFFAOYSA-N copper monosulfide Chemical compound [Cu]=S BWFPGXWASODCHM-UHFFFAOYSA-N 0.000 description 1
- YOCUPQPZWBBYIX-UHFFFAOYSA-N copper nickel Chemical compound [Ni].[Cu] YOCUPQPZWBBYIX-UHFFFAOYSA-N 0.000 description 1
- 238000002474 experimental method Methods 0.000 description 1
- 238000009852 extractive metallurgy Methods 0.000 description 1
- 238000007667 floating Methods 0.000 description 1
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 1
- 229910052737 gold Inorganic materials 0.000 description 1
- 239000010931 gold Substances 0.000 description 1
- 229910052595 hematite Inorganic materials 0.000 description 1
- 239000011019 hematite Substances 0.000 description 1
- 238000009854 hydrometallurgy Methods 0.000 description 1
- 239000000543 intermediate Substances 0.000 description 1
- 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 1
- 229910052749 magnesium Inorganic materials 0.000 description 1
- 238000012423 maintenance Methods 0.000 description 1
- 230000035800 maturation Effects 0.000 description 1
- 238000005554 pickling Methods 0.000 description 1
- 239000000843 powder Substances 0.000 description 1
- 150000003242 quaternary ammonium salts Chemical class 0.000 description 1
- 239000000376 reactant Substances 0.000 description 1
- 239000011347 resin Substances 0.000 description 1
- 229920005989 resin Polymers 0.000 description 1
- 238000000926 separation method Methods 0.000 description 1
- 229910052710 silicon Inorganic materials 0.000 description 1
- 239000010703 silicon Substances 0.000 description 1
- 239000010802 sludge Substances 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 230000001360 synchronised effect Effects 0.000 description 1
- 150000003512 tertiary amines Chemical class 0.000 description 1
- 238000009279 wet oxidation reaction Methods 0.000 description 1
- 229910052725 zinc Inorganic materials 0.000 description 1
- 239000011701 zinc Substances 0.000 description 1
Classifications
-
- 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
- Manufacture And Refinement Of Metals (AREA)
Abstract
The present invention relates to a method for extracting platinum group metals, copper, nickel, cobalt and other noble metals from the sulfide ore containing platinum group noble metals by a fully wet method, which comprises the process that firstly, material is pressurized for oxidation and boiled; secondly, Cu, Ni, Co, Fe and other heavy nonferrous metals are leached out from pressured and boiled slag; thirdly, platinum group metals are leached out from acid leached slag under normal pressure normal by adding acid and an oxidizing agent under normal pressure; fourthly, the traditional technology is used for enriching and refining platinum group metals. The method of the present invention has the characteristics of convenient technical process, no toxic substane, no pollution and strong adaptability of material. The recovery rate of platinum is larger than 94%, the recovery rate of palladium is larger than 96%, and the leaching rates of copper, nickel and cobalt reach more than 98%.
Description
Technical field
The present invention relates to the full wet processing of a kind of employing and handle low-grade platinum metals sulphide ores or its flotation concentrate, extract the method for platinum metals and base metal, particularly extract the method for platinum, palladium and copper, nickel, cobalt valuable metal.
Background technology
The tradition of platinum group metallic copper Ni sulphide mine is smelted extraction process and is all adopted the enrichment of pyrometallurgical smelting technology usually earlier.Its process system with the selecting and smelting technology of heavy non-ferrous metal as carrier.Flotation concentrate is earlier after roasting or oven dry, with the low sulfonium of output copper nickel behind electric furnace or the flash stove matte smelting divided silicon hydrochlorate, the platinum metals is trapped in the low sulfonium, low then sulfonium obtains the higher high sulfonium of precious metal grade through oxygen blowing, handle high sulfonium with the method for hydrometallurgy again, constantly heavy non-ferrous metal such as separating nickel, copper, cobalt obtain the concentration of precious metal thing, also claim the precious metal concentrate, that carries out at last precious metal again is separated from each other the purification refining.The tradition pyrometallurgical smelting process is for extracting the platinum metals, and flow process is too tediously long, and technical sophistication be difficult to obtain higher metal recovery rate, and environmental pollution is serious.When heavy non-ferrous metal grades such as the copper of association, nickel are on the low side, the pyrometallurgical smelting difficulty, the technology cost up, economic benefit is affected.And application HCl+Cl
2When the platinum metals sulphide ores is directly handled in the wet oxidation acidleach, must treat all heavy non-ferrous metal sulfide minerals and CaO, MgO, Al
2O
3Behind whole molten the finishing of oxide compound, precious metal could dissolve, and the reagent consumption is very big, the equipment anticorrosion difficulty, and the solution composition complexity, environmental pollution is serious, and from economy and technical standpoint analysis, its use value is subjected to bigger restriction.
Pressurization is leached as a kind of non-ferrous extractive metallurgy technology, has with the air or the oxygen-rich air of cheapness and makes oxygenant, and speed of response is fast, and flow process is short, operating environment close friend, advantage such as construction investment is little.Be used to handle not the heavy non-ferrous metal sulphide ores and the difficult-treating gold mine of platinum group metal, developed into quite sophisticated modern metallurgical industry new technology in the world.
In recent years, the range of application of pressurization leaching-out technique progressively expands to the platinum group metal sulfide.Chinese patent application: publication number CN1417356A has reported a kind of method of extracting platinum metals and copper, nickel, cobalt from low-grade Pt-Pd sulfide flotation concentrate.At first adopt pressure oxidation acidleach or fiery wet method combined pre-treatment flotation concentrate, leach heavy non-ferrous metal copper, nickel, cobalt, leach the platinum metals with the pressure cyanide method then.Platinum, palladium leaching yield can reach Pt96%, Pd99%, and copper, nickel, cobalt leaching yield also reach more than 99%.Its advantage is reflected in the cupric sulfide nickel flotation concentrate of pressure oxidation acidleach technical finesse platiniferous palladium, can make sulphide ores all destroyed is vitriol, at high temperature then most of pyrrhosiderite and the hematite precipitation of forming of iron, help from pickling liquor, reclaiming copper and mickel, also help from leached mud, after follow-up pressure cyanide is handled, extracting platinum and palladium.But, make the application of this Technology be subjected to the influence of factors such as environment protection and operator safety protection, thereby increased the technology cost owing to use huge malicious prussiate.
Summary of the invention
The object of the invention provides a kind of technological operation is simply extracted platinum and palladium and weighed coloured valuable metal copper, nickel, cobalt from the sulphide ores of platinum group metal method.
The method of the extraction platinum from the sulphide ores of platinum group metal of foregoing invention purpose and palladium and heavy coloured valuable metal copper, nickel, cobalt comprises following several steps successively:
1. the digestion mineral aggregate pressurizes: mineral aggregate and digestion liquid were put into 50～250 ℃ of digestions of closed reactor 0.5～20 hour, and digestion liquid is selected from H
2SO
4, NaOH, water any, the weight ratio of this step mineral aggregate and this step digestion liquid is solid: liquid=1: 2～6 are filled with air, oxygen-rich air or pure oxygen gas in the reactor;
2. the solid slag that 1. obtains of 50～100 ℃ of digestion steps is 0.5～8 hour, and the digestion liquid that digestion is selected for use is HCl, H
2SO
4At least a, the weight ratio of the digestion liquid that the solid slag that 1. above-mentioned steps obtains and this step digestion are selected for use is solid: liquid=1: 2～6;
3. with the mixture solid slag that 2. the digestion step obtains under 40～100 ℃ of temperature of digestion liquid and oxygenant 0.5～8 hour, this step digestion liquid was selected from HCl, H
2SO
4, HCl and H
2SO
4Mixture, HCl and HNO
3Any of mixture, above-mentioned oxygenant is selected from H
2O
2, HClO
3, HClO, HNO
3, NaClO
3, NaClO, Na
2S
2O
8, chlorinated lime, Cl
2The solid-liquid weight ratio of the solid slag that 2. any of gas, above-mentioned steps obtain and this step digestion liquid is for solid: liquid=1: 2～6;
4. platinum, palladium during the leach liquor that 3. obtains with traditional enrichment, method of refining extraction step contains.
Behind each technical parameter, can further improve the productive rate of extract in the preferred technique scheme.
To step 1., its H
2SO
4, NaOH concentration be preferably 1wt%～40wt%, the gaseous tension in the reactor is preferably 1～5MPa.
To step 3., oxygenant H
2O
2, HClO
3, HClO, HNO
3, NaClO
3, NaClO, Na
2S
2O
8, chlorinated lime, Cl
2The consumption of gas is preferably 0.5%～50% of solid slag weight that 2. step obtain.
To step oxygenant H 3.
2O
2, HClO
3, HClO, HNO
3, NaClO
3, NaClO, Na
2S
2O
8, in this step digestion liquid of the selected direct adding of chlorinated lime, oxygenant Cl
2In the digestion liquid of selected continuous this step of directly feeding of gas.
Further preferred steps consolidating in 1.: liquid=1: 3～5, step consolidating in 2.: liquid=1: 3～5, step consolidating in 3.: liquid=1: 3～5.
Can also preferred steps H in 1.
2SO
4Concentration is 15wt%～25wt%, and gaseous tension is 1～3MPa oxygen in the reactor, and digestion temperature is 180～200 ℃, and the digestion time is 8～15 hours.
Also can preferred steps H in 2.
2SO
4Concentration is 20wt%～50wt%, and digestion temperature is 80～100 ℃, 4～6 hours digestion time.
Can also preferred steps described digestion temperature in 3. be 80～100 ℃, 4～6 hours described digestion time.
Chlorinated lime in the technique scheme is meant Ca (ClO)
2CaCl
22H
2O.
In the technique scheme, step HCl and H 2.
2SO
4In the mixture, 0＜HCl concentration＜60wt%, 0＜H
2SO
4Concentration＜60wt%; Step HCl and H 3.
2SO
4In the mixture, 0＜HCl concentration＜100wt%, 0＜H
2SO
4Concentration＜100wt%; Step HCl and HNO 3.
3In the mixture, 0＜HCl concentration＜100wt%, 0＜HNO
3Concentration＜100wt%.
State on the implementation in the technical scheme process, because the oxygen in the step gas 1. participates in reaction process, so the gaseous tension in the reactor selects to remain a certain steady state value selected in appropriate technical solution gaseous tension parameter area.Help reaction like this and normally carry out, guarantee to obtain higher leaching yield, realize that simultaneously rapid extraction reclaims heavy non-ferrous metal copper, nickel, the cobalt in the mineral aggregate.
In the technique scheme, 1. step contains the leach liquor of copper, nickel, cobalt respective compound with the 2. equal output of step, available traditional method, and as solvent extraction, copper, nickel, cobalt in the difference Separation and Recovery digestion liquid.
In the technique scheme, step oxygenant 3. can with digestion liquid synchronized mixes, also can be in temperature-rise period substep add in the digestion liquid.
In the technique scheme, step traditional enrichment, method of refining 4. can be substitution method, ion exchange method, solvent extration or chromatograph absorption method.Substitution method can use metal replacements such as powdery, sheet or blocky Zn, Mg, Al, Fe or Cu to obtain platinum, palladium.Ion exchange method comprises various anionite-exchange resin absorption and obtains platinum, palladium.Solvent extration mainly is to obtain the platinum palladium with quaternary ammonium salt, tertiary amine, TRPO or TBP extraction.Use traditional method of refining output platinum metals product again through the platinum metals enriched substance that above-mentioned any enrichment obtains.
The present invention is to the pressure oxidation acidleach pre-treatment slag of platinum group metal sulfide, adopt atmospheric pressure oxidation acidleach platinum metals, saved the high-temperature cyanidation process, thereby the potential hazard of having avoided using huge malicious prussiate and environment and operator being existed, and shortened technical process, adopt normal pressure leaching-out technique means greatly to reduce and implement Financial cost of the present invention.Process program technology maturation of the present invention, simple to operate, the equipment used maintenance and management is easy.Adopt the present invention to handle the platinum group metal sulfide, platinum, palladium recovery rate can reach Pt＞94%, Pd＞96%, and copper, nickel, cobalt leaching yield reach more than 98%.Invention has the adaptable characteristics of material.
The present invention can be used for handling the Cu, Ni and Co sulfide content below 15%, the low-grade platinum metals sulphide ores of platinum metals content below 0.05%, be more suitable for handling its flotation concentrate, simultaneously also be suitable for handling other material of platinum group metal, for example platinum group burning ore deposit, the anode sludge, smelting intermediates etc.The present invention all has no special requirements to employed each chemical reagent of reactant gases that comprises, above commercial reagent of technical pure or gas can satisfy the technical scheme requirement.
Embodiment
Select for use the low-grade Pt-Pd sulphide ores through the flotation concentrate of floating and enriching as the technical solution of the present invention process object.This kind flotation concentrate sulfur-bearing is about 15%, belongs to sulphide ores, and the Pt+Pd grade can reach 70～100g/t, but Cu and Ni grade are respectively＜5%, SiO
2Content is 26%, and MgO content is near 20%.The platinum palladium flotation concentrate material that experiment is handled is the toner powder darkly, and the sample quartering analytical results is listed in table 1.
Table 1 platinum palladium flotation concentrate multielement chemical analysis results (Pt, Pd:g/t, other element: wt%)
| Element | Pt | Pd | Cu | Fe | Ni | Co | S | SiO 2 | CaO | MgO |
| Content | 32.14 | 50.31 | 3.77 | 15.32 | 3.55 | 0.24 | 14.15 | 25.97 | 2.89 | 18.56 |
Embodiment 1
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 20%H
2SO
4Aqueous solution, the furnishing pulpous state is put into reactor.Mineral aggregate and H
2SO
4Aqueous solution solid-liquid weight ratio is 1: 4.Be heated to 200 ℃ of temperature, reacted 10 hours, gas is oxygen in the preceding reactor of heating, and the gaseous tension in the reaction process control reactor is constant to be 2.0MPa.
2. the acid leaching residue that 1. step is obtained under normal pressure with the H of weight percent concentration 20%
2SO
4Aqueous solution, solid-liquid weight ratio are 1: 4, and temperature of reaction is 90 ℃, 6 hours reaction times.
3. the acid leaching residue that 2. step is obtained is put into reactor, sizes mixing with weight percent concentration 50%HCl aqueous solution, and the solid-liquid weight ratio is 1: 5, is heated to 90 ℃ of temperature, adds weight and be the NaClO of the acid leaching residue 10% that 2. step obtain
3, normal pressure reacted 6 hours down.
4. the leach liquor zinc dust precipitation that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing.
Platinum, palladium leaching yield are respectively: Pt97.07%, Pd98.90%.
Copper, nickel, cobalt leaching yield are respectively: Cu98.0%, Ni98.6%, Co98.5%.
Embodiment 2
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 15%H
2SO
4Aqueous solution, the furnishing pulpous state is put into reactor.Mineral aggregate and H
2SO
4Aqueous solution solid-liquid weight ratio is 1: 5.Be heated to 200 ℃ of temperature, reacted 15 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 3.0MPa.
2. the acid leaching residue that 1. step is obtained under normal pressure with the HCl aqueous solution of concentration expressed in percentage by weight 20%, the solid-liquid weight ratio is 1: 4, temperature of reaction is 50 ℃, 4 hours reaction times;
3. the acid leaching residue that 2. step is obtained is put into reactor, mixes with weight percent concentration 100%HCl solution and sizes mixing, and the solid-liquid weight ratio is 1: 5, is heated to 40 ℃ of temperature, adds weight and be the H of the acid leaching residue 40% that 2. step obtain
2O
2, normal pressure reacted 4 hours down.
4. the leach liquor zinc dust precipitation that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing.
Platinum, palladium leaching yield are respectively: Pt94.55%, Pd97.30%.
Copper, nickel, cobalt leaching yield are respectively: Cu99.0%, Ni99.0%, Co98.5%.
Embodiment 3
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 25%H
2SO
4Aqueous solution, the furnishing pulpous state is put into reactor.Flotation concentrate and H
2SO
4Aqueous solution solid-liquid weight ratio is 1: 3.Be heated to 180 ℃ of temperature, reacted 8 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 2.5MPa.
2. the acid leaching residue that 1. step is obtained under normal pressure is 20% HCl+H with concentration expressed in percentage by weight
2SO
4Aqueous solution, solid-liquid weight ratio are 1: 4, and temperature of reaction is 80 ℃, react 4 hours.
3. the acid leaching residue that 2. step is obtained is put into reactor, mixes with weight percent concentration 50%HCl solution and sizes mixing, and the solid-liquid weight ratio is 1: 4, is heated to 80 ℃ of temperature, adds weight percent concentration 15%NaClO, and normal pressure reacted 8 hours down.
4. the leach liquor zinc dust precipitation that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing.
Platinum, palladium leaching yield are respectively: Pt94.0%, Pd96.5%.
Copper, nickel, cobalt leaching yield are respectively: Cu99.0%, Ni99.0%, Co98.5%.
Embodiment 4
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 15%NaOH aqueous solution, the furnishing pulpous state is put into reactor.Flotation concentrate and NaOH aqueous solution solid-liquid weight ratio are 1: 5.Be heated to 200 ℃ of temperature, reacted 20 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 2.5MPa.
2. the leached mud that 1. step is obtained under normal pressure is 50% H with concentration expressed in percentage by weight
2SO
4Aqueous solution, solid-liquid weight ratio are 1: 4, and temperature of reaction is 70 ℃, react 6 hours.
3. the leached mud that 2. step is obtained is put into reactor, with weight percent concentration 50%HCl+H
2SO
4Solution mixes sizes mixing, and the solid-liquid weight ratio is 1: 4, is heated to 80 ℃ of temperature, adds weight and be the HClO of the acid leaching residue 50% that 2. step obtain
3, normal pressure reacted 8 hours down.
4. the leach liquor zinc dust precipitation that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing.
Platinum, palladium leaching yield are respectively: Pt95.7%, Pd96.0%.
Copper, nickel, cobalt leaching yield are respectively: Cu98.6%, Ni99.0%, Co99.2%.
Embodiment 5
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 20%H
2SO
4Aqueous solution, the furnishing pulpous state is put into reactor.Flotation concentrate and H
2SO
4Aqueous solution solid-liquid weight ratio is 1: 4.Be heated to 200 ℃ of temperature, reacted 10 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 2.0MPa.
2. the acid leaching residue that 1. step is obtained under normal pressure with the H of weight percent concentration 30%
2SO
4Aqueous solution, solid-liquid weight ratio are 1: 4, and temperature of reaction is 90 ℃, 6 hours reaction times.
3. the acid leaching residue that 2. step is obtained is put into reactor, sizes mixing with weight percent concentration 100%HCl aqueous solution, and the solid-liquid weight ratio is 1: 5, is heated to 90 ℃ of temperature, feeds Cl in digestion liquid continuously
2, take out sample with sampler in the reaction process and send chemical analysis, with platinum metals leaching yield＞90% as reaction end.
4. the leach liquor zinc dust precipitation that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing.
Platinum, palladium leaching yield are respectively: Pt95.2%, Pd97.5%.
Copper, nickel, cobalt leaching yield are respectively: Cu98.0%, Ni98.6%, Co98.5%.
Embodiment 6
1. getting mineral aggregate 500 grams in the table 1, is that 2000 water that restrain mix with weight, and the furnishing pulpous state is put into reactor.Be heated to 200 ℃ of temperature, reacted 18 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 3.0MPa.
2. the leached mud that 1. step is obtained under normal pressure is 20% HCl aqueous solution with concentration expressed in percentage by weight, and the solid-liquid weight ratio is 1: 4, and temperature of reaction is 100 ℃, reacts 4 hours.
3. the leached mud that 2. step is obtained is put into reactor, with weight percent concentration 50%HCl+HNO
3Solution mixes sizes mixing, and the solid-liquid weight ratio is 1: 4, is heated to 60 ℃ of temperature, and normal pressure reacted 4 hours down.
4. the leach liquor solvent extraction that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing with zinc dust precipitation from strip liquor.
Platinum, palladium leaching yield are respectively: Pt95.4%, Pd96.2%.
Copper, nickel, cobalt leaching yield are respectively: Cu98.6%, Ni99.0%, Co99.2%.
Embodiment 7
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 20%H
2SO
4Aqueous solution, the furnishing pulpous state is put into reactor.Flotation concentrate and H
2SO
4Aqueous solution solid-liquid weight ratio is 1: 4.Be heated to 200 ℃ of temperature, reacted 10 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 2.0MPa.
2. the acid leaching residue that 1. step is obtained under normal pressure with the H of weight percent concentration 30%
2SO
4Aqueous solution, solid-liquid weight ratio are 1: 4, and temperature of reaction is 90 ℃, 6 hours reaction times.
3. the leached mud that 2. step is obtained is put into reactor, with weight percent concentration 50%H
2SO
4Solution mixes sizes mixing, and the solid-liquid weight ratio is 1: 6, is heated to 100 ℃ of temperature, adds weight and be the NaClO of the acid leaching residue 10% that 2. step obtain
3, normal pressure reacted 8 hours down.
4. the leach liquor zinc dust precipitation that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing.
Platinum, palladium leaching yield are respectively: Pt94.5%, Pd95.2%.
Copper, nickel, cobalt leaching yield are respectively: Cu98.6%, Ni99.0%, Co99.2%.
Embodiment 8
1. get mineral aggregate 500 grams in the table 1, with weight percent concentration 15%NaOH aqueous solution, the furnishing pulpous state is put into reactor.Flotation concentrate and NaOH aqueous solution solid-liquid weight ratio are 1: 5.Be heated to 200 ℃ of temperature, reacted 20 hours, gas is oxygen in the preceding reactor of heating, and the constant system pressure of reaction process is 2.5MPa.
2. the leached mud that 1. step is obtained under normal pressure is 50% HCl aqueous solution with concentration expressed in percentage by weight, and the solid-liquid weight ratio is 1: 4, and temperature of reaction is 70 ℃, reacts 6 hours.
3. the leached mud that 2. step is obtained is put into reactor, with weight percent concentration 95%HCl+HNO
3Solution mixes sizes mixing, and the solid-liquid weight ratio is 1: 4, is heated to 100 ℃ of temperature, and normal pressure reacted 2 hours down.
4. the leach liquor solvent extraction that 3. step is obtained obtains separating the purification platinum metals with traditional technology again behind the concentration of precious metal thing with zinc dust precipitation from strip liquor.
Platinum, palladium leaching yield are respectively: Pt96.4%, Pd97.0%.
Copper, nickel, cobalt leaching yield are respectively: Cu98.6%, Ni99.0%, Co99.2%.
Claims (8)
1. method of extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal comprises following process steps successively:
1. the digestion mineral aggregate pressurizes: mineral aggregate and digestion liquid were put into 50～250 ℃ of digestions of closed reactor 0.5 ~ 20 hour, and described digestion liquid is selected from H
2SO
4, NaOH, water any, the weight ratio of described mineral aggregate and described digestion liquid is solid: liquid=1: 2～6 are filled with air, oxygen-rich air or pure oxygen gas in the described reactor;
2. the solid slag that 1. obtains of 50～100 ℃ of digestion steps is 0.5～8 hour, and the digestion liquid that described digestion is selected for use is HCl, H
2SO
4Or its mixture, the weight ratio of the digestion liquid that solid slag that 1. described step obtains and described digestion are selected for use is solid: liquid=1: 2～6;
3. with the mixture solid slag that 2. the digestion step obtains under 40～100 ℃ of temperature of digestion liquid and oxygenant 0.5～8 hour, described digestion liquid was selected from HCl, H
2SO
4, HCl and H
2SO
4Mixture, HCl and HNO
3Any of mixture, described oxygenant is selected from H
2O
2, HClO
3, HClO, HNO
3, NaClO
3, NaClO, Na
2S
2O
8, chlorinated lime, Cl
2Any of gas, the solid-liquid weight ratio of solid slag that 2. described step obtains and described digestion liquid is for solid: liquid=1: 2～6;
4. platinum, the palladium in the leach liquor that 3. obtains with traditional enrichment, method of refining extraction step.
2. the method for extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal as claimed in claim 1 is characterized in that the 1. described H of step
2SO
4, NaOH concentration be 1wt%～40wt%, the gaseous tension in the described reactor is 1～5MPa.
3. the method for extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal as claimed in claim 1 is characterized in that the 3. described oxygenant H of step
2O
2, HClO
3, HClO, HNO
3, NaClO
3, NaClO, Na
2S
2O
8, chlorinated lime, Cl
2The consumption of gas is 0.5%～50% of the solid slag weight that 2. obtains of described step.
4. the method for extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal as claimed in claim 3 is characterized in that the 3. described oxygenant H of step
2O
2, HClO
3, HClO, HNO
3, NaClO
3, NaClO, Na
2S
2O
8, chlorinated lime is introduced directly in the described digestion liquid described oxygenant Cl
2Gas is directly fed in the described digestion liquid continuously.
5. as the described method of from the sulphide ores of platinum group metal, extracting platinum, palladium, copper, nickel, cobalt of one of claim 1 to 4, it is characterized in that step is 1. described solid: liquid=1: 3～5, step is 2. described solid: liquid=1: 3～5, step are 3. described solid: liquid=1: 3～5.
6. the method for extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal as claimed in claim 5 is characterized in that the 1. described H of step
2SO
4Concentration is 15wt%～25wt%, and gaseous tension is 1～3MPa oxygen in the described reactor, and described digestion temperature is 180～200 ℃, and the described digestion time is 8～15 hours.
7. the method for extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal as claimed in claim 5 is characterized in that the 2. described H of step
2SO
4Concentration is 20wt%～50wt%, and described digestion temperature is 80～100 ℃, 4～6 hours described digestion time.
8. the method for extracting platinum, palladium, copper, nickel, cobalt from the sulphide ores of platinum group metal as claimed in claim 5 is characterized in that the 3. described digestion temperature of step is 80～100 ℃, 4～6 hours described digestion time.
Priority Applications (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CNB2004100401018A CN1328398C (en) | 2004-06-26 | 2004-06-26 | Method for extracting platinum-palladium and base metal from platinum metal sulphide ore |
| ZA200505141A ZA200505141B (en) | 2004-06-26 | 2005-06-24 | Hydrometallurgical leaching method for extracting platinum, palladium, copper and nickel from the sulfide flotation concentrates containing platinum metals |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CNB2004100401018A CN1328398C (en) | 2004-06-26 | 2004-06-26 | Method for extracting platinum-palladium and base metal from platinum metal sulphide ore |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN1594608A CN1594608A (en) | 2005-03-16 |
| CN1328398C true CN1328398C (en) | 2007-07-25 |
Family
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| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CNB2004100401018A Expired - Lifetime CN1328398C (en) | 2004-06-26 | 2004-06-26 | Method for extracting platinum-palladium and base metal from platinum metal sulphide ore |
Country Status (2)
| Country | Link |
|---|---|
| CN (1) | CN1328398C (en) |
| ZA (1) | ZA200505141B (en) |
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| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1304612C (en) * | 2005-09-01 | 2007-03-14 | 徐致钢 | Process for extracting platinum metals from ore containing platinum metal |
| CN101509077B (en) * | 2009-02-19 | 2010-08-25 | 昆明贵金属研究所 | Method for extracting platinum, palladium, rhodium from automotive catalyst of ore phase reconstruction |
| JP2014507564A (en) * | 2011-02-03 | 2014-03-27 | ウェスタン プラチナム リミテッド | Method for refining platinum group metal concentrates |
| CN103436711B (en) * | 2013-08-22 | 2014-10-29 | 中南大学 | Method for enriching gold in gold cyanide sludge |
| CN103572066A (en) * | 2013-11-11 | 2014-02-12 | 广州有色金属研究院 | Method for enriching platinum family elements from platinum family concentrate |
| CN104263958B (en) * | 2014-08-30 | 2016-04-20 | 广东省工业技术研究院(广州有色金属研究院) | A kind of method being separated Cu, Ni and Co and platinum family element from platinum family concentrate |
| CN106244820A (en) * | 2016-08-29 | 2016-12-21 | 金川集团股份有限公司 | A kind of method of chlorine pressure rapid solution complexity rare precious metal concentrate |
| CN107475512B (en) * | 2017-08-30 | 2018-10-16 | 江西铜业股份有限公司 | A kind of method of comprehensive exploitation low-grade Pt-Pd concentrate |
| CN113802008A (en) * | 2021-09-16 | 2021-12-17 | 兰州大学 | Method for treating waste liquid containing platinum group noble metal |
Citations (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1061044A (en) * | 1990-10-29 | 1992-05-13 | 中国有色金属工业总公司昆明贵金属研究所 | From copper anode mud, reclaim Au, Pt, Pd and tellurium |
| CN1158904A (en) * | 1996-09-16 | 1997-09-10 | 昆明贵金属研究所 | Recovering method of platinum and lead in anode sludge from electrolytic hydrogen peroxide solution production |
| RU2167212C2 (en) * | 1999-06-28 | 2001-05-20 | Институт химии и технологии редких элементов и минерального сырья им.И.В.Тананаева Кольского научного центра РАН | Method of processing material based on carbon and containing noble metal |
| CN1417356A (en) * | 2001-11-07 | 2003-05-14 | 昆明贵金属研究所 | Extraction of platinum family metals and Cu, Ni and Co from sulfide ore or floated concentrate of platinum family metals |
| WO2004022795A1 (en) * | 2002-09-06 | 2004-03-18 | Placer Dome Technical Services Limited | Process for recovering platinum group metals from material containing base metals |
-
2004
- 2004-06-26 CN CNB2004100401018A patent/CN1328398C/en not_active Expired - Lifetime
-
2005
- 2005-06-24 ZA ZA200505141A patent/ZA200505141B/en unknown
Patent Citations (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1061044A (en) * | 1990-10-29 | 1992-05-13 | 中国有色金属工业总公司昆明贵金属研究所 | From copper anode mud, reclaim Au, Pt, Pd and tellurium |
| CN1158904A (en) * | 1996-09-16 | 1997-09-10 | 昆明贵金属研究所 | Recovering method of platinum and lead in anode sludge from electrolytic hydrogen peroxide solution production |
| RU2167212C2 (en) * | 1999-06-28 | 2001-05-20 | Институт химии и технологии редких элементов и минерального сырья им.И.В.Тананаева Кольского научного центра РАН | Method of processing material based on carbon and containing noble metal |
| CN1417356A (en) * | 2001-11-07 | 2003-05-14 | 昆明贵金属研究所 | Extraction of platinum family metals and Cu, Ni and Co from sulfide ore or floated concentrate of platinum family metals |
| WO2004022795A1 (en) * | 2002-09-06 | 2004-03-18 | Placer Dome Technical Services Limited | Process for recovering platinum group metals from material containing base metals |
Non-Patent Citations (1)
| Title |
|---|
| 加压湿法冶金处理含铂族金属铜镍硫化矿的应用及研究进展 黄昆等，稀有金属，第27卷第6期 2003 * |
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|---|---|
| ZA200505141B (en) | 2007-01-31 |
| CN1594608A (en) | 2005-03-16 |
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Application publication date: 20050316 Assignee: Kunming Guiyan New Material Technology Co., Ltd. Assignor: Kunming Institute of Precious Metals Contract record no.: X2019530000002 Denomination of invention: Method for extracting platinum-palladium and base metal from platinum metal sulphide ore Granted publication date: 20070725 License type: Common License Record date: 20190929 |
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