CN114438318A - Method for starting zinc hydrometallurgy - Google Patents
Method for starting zinc hydrometallurgy Download PDFInfo
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- CN114438318A CN114438318A CN202111652236.XA CN202111652236A CN114438318A CN 114438318 A CN114438318 A CN 114438318A CN 202111652236 A CN202111652236 A CN 202111652236A CN 114438318 A CN114438318 A CN 114438318A
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- Prior art keywords
- leaching
- zinc
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- sulfuric acid
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- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 142
- 239000011701 zinc Substances 0.000 title claims abstract description 131
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 129
- 238000000034 method Methods 0.000 title claims abstract description 120
- 238000009854 hydrometallurgy Methods 0.000 title claims abstract description 21
- 238000002386 leaching Methods 0.000 claims abstract description 169
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 113
- 239000007788 liquid Substances 0.000 claims abstract description 71
- 238000002156 mixing Methods 0.000 claims abstract description 71
- 230000008569 process Effects 0.000 claims abstract description 71
- 238000001556 precipitation Methods 0.000 claims abstract description 64
- 239000006228 supernatant Substances 0.000 claims abstract description 62
- 229910052742 iron Inorganic materials 0.000 claims abstract description 43
- 239000003792 electrolyte Substances 0.000 claims abstract description 38
- 229910052738 indium Inorganic materials 0.000 claims abstract description 35
- APFVFJFRJDLVQX-UHFFFAOYSA-N indium atom Chemical compound [In] APFVFJFRJDLVQX-UHFFFAOYSA-N 0.000 claims abstract description 35
- 239000002253 acid Substances 0.000 claims abstract description 32
- 239000002699 waste material Substances 0.000 claims abstract description 31
- 229910052760 oxygen Inorganic materials 0.000 claims abstract description 30
- 239000002893 slag Substances 0.000 claims abstract description 30
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims abstract description 29
- 239000001301 oxygen Substances 0.000 claims abstract description 29
- 230000009467 reduction Effects 0.000 claims abstract description 29
- 230000007935 neutral effect Effects 0.000 claims abstract description 22
- 229910052595 hematite Inorganic materials 0.000 claims abstract description 20
- 239000011019 hematite Substances 0.000 claims abstract description 20
- LIKBJVNGSGBSGK-UHFFFAOYSA-N iron(3+);oxygen(2-) Chemical compound [O-2].[O-2].[O-2].[Fe+3].[Fe+3] LIKBJVNGSGBSGK-UHFFFAOYSA-N 0.000 claims abstract description 19
- 229910052802 copper Inorganic materials 0.000 claims abstract description 17
- 239000010949 copper Substances 0.000 claims abstract description 17
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 12
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 138
- 239000008394 flocculating agent Substances 0.000 claims description 81
- RAHZWNYVWXNFOC-UHFFFAOYSA-N Sulphur dioxide Chemical compound O=S=O RAHZWNYVWXNFOC-UHFFFAOYSA-N 0.000 claims description 40
- 230000002829 reductive effect Effects 0.000 claims description 39
- 230000008719 thickening Effects 0.000 claims description 34
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 claims description 27
- 229910000368 zinc sulfate Inorganic materials 0.000 claims description 25
- 239000012535 impurity Substances 0.000 claims description 24
- 229960001763 zinc sulfate Drugs 0.000 claims description 21
- 235000019738 Limestone Nutrition 0.000 claims description 18
- 239000006028 limestone Substances 0.000 claims description 18
- 150000002500 ions Chemical class 0.000 claims description 14
- 238000005868 electrolysis reaction Methods 0.000 claims description 13
- 239000002002 slurry Substances 0.000 claims description 12
- LWUVWAREOOAHDW-UHFFFAOYSA-N lead silver Chemical compound [Ag].[Pb] LWUVWAREOOAHDW-UHFFFAOYSA-N 0.000 claims description 8
- 229910052793 cadmium Inorganic materials 0.000 claims description 5
- SOLUNJPVPZJLOM-UHFFFAOYSA-N trizinc;distiborate Chemical compound [Zn+2].[Zn+2].[Zn+2].[O-][Sb]([O-])([O-])=O.[O-][Sb]([O-])([O-])=O SOLUNJPVPZJLOM-UHFFFAOYSA-N 0.000 claims description 3
- 238000004886 process control Methods 0.000 claims 1
- 238000006386 neutralization reaction Methods 0.000 abstract description 13
- 238000000605 extraction Methods 0.000 abstract description 2
- 230000009977 dual effect Effects 0.000 abstract 1
- 239000000243 solution Substances 0.000 description 152
- 235000010269 sulphur dioxide Nutrition 0.000 description 18
- 229910001308 Zinc ferrite Inorganic materials 0.000 description 16
- WGEATSXPYVGFCC-UHFFFAOYSA-N zinc ferrite Chemical compound O=[Zn].O=[Fe]O[Fe]=O WGEATSXPYVGFCC-UHFFFAOYSA-N 0.000 description 15
- 238000000926 separation method Methods 0.000 description 14
- 239000002562 thickening agent Substances 0.000 description 12
- 238000006243 chemical reaction Methods 0.000 description 11
- 238000000746 purification Methods 0.000 description 11
- 238000005352 clarification Methods 0.000 description 9
- 230000000694 effects Effects 0.000 description 9
- 239000007787 solid Substances 0.000 description 9
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 description 8
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 description 8
- 150000004965 peroxy acids Chemical class 0.000 description 8
- 239000000084 colloidal system Substances 0.000 description 7
- 238000011084 recovery Methods 0.000 description 7
- 238000004062 sedimentation Methods 0.000 description 7
- 238000004519 manufacturing process Methods 0.000 description 6
- 229910052751 metal Inorganic materials 0.000 description 6
- 229920006395 saturated elastomer Polymers 0.000 description 6
- 229910052709 silver Inorganic materials 0.000 description 6
- 239000004332 silver Substances 0.000 description 6
- 239000011686 zinc sulphate Substances 0.000 description 6
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 description 5
- 238000003723 Smelting Methods 0.000 description 5
- 230000009286 beneficial effect Effects 0.000 description 5
- 239000003795 chemical substances by application Substances 0.000 description 5
- 238000004090 dissolution Methods 0.000 description 5
- 238000005189 flocculation Methods 0.000 description 5
- 238000006460 hydrolysis reaction Methods 0.000 description 5
- 239000000843 powder Substances 0.000 description 5
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 4
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 description 4
- 235000011941 Tilia x europaea Nutrition 0.000 description 4
- 230000002411 adverse Effects 0.000 description 4
- 229910052787 antimony Inorganic materials 0.000 description 4
- 229910052785 arsenic Inorganic materials 0.000 description 4
- 229910001431 copper ion Inorganic materials 0.000 description 4
- 230000016615 flocculation Effects 0.000 description 4
- 239000001257 hydrogen Substances 0.000 description 4
- 229910052739 hydrogen Inorganic materials 0.000 description 4
- 230000007062 hydrolysis Effects 0.000 description 4
- 239000004571 lime Substances 0.000 description 4
- 239000002184 metal Substances 0.000 description 4
- 230000003472 neutralizing effect Effects 0.000 description 4
- 239000002245 particle Substances 0.000 description 4
- 230000001376 precipitating effect Effects 0.000 description 4
- 239000002994 raw material Substances 0.000 description 4
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 4
- 239000011787 zinc oxide Substances 0.000 description 4
- 229910052984 zinc sulfide Inorganic materials 0.000 description 4
- UFHFLCQGNIYNRP-UHFFFAOYSA-N Hydrogen Chemical compound [H][H] UFHFLCQGNIYNRP-UHFFFAOYSA-N 0.000 description 3
- 239000005083 Zinc sulfide Substances 0.000 description 3
- 239000012141 concentrate Substances 0.000 description 3
- 238000000151 deposition Methods 0.000 description 3
- 230000008021 deposition Effects 0.000 description 3
- 238000005516 engineering process Methods 0.000 description 3
- 230000000670 limiting effect Effects 0.000 description 3
- 230000036284 oxygen consumption Effects 0.000 description 3
- 230000036961 partial effect Effects 0.000 description 3
- UGZADUVQMDAIAO-UHFFFAOYSA-L zinc hydroxide Chemical compound [OH-].[OH-].[Zn+2] UGZADUVQMDAIAO-UHFFFAOYSA-L 0.000 description 3
- 229940007718 zinc hydroxide Drugs 0.000 description 3
- 229910021511 zinc hydroxide Inorganic materials 0.000 description 3
- DRDVZXDWVBGGMH-UHFFFAOYSA-N zinc;sulfide Chemical compound [S-2].[Zn+2] DRDVZXDWVBGGMH-UHFFFAOYSA-N 0.000 description 3
- 229910001316 Ag alloy Inorganic materials 0.000 description 2
- 241000196324 Embryophyta Species 0.000 description 2
- 241000209094 Oryza Species 0.000 description 2
- 235000007164 Oryza sativa Nutrition 0.000 description 2
- PTFCDOFLOPIGGS-UHFFFAOYSA-N Zinc dication Chemical compound [Zn+2] PTFCDOFLOPIGGS-UHFFFAOYSA-N 0.000 description 2
- 230000009471 action Effects 0.000 description 2
- 229910052782 aluminium Inorganic materials 0.000 description 2
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 description 2
- 150000001462 antimony Chemical class 0.000 description 2
- 238000003490 calendering Methods 0.000 description 2
- 238000002425 crystallisation Methods 0.000 description 2
- 230000008025 crystallization Effects 0.000 description 2
- 238000010586 diagram Methods 0.000 description 2
- 238000004070 electrodeposition Methods 0.000 description 2
- 229910001449 indium ion Inorganic materials 0.000 description 2
- JEIPFZHSYJVQDO-UHFFFAOYSA-N iron(III) oxide Inorganic materials O=[Fe]O[Fe]=O JEIPFZHSYJVQDO-UHFFFAOYSA-N 0.000 description 2
- 239000000463 material Substances 0.000 description 2
- 239000011259 mixed solution Substances 0.000 description 2
- 229910052759 nickel Inorganic materials 0.000 description 2
- 238000004064 recycling Methods 0.000 description 2
- 235000009566 rice Nutrition 0.000 description 2
- 235000009529 zinc sulphate Nutrition 0.000 description 2
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 230000004075 alteration Effects 0.000 description 1
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 description 1
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 1
- GCPXMJHSNVMWNM-UHFFFAOYSA-N arsenous acid Chemical compound O[As](O)O GCPXMJHSNVMWNM-UHFFFAOYSA-N 0.000 description 1
- 230000015556 catabolic process Effects 0.000 description 1
- 239000007795 chemical reaction product Substances 0.000 description 1
- 238000010924 continuous production Methods 0.000 description 1
- 238000000354 decomposition reaction Methods 0.000 description 1
- 238000006731 degradation reaction Methods 0.000 description 1
- 238000009792 diffusion process Methods 0.000 description 1
- 238000006073 displacement reaction Methods 0.000 description 1
- 239000008151 electrolyte solution Substances 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 230000002349 favourable effect Effects 0.000 description 1
- 235000003891 ferrous sulphate Nutrition 0.000 description 1
- 239000011790 ferrous sulphate Substances 0.000 description 1
- 239000007789 gas Substances 0.000 description 1
- 229910052598 goethite Inorganic materials 0.000 description 1
- 230000005484 gravity Effects 0.000 description 1
- XLYOFNOQVPJJNP-ZSJDYOACSA-N heavy water Substances [2H]O[2H] XLYOFNOQVPJJNP-ZSJDYOACSA-N 0.000 description 1
- 150000002431 hydrogen Chemical class 0.000 description 1
- 230000003301 hydrolyzing effect Effects 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 description 1
- 230000000640 hydroxylating effect Effects 0.000 description 1
- ZMFWDTJZHRDHNW-UHFFFAOYSA-N indium;trihydrate Chemical compound O.O.O.[In] ZMFWDTJZHRDHNW-UHFFFAOYSA-N 0.000 description 1
- KFZAUHNPPZCSCR-UHFFFAOYSA-N iron zinc Chemical compound [Fe].[Zn] KFZAUHNPPZCSCR-UHFFFAOYSA-N 0.000 description 1
- BAUYGSIQEAFULO-UHFFFAOYSA-L iron(2+) sulfate (anhydrous) Chemical compound [Fe+2].[O-]S([O-])(=O)=O BAUYGSIQEAFULO-UHFFFAOYSA-L 0.000 description 1
- 229910000359 iron(II) sulfate Inorganic materials 0.000 description 1
- 229910052935 jarosite Inorganic materials 0.000 description 1
- 239000007791 liquid phase Substances 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 229920002401 polyacrylamide Polymers 0.000 description 1
- 239000002244 precipitate Substances 0.000 description 1
- 239000000047 product Substances 0.000 description 1
- 238000009853 pyrometallurgy Methods 0.000 description 1
- 230000002441 reversible effect Effects 0.000 description 1
- 239000007790 solid phase Substances 0.000 description 1
- 238000001179 sorption measurement Methods 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 239000004291 sulphur dioxide Substances 0.000 description 1
- 235000011149 sulphuric acid Nutrition 0.000 description 1
- 239000001117 sulphuric acid Substances 0.000 description 1
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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
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
-
- 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
- C22B1/00—Preliminary treatment of ores or scrap
-
- 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
-
- 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
- C22B15/0084—Treating solutions
- C22B15/0089—Treating solutions by chemical methods
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
- C22B3/46—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes by substitution, e.g. by cementation
-
- 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
- C22B58/00—Obtaining gallium or indium
-
- C—CHEMISTRY; METALLURGY
- C25—ELECTROLYTIC OR ELECTROPHORETIC PROCESSES; APPARATUS THEREFOR
- C25C—PROCESSES FOR THE ELECTROLYTIC PRODUCTION, RECOVERY OR REFINING OF METALS; APPARATUS THEREFOR
- C25C1/00—Electrolytic production, recovery or refining of metals by electrolysis of solutions
- C25C1/16—Electrolytic production, recovery or refining of metals by electrolysis of solutions of zinc, cadmium or mercury
-
- 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
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- Metallurgy (AREA)
- Materials Engineering (AREA)
- Manufacturing & Machinery (AREA)
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- Geochemistry & Mineralogy (AREA)
- General Life Sciences & Earth Sciences (AREA)
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Abstract
The invention discloses a method for starting a zinc hydrometallurgy process, which comprises the following steps: mixing part of zinc calcine with dilute acid for leaching, and sending the mixed leaching bottom flow to SO2A reduction leaching system, wherein the reduction leaching supernatant is subjected to iron powder copper precipitation, pre-neutralization, neutralization indium precipitation and hematite iron precipitation respectively to obtain hematite slag and iron precipitation liquid, and the iron precipitation liquid returns to the mixing leaching; reduction leaching underflow flowAnd (5) completing the start-up of the slag treatment system by a high-acid leaching system. Mixing the other part of the zinc calcine with dilute acid, mixed leaching supernatant and oxygen for neutral leaching, and returning the neutral leaching underflow to the mixed leaching; and the neutral leaching supernatant is purified, purified and then electrolyzed to obtain zinc and waste electrolyte, and part of the waste electrolyte is returned to the mixed leaching and/or neutral leaching, so that the start-up of the main process system is completed. The parallel start-up mode of the main process and the slag treatment dual system has the advantages of short start-up time, quick zinc extraction of the system, high total leaching rate of zinc, low start-up cost and the like.
Description
Technical Field
The invention belongs to the field of non-ferrous metal smelting, and particularly relates to a method for starting a zinc hydrometallurgy process.
Background
The zinc smelting process comprises pyrometallurgy and wet zinc smelting, wherein the wet zinc smelting accounts for more than 85%; the zinc hydrometallurgy process mainly comprises a conventional leaching method, a hot acid leaching method and an oxygen pressure leaching method; according to the treatment process of the acid leaching slag, the method mainly comprises the processes of conventional leaching-volatilization kiln/fuming furnace, hot acid leaching-goethite method, hot acid leaching-jarosite method, hot acid leaching-hematite method and the like. The traditional zinc hydrometallurgy process can not realize the direct utilization of iron slag, the conventional zinc hydrometallurgy has the problems of low recovery rate, high energy consumption, difficult stacking and utilization of residues and the like caused by the fact that acid leaching slag needs to be treated by a fire method, and the problems of poor raw material adaptability, difficult utilization and difficult stacking of produced sulfur slag in the process of zinc hydrometallurgy are also obvious.
Different from most domestic zinc hydrometallurgy processes, acid leaching residues are subjected to reduction leaching, and zinc and iron enter a solution at high temperature in a reducing atmosphere; the iron in solution sinks as hematite in the slag under high temperature, high acid and high pressure conditions in the autoclave. At present, only the zinc plant in the rice island of autumn field of Japan runs.
Since the technology of the zinc plant in the rice island of autumn Japan is blocked all the time, SO is adopted for the acid leaching slag of zinc hydrometallurgy in China at present on the aspect of industrial application2The reduction leaching and hematite iron removal process technology is always in a blank state and needs further research.
Disclosure of Invention
The present invention is directed to solving, at least to some extent, one of the technical problems in the related art. Therefore, the invention aims to provide a method for starting a zinc hydrometallurgy process, and the method for starting the zinc hydrometallurgy process has the advantages of short starting time, quick zinc extraction of a system, high total leaching rate of zinc, low starting cost and the like.
The invention provides a method for starting a zinc hydrometallurgy process. According to the embodiment of the invention, the method for starting the zinc hydrometallurgy comprises the following steps:
(1) mixing and leaching a part of the zinc calcine with a first sulfuric acid solution to obtain a first mixed leaching solution;
(2) mixing the first mixed leaching solution with a first flocculating agent and then carrying out first thickening treatment so as to obtain a first supernatant containing zinc sulfate and a first thickened underflow;
(3) mixing the other part of the zinc calcine with a second sulfuric acid solution, the first supernatant containing zinc sulfate and oxygen for neutral leaching so as to obtain a second mixed leaching solution;
(4) mixing the second mixed leaching solution with a second flocculating agent, then carrying out second thickening treatment to obtain a second supernatant and a second thickened underflow, and returning the second thickened underflow to the step (1) for carrying out the mixed leaching;
(5) purifying the second supernatant to remove impurities so as to obtain purified liquid;
(6) mixing the purified liquid with a third sulfuric acid solution for electrolysis so as to obtain zinc and waste electrolyte, and returning part of the waste electrolyte to the step (1) and/or the step (3).
According to the method for starting the zinc hydrometallurgy, a part of zinc calcine is mixed and leached with a first sulfuric acid solution to obtain a first mixed leaching solution, and the first mixed leaching solution is mixed with a first flocculating agent and then subjected to first thickening treatment to obtain a first supernatant containing zinc sulfate and a first thickening underflow. Meanwhile, the other part of the zinc calcine is mixed with a second sulfuric acid solution, a first supernatant containing zinc sulfate and oxygen for neutral leaching to obtain a second mixed leachate, the second mixed leachate is mixed with a second flocculating agent for second thickening treatment to obtain a second supernatant and a second thickened underflow, and the second thickened underflow is returned to the mixed leaching, so that the mixed leaching and the neutral leaching can be started simultaneously, the start time is shortened, and the production efficiency is improved. And then purifying and removing impurities from the second supernatant to obtain purified liquid, mixing the purified liquid with a third sulfuric acid solution for electrolysis to obtain zinc and electrolytic waste liquid, and returning part of the waste electrolyte to be mixed and leached or neutral for leaching, so that the recycling of the waste electrolyte is realized, and the production cost is saved.
In addition, the method for starting the hydrometallurgy of the zinc according to the embodiment of the invention can also have the following additional technical characteristics:
in some embodiments of the invention, further comprising: (7) mixing another part of the waste electrolyte with the first concentrated underflow, sulfur dioxide and a fourth sulfuric acid solution for reduction leaching so as to obtain a reduced solution; (8) mixing the reduced liquid with a third flocculating agent and then carrying out third thickening treatment so as to obtain a third supernatant and a third thickened underflow; (9) mixing the third supernatant with iron powder for copper deposition to obtain a copper deposited solution; (10) mixing the copper-precipitation solution with first limestone ore pulp for preneutralization so as to obtain a neutralized solution; (11) mixing the neutralized liquid with second limestone ore pulp for indium precipitation so as to obtain an indium precipitation liquid; (12) mixing the liquid after indium precipitation with oxygen to precipitate iron so as to obtain hematite slag and liquid after iron precipitation, and supplying the liquid after iron precipitation to the step (1) for mixed leaching.
In some embodiments of the invention, further comprising: (13) mixing the third concentrated underflow with a fifth sulfuric acid solution for high-acid leaching so as to obtain a high-acid leaching solution; (14) and (4) mixing the high-acid leaching solution and a fourth flocculating agent for fourth thickening treatment so as to obtain a fourth supernatant and lead-silver slag, and returning the fourth supernatant to the step (7) for reduction leaching. Therefore, the metal lead and silver can be enriched and recovered, and the total leaching rate of zinc is improved.
In some embodiments of the present invention, in the step (1), the concentration of the first sulfuric acid solution is 150 to 180 g/L. Therefore, on one hand, the zinc in the zinc calcine can be ensured to be dissolved more completely, and on the other hand, the influence on neutral leaching caused by the over-high content of harmful impurities in the first supernatant can be avoided.
In some embodiments of the invention, in step (1), the solid-to-liquid ratio of the mixed leaching process is 1t: (3 to 6) m3. Thus, the sufficiency of the mixed leaching can be ensured.
In some embodiments of the invention, in the step (2), the mass concentration of the first flocculating agent is 0.5 to 2 per thousand, and the amount of the first flocculating agent added is 7 to 15mL based on 1L of the first mixed leaching solution. Therefore, the sedimentation of the solid is facilitated, and the solid-liquid separation speed of the first mixed leaching solution is increased.
In some embodiments of the present invention, in the step (3), the concentration of the second sulfuric acid solution is 150 to 180 g/L. Therefore, on one hand, the zinc in the zinc calcine can be dissolved in the solution as quickly and completely as possible, and the high leaching rate is obtained; on the other hand, a hydrolysis precipitation method can be used for removing partial impurities (such as iron, arsenic, antimony and the like) so as to reduce the burden of solution purification.
In some embodiments of the invention, in step (3), the ratio of the zinc calcine to the second sulfuric acid solution and the zinc sulfate-containing first supernatant is 1t (2-5) m3：(2～5)m3. Thus, the sufficiency of neutral leaching can be ensured.
In some embodiments of the invention, in the step (3), the amount of the oxygen is 4 to 8Nm based on 1t of the zinc calcine3. Therefore, partial impurities can be removed by utilizing a hydrolysis precipitation method to the maximum extent so as to reduce the burden of subsequent solution purification.
In some embodiments of the invention, in the step (4), the mass concentration of the second flocculating agent is 0.5 to 2 per thousand, and the amount of the second flocculating agent added is 7 to 15mL based on 1L of the second mixed leaching solution. Therefore, the sedimentation of the solid is facilitated, and the solid-liquid separation speed of the second mixed leaching solution is increased.
In some embodiments of the invention, in the step (5), the purifying and impurity removing process includes a first-stage purifying process, a second-stage purifying process and a third-stage purifying process, wherein the first-stage purifying process is to add zinc powder to remove Cu and Cd, the second-stage purifying process is to add zinc powder and antimonate to remove Co, the third-stage purifying process is to add zinc powder to remove the balance Cd, and the purified solution ensures that Cd is less than 1mg/L and Co is less than 0.5 mg/L. Therefore, the smooth proceeding of the electrolytic process is ensured, and the high-quality precipitated zinc is obtained.
In some embodiments of the present invention, in the step (6), the concentration of the third sulfuric acid solution is 150 to 180 g/L. Therefore, on one hand, the zinc sulfate can be prevented from being hydrolyzed into zinc hydroxide; on the other hand, current efficiency can be ensured.
In some embodiments of the invention, in the step (6), the volume ratio of the purified liquid to the third sulfuric acid solution is 1 (20-30). Therefore, on one hand, the zinc sulfate can be prevented from being hydrolyzed into zinc hydroxide; on the other hand, current efficiency can be ensured.
In some embodiments of the present invention, in the step (6), the current density of the electrolysis process is controlled to be 360-420A/m2The ratio of zinc to zinc is 3.0 to 3.5. Thereby, high-quality precipitated zinc can be obtained.
In some embodiments of the present invention, in the step (7), the concentration of the fourth sulfuric acid solution is 150 to 180 g/L. Therefore, the sedimentation of the solid is facilitated, and the solid-liquid separation speed of the second mixed leaching solution is increased.
In some embodiments of the invention, in step (7), the volume ratio of the spent electrolyte to the first concentrated underflow and the fourth sulfuric acid solution is 1 (2-5): 1. Thus, the zinc in the zinc ferrite is dissolved out and the reduction behavior of Fe (III) is facilitated under the action of sulfur dioxide.
In some embodiments of the invention, in step (7), the amount of sulfur dioxide used is 30 to 50L based on 1L of the first concentrated underflow. Thus, the zinc in the zinc ferrite is dissolved out and the reduction behavior of Fe (III) is facilitated under the action of sulfur dioxide.
In some embodiments of the invention, in the step (8), the mass concentration of the third flocculating agent is 0.5-2% per mill, and the amount of the third flocculating agent added is 10-15 mL based on 1L of the reduced solution. Therefore, the method is beneficial to the sedimentation of the solid and accelerates the speed of liquid-solid-liquid separation after reduction.
In some embodiments of the present invention, in the step (9), the iron powder is added in an amount of 2-5 g based on 1L of the third supernatant. Thus, the sufficiency of the replacement copper deposition can be ensured.
In some embodiments of the invention, in step (10), the pre-neutralized end-point acidity is 5 to 15 g/L. Therefore, the subsequent indium neutralization and precipitation are facilitated.
In some embodiments of the invention, in the step (11), the volume ratio of the neutralized liquid to the second limestone pulp is (20-30): 1, and the indium precipitation end point pH is 4.5-5.2. Therefore, the method is beneficial to ensuring the sufficiency of neutralizing and precipitating indium.
In some embodiments of the invention, in the step (12), the amount of the oxygen is 5-8L based on the 1L of the indium precipitation solution. Thus, the iron in the solution can be removed by maximally hydrolyzing and precipitating the ferrous sulfate in the solution.
In some embodiments of the present invention, in the step (13), the concentration of the fifth sulfuric acid solution is 150 to 180 g/L. Therefore, zinc ferrite, zinc sulfide and the like which are not dissolved in the third concentrated underflow can be dissolved, and the zinc leaching rate is improved.
In some embodiments of the invention, in step (13), the volume ratio of the third concentrated underflow to the fifth sulfuric acid solution is 1 (2-5). Therefore, zinc ferrite, zinc sulfide and the like which are not dissolved in the third concentrated underflow can be dissolved, and the zinc leaching rate is improved.
In some embodiments of the invention, in the step (14), the mass concentration of the fourth flocculating agent is 0.5 to 2 per thousand, and the amount of the fourth flocculating agent added is 5 to 15mL based on 1L of the peracid leaching solution. Therefore, the method is beneficial to the sedimentation of solids and accelerates the solid-liquid separation speed of the high-acid leachate.
Additional aspects and advantages of the invention will be set forth in part in the description which follows and, in part, will be obvious from the description, or may be learned by practice of the invention.
Drawings
The above and/or additional aspects and advantages of the present invention will become apparent and readily appreciated from the following description of the embodiments, taken in conjunction with the accompanying drawings of which:
FIG. 1 is a schematic flow diagram of a method for starting up a hydrometallurgical zinc process according to an embodiment of the present invention;
FIG. 2 is a schematic flow chart of a method for starting up a hydrometallurgical zinc process according to yet another embodiment of the present invention;
FIG. 3 is a schematic flow chart of a method for starting up a hydrometallurgical zinc process according to yet another embodiment of the present invention;
fig. 4 is a schematic process flow diagram for the start-up of the hydrometallurgical zinc process according to one embodiment of the present invention.
Detailed Description
Reference will now be made in detail to embodiments of the present invention, examples of which are illustrated in the accompanying drawings, wherein like or similar reference numerals refer to the same or similar elements or elements having the same or similar function throughout. The embodiments described below with reference to the drawings are illustrative and intended to be illustrative of the invention and are not to be construed as limiting the invention.
Furthermore, the terms "first", "second" and "first" are used for descriptive purposes only and are not to be construed as indicating or implying relative importance or implicitly indicating the number of technical features indicated. Thus, a feature defined as "first" or "second" may explicitly or implicitly include at least one such feature. In the description of the present invention, "a plurality" means at least two, e.g., two, three, etc., unless specifically limited otherwise.
The invention provides a method for starting a zinc hydrometallurgy process. Referring to fig. 1-3, the method includes, according to an embodiment of the invention:
s100: mixing and leaching part of zinc calcine and first sulfuric acid solution
In the step, a part of zinc calcine is mixed and leached with a first sulfuric acid solution, so that soluble zinc in the zinc calcine can react as follows to the maximum extent: ZnO + H2SO4＝ZnSO4+H2And introducing O into the solution to obtain a first mixed leaching solution containing zinc sulfate. Further, the concentration of the first sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, and more preferably 165-170 g/L. The inventor finds that the higher the concentration of the sulfuric acid is, the more beneficial the leaching of zinc is, but the too high acidity can cause a large amount of impurity elements such as iron to enter the process, and can corrode equipment, cause crystallization and block pipelines. In addition, the solid-liquid ratio in the mixed leaching process is 1t: (3 to 6) m3Preferably 1t: (3 to 5) m3More preferably 1t: (3 to 4) m3The inventor finds that if the solid-liquid ratio is too high, the viscosity of the ore pulp is increased, so that the mass transfer speed between solid and liquid phases is reduced, the leaching rate is reduced, and the subsequent solid-liquid separation is difficult; and if the solid-liquid ratio is too low, the load of leaching and solid-liquid separation equipment or the loss of a leaching agent is increased. Therefore, the solid-liquid ratio of the method can improve the leaching rate on one hand; on the other hand, the load of leaching and solid-liquid separation equipment or the loss increase of a leaching agent can be avoided.
S200: mixing the first mixed leaching solution and a first flocculating agent and then carrying out first thickening treatment
In the step, the first mixed leaching solution and the first flocculating agent are mixed and then subjected to first thickening treatment to achieve the aim of solid-liquid separation. Preferably the process is carried out in a thickener and the sedimentation is carried out by gravity of the solid particles themselves, the addition of a flocculating agent causing the fine suspended particles to agglomerate into larger particles, thus increasing the sedimentation rate and thus obtaining a first supernatant containing zinc sulphate and a first concentrate underflow. It should be noted that the specific type of the first flocculating agent can be selected by those skilled in the art according to actual needs, as long as the speed of concentrating and clarifying the first mixed solution is increased, for example, the first flocculating agent is polyacrylamide. Meanwhile, the specific type of the thickener can be selected by a person skilled in the art according to actual needs, as long as the function can be realized. Further, the mass concentration of the first flocculating agent is 0.5-2 per mill, and the amount of the first flocculating agent to be added is 7-15 mL, preferably 7-12 mL, and more preferably 10mL based on 1L of the first mixed leaching solution, and the inventor finds that if the concentration or the addition amount of the first flocculating agent is too low, the flocculation effect is not ideal, and thus the concentration and clarification time is increased; if the concentration of the first flocculating agent is too high or the addition amount of the first flocculating agent is too high, the flocculating agent is also a transparent colloid, when the addition amount is too large, the first mixed leaching solution becomes very viscous if a saturated state is achieved, and at the moment, the flocculating agent cannot be gathered together and precipitated in the first mixed leaching solution and can block equipment such as pipelines. Therefore, by adopting the concentration and the input amount of the first flocculating agent, on one hand, the time for concentration and clarification can be shortened; on the other hand, the blockage of the pipeline can be avoided.
S300: mixing the other part of the zinc calcine with a second sulfuric acid solution, a first supernatant containing zinc sulfate and oxygen for neutral leaching
In this step, most of the zinc oxide can be dissolved by leaching by mixing the other part of the zinc calcine with the second sulfuric acid solution, the first supernatant containing zinc sulfate. In the leaching process, besides the zinc enters the solution, the metal impurities are dissolved to different degrees and follow the zinc IAnd comes into solution. These impurities adversely affect the zinc electrodeposition process and therefore it is necessary to remove as much as possible the harmful impurities before electrodeposition is carried out. During leaching, partial impurities (such As Fe, As, Sb and the like) are removed by using a hydrolysis precipitation method As much As possible so As to reduce the burden of solution purification. The pH value of the leaching end point is controlled to be close to neutral by adjusting the adding amount of acid, so that Fe can be ensured3+In the form of Fe (OH)3Hydrolysis precipitation, the reaction taking place is:
fe when the pH is 5.2-5.43+Can be completely removed, but Fe2+Cannot be removed, for which purpose Fe must be added2+Is oxidized into Fe3+Can be removed, so that oxygen is required to be introduced in the process. As and Sb in solution may co-precipitate with Fe, Fe (OH)3Is a colloid in Fe (OH)3The flocculation process of colloid has strong adsorption capacity, at the moment, the hydroxide of As and Sb can be adsorbed and co-precipitated, and Fe (OH) is formed in the leaching solution3Colloid and As3+The following reactions occur: 4Fe (OH)3+H3AsO3＝Fe4O5(OH)5As↓+5H2O, Sb also react similarly. By neutral leaching, the second mixed leaching solution with better quality can be ensured, the leaching rate of zinc is improved, and the zinc content of the middle leaching residue is reduced.
Furthermore, the concentration of the second sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, and more preferably 165-170 g/L, and the inventor finds that if the concentration is too high, a large amount of impurities such as iron and the like enter the leachate, further difficulty is brought to the clarification of ore pulp, the quality of the zinc sulfate solution is reduced, the technical and economic indexes of wet zinc smelting are influenced, in addition, equipment is corroded, crystallization is caused, and pipelines are blocked; if the concentration is too low, the leaching speed is slow, and the leaching is insufficient, so that the leaching rate of zinc is reduced. Therefore, the quality of the zinc sulfate solution and the zinc leaching rate can be guaranteed by adopting the second sulfuric acid concentration. Meanwhile, the ratio of the zinc calcine to the second sulfuric acid solution to the first supernatant containing zinc sulfate is 1t (2-5) m3：(2～5)m3. The inventor finds that by adopting the ratio in the range of the application, on one hand, the sufficiency of neutral leaching can be ensured, and on the other hand, the larger load required by the subsequent thickening treatment can be avoided, and the loss of raw materials can be reduced. In addition, based on 1t of the zinc calcine, the using amount of the oxygen is 4-8 Nm3Preferably 5 to 8Nm3More preferably 5 to 6Nm3The inventors have found that if the amount of oxygen is too low, a large amount of Fe still remains in the solution2+The water is not oxidized and cannot be removed from the solution by a hydrolysis precipitation method, the burden of subsequent solution purification is further increased, and certain influence is caused on the product quality; if the amount of oxygen is too high, unnecessary waste of oxygen will be caused. Therefore, the oxygen consumption of the method can reduce the burden of subsequent solution purification; on the other hand, cost can be saved.
S400: mixing the second mixed leaching solution with a second flocculating agent, then carrying out second thickening treatment, and returning the second thickened underflow to the step S100 for mixed leaching
In the step, the purpose of solid-liquid separation is achieved by mixing the second mixed leaching solution and the second flocculating agent and then carrying out second thickening treatment. Preferably, the process is carried out in a thickener, a second supernatant and a second concentrated underflow are obtained after the thickening treatment, and the second concentrated underflow is returned to the step S100 for mixed leaching, so that the zinc oxide which is not dissolved in the neutral leaching process is further dissolved to the maximum extent, and the leaching rate of zinc is improved. It should be noted that the type of thickener and second flocculant is the same as described above and will not be described further herein. Further, the mass concentration of the second flocculating agent is 0.5-2 per thousand, and the amount of the second flocculating agent to be added is 7-15 mL, preferably 10-15 mL, and more preferably 10-13 mL based on 1L of the second mixed leaching solution, and the inventor finds that if the concentration or the addition amount of the second flocculating agent is too low, the flocculation effect is not ideal, so that the concentration and clarification time is increased; if the concentration of the second flocculating agent is too high or the addition amount of the second flocculating agent is too high, the flocculating agent is also a transparent colloid, when the addition amount is too large, the second mixed leaching solution becomes very viscous if a saturated state is reached, and at the moment, the flocculating agent cannot be gathered together and precipitated in the second mixed leaching solution and can block equipment such as pipelines. Therefore, by adopting the concentration and the input amount of the second flocculating agent, on one hand, the time for concentration and clarification can be shortened; on the other hand, the blockage of the pipeline can be avoided.
S500: purifying the second supernatant to remove impurities
In the step, the second supernatant obtained by neutral leaching contains a large amount of impurities harmful to electrolysis, and the purified liquid can be obtained by three-stage continuous purification of the second supernatant. The three-stage purification principle is that by utilizing the principle of metal electrode potential difference, zinc powder is used for replacing impurities such as Cu, Cd, Co, Ni and the like in the second supernatant, the specific process is that first-stage zinc powder is added for replacing and removing Cu and Cd, second-stage zinc powder and antimonate are added for removing Co, and third-stage purification is performed by adding a small amount of zinc powder for removing the balance Cd, so that the content of impurities such as Cu, Cd, Co, Ni and the like can be reduced to the range allowed by electrolysis, and high-quality precipitated zinc is obtained. Meanwhile, through the purification and impurity removal process, some valuable metal elements in the raw materials can be further enriched, so that the valuable metal elements can be recovered from the purification slag. The concentration of main impurity ions of the purified liquid meets the following requirements: cd is less than 1mg/L, and Co is less than 0.5 mg/L. It should be noted that the addition amounts of the zinc powder and the antimony salt in the three-stage purification process are not particularly limited, and those skilled in the art can select the zinc powder and the antimony salt according to actual needs.
S600: mixing the purified solution with a third sulfuric acid solution for electrolysis, and returning a part of the waste electrolyte to step S100 and/or step S300
In the step, the purified solution is continuously supplied to an electrolytic bath, a third sulfuric acid solution is added in a certain proportion and mixed to be used as an electrolyte, a lead-silver alloy plate is used as an anode, a pure calendering aluminum plate is used as a cathode, and when direct current is passed, positive ions and negative ions in the solution respectively start to move in opposite directions: positive ions move to the cathode and negative ions move to the anode. Metallic zinc is precipitated on the cathode, and the reaction is as follows: zn2++2e ═ Zn; oxygen evolution on the anode takes place as follows: 2OH--2e＝H2O+1/2O2×) totalThe reaction formula is as follows: ZnSO4+H2O＝Zn+H2SO4+1/2O2×) @. Along with the continuous process, the zinc content in the electrolyte is continuously reduced, and the sulfuric acid content is continuously increased. The electrolyte becomes waste electrolyte after electrolytic deposition, and part of the waste electrolyte returns to the step S100 and/or the step S300 to carry out mixed leaching and/or neutral leaching on the zinc calcine, thereby realizing the recycling of the waste electrolyte and greatly saving the production cost. It should be noted that the electrolytic cell, the cathode and the anode and the specific types can be selected by those skilled in the art according to actual needs, as long as the above functions are realized. Furthermore, the concentration of the third sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, and more preferably 165-170 g/L, and the inventor finds that if the concentration of the third sulfuric acid solution is too high, the acidity of the electrolyte is too high, the reverse dissolution of precipitated zinc is increased, the possibility of hydrogen precipitation on the cathode is increased, and the current efficiency is remarkably reduced; if the concentration is too low, the acidity of the electrolyte is too low, zinc sulfate is hydrolyzed to generate zinc hydroxide, cathode zinc is in a sponge state, and the conductivity of the electrolyte is reduced. From this, adopt the third sulfuric acid solution concentration of this application can avoid the cathodic zinc to be the reduction of sponge state and electrolyte conductivity when guaranteeing current efficiency. Meanwhile, the mixing volume ratio of the purified liquid to the third sulfuric acid solution is 1 (20-30), preferably 1 (20-28), and more preferably 1 (25-28). the inventor finds that if the mixing volume ratio is too high, the acidity of the electrolyte is too low; if the volume ratio is too low, the acidity of the electrolyte solution becomes too high. Therefore, the electrolyte can be ensured to have proper acidity by adopting the mixing volume ratio of the electrolyte. In addition, the current density of the electrolysis process is controlled to be 360-420A/m according to the precipitation condition2The ratio of zinc to zinc is 3.0 to 3.5. The inventor finds that if the current density is too high, the temperature of the electrolyte is increased, and the separated zinc is dissolved again; and if the current density is too low, hydrogen is favorably precipitated on the cathode, and the electric efficiency is reduced. Meanwhile, if the zinc ratio is too low, the concentration of zinc ions is high, so that the cell voltage is increased; if the ratio of zinc to zinc is too high, the concentration of zinc ions is reduced, the hydrogen is precipitated on the cathode,causing the precipitated zinc to be re-dissolved and reducing the current efficiency. Thus, with the current density and zinc ratio of the present application, current efficiency degradation and cell voltage increase can be avoided.
Further, referring to fig. 2, the start-up method of the zinc hydrometallurgy includes:
s700: mixing another part of the waste electrolyte with the first concentrated underflow, sulfur dioxide and fourth sulfuric acid solution for reduction leaching
In the step, a part of the waste electrolyte is mixed with the first concentrated underflow and the fourth sulfuric acid solution obtained in the step S200, and sulfur dioxide gas is introduced into the mixed solution, so that the first concentrated underflow is subjected to reduction leaching, and a reduced solution is obtained. Zinc ferrite is inevitably generated in the roasting process, the structure of the zinc ferrite has quite large stability, and the zinc ferrite is difficult to dissolve out only by adopting mixed leaching. The dissolution of zinc ferrite is carried out by the following three steps: 1) hydroxylating water molecules on the surface of the zinc ferrite solid to form complex molecules; 2) the complex molecule is dissociated and diffused on the surface of the zinc ferrite solid; 3) the reaction product is dissociated. In these three steps, Fe3+The transfer process of ions from the surface of the zinc ferrite solid particles to the bulk of the solution is a limiting link of the whole dissolution process. When the solution contains a large amount of Fe3+When the ion is generated, the transfer process is slow; when the solution contains a large amount of Fe2+During ion, the transfer process is faster, so that the treatment of zinc ferrite by adopting a sulfur dioxide reduction method can complete Fe while realizing high-efficiency decomposition of the zinc ferrite3+Ion direction to Fe2+And (4) ion conversion. Reducing and decomposing zinc ferrite, and enabling iron in the reduced solution to exist in the form of Fe (II), thereby not only improving the recovery rate of zinc, but also creating favorable conditions for the subsequent separation of zinc, indium and iron. Furthermore, the volume ratio of the waste electrolyte to the first concentrated underflow and the fourth sulfuric acid solution is 1 (2-5): 1, preferably 1 (3-4): 1, and more preferably 1:3: 1. The inventors have found that the use of a volume ratio in the range of the present application is advantageous for the dissolution of zinc from zinc ferrite and the reduction behaviour of Fe (iii) under the influence of sulphur dioxide. Meanwhile, based on 1L of the first dense underflow, the amount of sulfur dioxide is 30 to 50L, preferably 30 to E40L, more preferably 35-40L, and the inventor finds that if the using amount is too low, the leaching rate of zinc and the reduction rate of Fe (III) are low; if the dosage is too high, the dissolved sulfur dioxide in the solution tends to be saturated, so that the promotion effect of continuously increasing the introduction amount of the sulfur dioxide on reduction leaching is small, and the waste of resources is caused. Therefore, by adopting the sulfur dioxide dosage, on one hand, higher zinc dissolution rate and Fe (III) reduction rate can be ensured; on the other hand, cost can be saved. In addition, the sulfur dioxide reduction leaching is carried out at the temperature of 100-140 ℃ for reaction. The inventors have found that the use of a reaction temperature in the range of the present application facilitates diffusion of the substance and increases the leaching rate.
S800: mixing the reduced liquid with a third flocculating agent and then carrying out third thickening treatment
In the step, the reduced liquid and a third flocculating agent are mixed and then subjected to third thickening treatment to achieve the purpose of solid-liquid separation. Preferably, the process is carried out in a thickener, which gives a third supernatant and a third thickened underflow after thickening. It should be noted that the type of thickener and third flocculant are the same as described above and will not be described further herein. Further, the mass concentration of the third flocculating agent is 0.5-2 per mill, and the amount of the third flocculating agent to be added is 10-15 mL based on 1L of the reduced solution. The inventor finds that if the concentration of the third flocculating agent is too low or the addition amount is too low, the flocculation effect is not ideal, and the concentration and clarification time is increased; if the concentration of the third flocculating agent is too high or the addition amount of the third flocculating agent is too high, the flocculating agent is also a transparent colloid, when the addition amount is too large, the reduced liquid becomes very viscous if a saturated state is reached, and at the moment, the flocculating agent cannot be gathered together and precipitated in the reduced liquid, and can block equipment such as pipelines. Therefore, by adopting the concentration and the input amount of the third flocculating agent, on one hand, the time for concentration and clarification can be shortened; on the other hand, the blockage of the pipeline can be avoided.
S900: mixing the third supernatant with iron powder for copper precipitation
In the step, the third supernatant is mixed with iron powder, so that copper ions in the solution are replaced by the iron powder and dissolvedPrecipitating out in the solution, and changing iron powder into Fe2+Ions enter the solution, so that the adverse effect of the existence of copper ions in the solution on the subsequent electrolysis process is avoided. Further, the mass of the iron powder added into 1L of the third supernatant is 2-5 g, preferably 2-4 g, and more preferably 3-4 g, and the inventor finds that if the amount of the iron powder added is too small, a large amount of copper ions in the third supernatant cannot be precipitated and separated out, and the copper ions in the electrolyte can be separated out at the cathode in the subsequent electrolysis process, so that the hydrogen separation is accelerated, the current efficiency is reduced, the separated zinc is loose and black, and holes are formed in the serious zinc; and if the adding amount is too much, the waste of the iron powder is caused. Therefore, by adopting the iron powder input amount, on one hand, higher current efficiency and zinc precipitation quality can be ensured; on the other hand, cost can be saved.
S1000: mixing the copper-precipitated solution with the first limestone slurry for preneutralization
In the step, because the acidity of the solution after copper precipitation is too high, if the indium is directly neutralized and precipitated, the taste of the indium slag is reduced due to the excessive using amount of the neutralizing agent, and therefore, a pre-neutralization process is performed before the indium precipitation by neutralization. The process is a principle of reacting the first limestone slurry with sulfuric acid in the solution after copper precipitation, and the acidity of the solution after copper precipitation is reduced to a certain degree, so that the reduction of the slag amount in the subsequent indium precipitation process is facilitated. After pre-neutralization, a neutralized liquid can be obtained. Specifically, the first lime slurry is 20% by mass of lime slurry (limestone powder containing CaCO)3> 93 wt%). Furthermore, the acidity of the pre-neutralized end point is 5-15 g/L, preferably 8-15 g/L, and more preferably 8-12 g/L, and the inventor finds that if the acidity of the end point is too low, the consumption of the neutralizing agent in the subsequent process is too much, and the taste of the indium slag is reduced; if the end point acidity is too high, a small amount of indium is lost to the slag. Therefore, by adopting the pre-neutralization end point acidity of the method, on one hand, the taste and the recovery rate of the subsequent indium slag can be ensured to be higher; on the other hand, the dosage of the neutralizing agent can be saved.
S1100: mixing the neutralized liquid with a second limestone slurry for indium precipitation
In this step, the neutralized liquid is mixed with the second limestone slurryAdjusting the pH value of the neutralized liquid to 4.5-5.2, wherein at the moment, the indium ions in the neutralized liquid can be hydrolyzed and precipitated, so that the indium-precipitated liquid is obtained, and the chemical reaction equation is as follows: in3++3H2O＝In(OH)3+3H+. Specifically, the second lime slurry is 20% lime slurry (the limestone powder contains CaCO)3> 93 wt%). Furthermore, the volume ratio of the neutralized liquid to the second limestone slurry is (20-30): 1 is preferably (22-28): 1, and is more preferably (24-26): 1, and the inventor finds that if the volume ratio is too high, insufficient indium precipitation due to neutralization can be caused, a large amount of indium ions in the neutralized liquid can remain in the solution, cannot be enriched, and can have adverse effects on subsequent processes; and if the volume ratio is too low, the taste of the indium in the slag is low. Therefore, the volume ratio of the method can ensure that the indium slag is high in taste and recovery rate.
S1200: mixing the solution after indium precipitation with oxygen to precipitate iron, and supplying the solution after iron precipitation to the step (1) for mixed leaching
In the step, the indium-precipitated liquid is mixed with oxygen to precipitate iron, and the mixture reacts at the temperature of 170-200 ℃, wherein the chemical reaction equation is as follows: 2FeSO4+0.5O2+2H2O＝Fe2O3+2H2SO4And hematite slag and the liquid after iron precipitation are obtained, and the liquid after iron precipitation is supplied to S100 for mixed leaching, so that the recovery rate of zinc is improved. Preferably, the hematite iron removal process is carried out in a hematite iron removal reactor. It is noted that the specific type of hematite iron removal reactor may be selected by those skilled in the art as required by the application, provided that the above-described functions are achieved. Further, based on 1L of the solution after indium precipitation, the amount of oxygen is 5-8L, preferably 6-8L, and the inventor finds that if the amount of oxygen is too low, Fe in the solution after indium precipitation cannot be guaranteed2+Ions are sufficiently oxidized, hydrolyzed and precipitated, thereby having adverse effects on subsequent processes; if the oxygen consumption is too high, the dissolved oxygen in the solution tends to be saturated, so that the promotion effect of continuously increasing the oxygen input on the iron precipitation of the hematite is small, and the waste of resources is caused. Thus, the present application is adoptedThe oxygen consumption ensures that the iron precipitation is carried out more thoroughly; on the other hand, resource waste can be avoided. In addition, hematite iron precipitation is carried out at the temperature of 170-200 ℃. The inventors have found that the reaction temperature in the range of the present application is beneficial for increasing the speed of hematite iron precipitation.
Further, referring to fig. 3, the start-up method for zinc hydrometallurgy includes:
s1300: mixing the third concentrated underflow with a fifth sulfuric acid solution for high-acid leaching
In the step, zinc sulfide in the third concentrated underflow and zinc ferrite which is not decomposed in the reduction leaching process of sulfur dioxide can be dissolved by performing the peracid leaching on the third concentrated underflow and a fifth sulfuric acid solution to obtain a peracid leaching solution, so that the zinc-iron leaching rate of the third concentrated underflow is improved, the leaching residue amount is greatly reduced, and the chemical reaction equations are as follows: ZnS + H2SO4＝ZnSO4+H2S and ZnO. Fe2O3+4H2SO4＝ZnSO4+Fe(SO4)3+4H2And O. Further, the concentration of the fifth sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, and more preferably 165-170 g/L, and the inventors found that if the concentration is too high, a part of lead and silver in the third concentrated underflow may be leached, so that the enrichment degree of lead and silver is reduced; and if the concentration is too low, the rate of peracid leaching will be reduced. Therefore, the concentration of the fifth sulfuric acid solution can be used for improving the recovery rate of lead and silver while ensuring the high acid leaching rate. In addition, the volume ratio of the third concentrated underflow to the fifth sulfuric acid solution is 1 (2-5), preferably 1 (3-5), and more preferably 1 (3-4), and the inventor finds that if the volume ratio is too high, the high-acid leaching may be insufficient, so that the total leaching rate of zinc is reduced, and the amount of leaching residues is increased; if the volume ratio is too low, the load of the subsequent thickening treatment is increased. Therefore, by adopting the volume ratio of the method, on one hand, the total leaching rate of zinc can be improved; on the other hand, the increase of the load of the subsequent thickening treatment can be avoided.
S1400: mixing the high-acid leachate with a fourth flocculating agent for fourth thickening treatment, and returning a fourth supernatant to the step S700 for reduction leaching
In the step, the purpose of solid-liquid separation is achieved by carrying out fourth thickening treatment on the peracid leaching solution and a fourth flocculating agent. Preferably, the process is carried out in a thickener, and a fourth supernatant and lead-silver slag are obtained after the thickening treatment, so that the lead and the silver in the raw materials can be enriched and recovered, and the fourth supernatant is returned to the step S700 for reduction leaching, so that the total leaching rate of zinc is further improved. It should be noted that the type of thickener and the fourth flocculant are the same as described above and will not be described herein again. Further, the mass concentration of the fourth flocculating agent is 0.5-2 per thousand, and the amount of the fourth flocculating agent to be added is 5-15 mL, preferably 5-10 mL, and more preferably 6-8 mL based on 1L of the peracid leaching solution, and the inventor finds that if the concentration or the addition amount of the fourth flocculating agent is too low, the flocculation effect is not ideal, so that the concentration and clarification time is increased; if the concentration of the fourth flocculating agent is too high or the addition amount of the fourth flocculating agent is too high, the flocculating agent is also a transparent colloid, when the addition amount is too large, the high-acid leachate becomes viscous if a saturated state is reached, and at the moment, the flocculating agents cannot be gathered together and precipitated in the high-acid leachate and can block equipment such as pipelines and the like. Therefore, by adopting the concentration and the input amount of the fourth flocculating agent, on one hand, the time for concentration and clarification can be shortened; on the other hand, the blockage of the pipeline can be avoided.
The inventors have found that a first mixed leach solution is obtained by mixing a portion of the zinc calcine with a first sulphuric acid solution and leaching, and that a first concentrate treatment is performed after mixing the first mixed leach solution with a first flocculating agent to obtain a first supernatant containing zinc sulphate and a first concentrate underflow. Then sending the first concentrated bottom stream to a sulfur dioxide reduction leaching system to obtain a reduced liquid, mixing the reduced liquid with a third flocculating agent, and then carrying out third concentration treatment to obtain a third supernatant and a third concentrated bottom stream, wherein the third supernatant is subjected to iron powder copper precipitation, pre-neutralization, neutralization indium precipitation and hematite iron precipitation respectively to obtain hematite slag and an iron precipitation liquid, and the iron precipitation liquid is returned to mixed leaching; and the third concentrated underflow is subjected to high-acid leaching to obtain a high-acid leaching solution, the high-acid leaching solution and a fourth flocculating agent are mixed for fourth concentration treatment to obtain a fourth supernatant and lead-silver slag, so that the start of a slag treatment system is completed, and the lead and the silver can be enriched and recovered.
Simultaneously mixing the other part of the zinc calcine with a second sulfuric acid solution, a first supernatant containing zinc sulfate and oxygen for neutral leaching to obtain a second mixed leachate, mixing the second mixed leachate with a second flocculating agent, and then carrying out second thickening treatment to obtain a second supernatant and a second thickened underflow, wherein the second thickened underflow returns to the mixed leaching step; and the second supernatant is purified and then is electrolyzed to obtain zinc and waste electrolyte, and part of the waste electrolyte is returned to be mixed and leached and/or neutral to complete the start of the main process system, and the waste electrolyte can be recycled, so that the production cost is saved.
By adopting the main process and slag treatment double-system parallel start-up mode, the start-up time is greatly shortened, the production efficiency is improved, and the production cost is reduced. The invention also provides a new zinc hydrometallurgy system starting scheme which takes the new technology of the new processes of sulfur dioxide reduction leaching, indium precipitation by neutralization and iron precipitation by hematite as the core, thereby improving the total leaching rate and recovery rate of zinc and reducing the zinc content of leaching slag.
Examples
Step 1: mixing and leaching a part of zinc calcine with the zinc content of 55 wt% and a first sulfuric acid solution with the concentration of 165g/L to obtain a first mixed leaching solution, wherein the solid-to-liquid ratio of the zinc calcine to the first sulfuric acid solution is 1: 4;
step 2: mixing the first mixed leaching solution and a first flocculating agent (Eisen 6000S) in a thickener, and then carrying out first thickening treatment, wherein the amount of the first flocculating agent (with the mass concentration of 1 per mill) to be added is 10mL based on 1L of the first mixed leaching solution, and a first supernatant containing zinc sulfate and a first thickening underflow are obtained after thickening treatment;
and step 3: mixing another part of zinc calcine with a second sulfuric acid solution with a concentration of 165g/L, a first supernatant containing zinc sulfate and oxygen for neutral leaching, and mixing the zinc calcine with the second sulfuric acidThe ratio of the solution to the first supernatant containing zinc sulfate was 1t: 3m3：2m3The end-point pH of the neutral leaching is controlled at 5.0 and the amount of oxygen is 6Nm based on 1g of zinc calcine3Thereby obtaining a second mixed leaching solution;
and 4, step 4: mixing the second mixed leaching solution and a second flocculating agent (Edison 655S) in a thickener, and then carrying out second thickening treatment, wherein the amount of the second flocculating agent (with the mass concentration of 1 per mill) to be added is 13mL based on 1L of the second mixed leaching solution, a second supernatant and a second thickened underflow are obtained after thickening treatment, and the second thickened underflow is returned to the step 1 for mixed leaching;
and 5: and purifying and removing impurities from the second supernatant to obtain purified liquid, wherein the purified and removed impurities are obtained by using a continuous three-section zinc powder displacement precipitation method, and the concentration of main impurity ions of the purified liquid meets the following requirements: cd is less than 1mg/L, Co is less than 0.5 mg/L;
step 6: mixing the purified solution with a third sulfuric acid solution with the concentration of 165g/L in an electrolytic bath for electrolysis, taking a pure calendering aluminum plate as a cathode and a lead-silver alloy plate as an anode, wherein the mixing volume ratio of the purified solution to the third sulfuric acid solution is 1:27, and controlling the current density to be 360-420A/m according to the precipitation condition in the electrolysis process2The ratio of zinc to zinc is 3.0-3.5, so that zinc and waste electrolyte are obtained, the zinc content in the obtained precipitated zinc is 99.995 wt%, and a part of the waste electrolyte is returned to the step 1 and/or the step 3;
and 7: mixing another part of the waste electrolyte with the first concentrated underflow, sulfur dioxide and a fourth sulfuric acid solution with the concentration of 165g/L at 110 ℃ for reduction leaching to obtain a reduced solution, wherein the volume ratio of the waste electrolyte to the first concentrated underflow to the fourth sulfuric acid solution is 1:3:1, and the amount of the sulfur dioxide is 38L based on 1L of the first concentrated underflow;
and 8: mixing the reduced solution and a third flocculating agent (Edson 6000S) in a thickener, and then carrying out third thickening treatment to obtain a third supernatant and a third thickened underflow, wherein the amount of the third flocculating agent (with the mass concentration of 1 per mill) added is 10mL based on 1L of the reduced solution;
and step 9: mixing the third supernatant with iron powderPrecipitating copper to obtain copper-precipitated solution, adding iron powder (reduced iron powder, containing Fe) into 1L of the third supernatant0> 95 wt%) of 3 g;
step 10: mixing the copper-precipitated solution with 20% limestone slurry (limestone powder containing CaCO)3More than 93 wt%) and mixing them to make preneutralization, and the terminal acidity of preneutralization is 12g/L so as to obtain neutralized liquor;
step 11: mixing the neutralized solution with 20% limestone slurry (limestone powder containing CaCO)3More than 93 wt%) and mixing, precipitating indium, wherein the end point pH of the indium precipitation by neutralization is 4.8, so as to obtain a liquid after indium precipitation, and the volume ratio of the liquid after neutralization to limestone slurry is 25: 1;
step 12: supplying the indium precipitation solution and oxygen to a hematite iron removal reactor, and mixing and precipitating iron at 180 ℃ to obtain hematite slag and an iron precipitation solution, wherein the zinc content of the hematite slag is 58-60 wt%, and supplying the iron precipitation solution to the step 1 for mixed leaching, wherein the amount of the oxygen is 8L based on 1L of the indium precipitation solution;
step 13: mixing the third concentrated underflow with a fifth sulfuric acid solution with the concentration of 165g/L for peracid leaching to obtain a peracid leaching solution, wherein the volume ratio of the third concentrated underflow to the fifth sulfuric acid solution is 1: 3;
step 14: mixing the high-acid leachate with a fourth flocculating agent (Eisen 6000S) in a thickener to perform fourth thickening treatment, wherein the amount of the fourth flocculating agent (with the mass concentration of 1 per thousand) is 7mL based on 1L of the high-acid leachate, obtaining a fourth supernatant and lead-silver residues after thickening treatment, the zinc content of the lead-silver residues is 3-5 wt%, and returning the fourth supernatant to the step 7 for reduction leaching.
In the description herein, references to the description of the term "one embodiment," "some embodiments," "an example," "a specific example," or "some examples," etc., mean that a particular feature, structure, material, or characteristic described in connection with the embodiment or example is included in at least one embodiment or example of the invention. In this specification, the schematic representations of the terms used above are not necessarily intended to refer to the same embodiment or example. Furthermore, the particular features, structures, materials, or characteristics described may be combined in any suitable manner in any one or more embodiments or examples. Furthermore, various embodiments or examples and features of different embodiments or examples described in this specification can be combined and combined by one skilled in the art without contradiction.
Although embodiments of the present invention have been shown and described above, it is understood that the above embodiments are exemplary and should not be construed as limiting the present invention, and that variations, modifications, substitutions and alterations can be made to the above embodiments by those of ordinary skill in the art within the scope of the present invention.
Claims (10)
1. A method for starting a zinc hydrometallurgy system is characterized by comprising the following steps:
(1) mixing and leaching a part of the zinc calcine with a first sulfuric acid solution to obtain a first mixed leaching solution;
(2) mixing the first mixed leaching solution with a first flocculating agent and then carrying out first thickening treatment so as to obtain a first supernatant containing zinc sulfate and a first thickened underflow;
(3) mixing the other part of the zinc calcine with a second sulfuric acid solution, the first supernatant containing zinc sulfate and oxygen for neutral leaching so as to obtain a second mixed leaching solution;
(4) mixing the second mixed leaching solution with a second flocculating agent, then carrying out second thickening treatment to obtain a second supernatant and a second thickened underflow, and returning the second thickened underflow to the step (1) for carrying out the mixed leaching;
(5) purifying the second supernatant to remove impurities so as to obtain purified liquid;
(6) mixing the purified liquid with a third sulfuric acid solution for electrolysis so as to obtain zinc and waste electrolyte, and returning part of the waste electrolyte to the step (1) and/or the step (3).
2. The method of claim 1, further comprising:
(7) mixing another part of the waste electrolyte with the first concentrated underflow, sulfur dioxide and a fourth sulfuric acid solution for reduction leaching so as to obtain a reduced solution;
(8) mixing the reduced liquid with a third flocculating agent and then carrying out third thickening treatment so as to obtain a third supernatant and a third thickened underflow;
(9) mixing the third supernatant with iron powder to precipitate copper so as to obtain a solution after copper precipitation;
(10) mixing the copper-precipitation solution with first limestone ore pulp for preneutralization so as to obtain a neutralized solution;
(11) mixing the neutralized liquid with second limestone ore pulp for indium precipitation so as to obtain an indium precipitation liquid;
(12) mixing the liquid after indium precipitation with oxygen to precipitate iron so as to obtain hematite slag and liquid after iron precipitation, and supplying the liquid after iron precipitation to the step (1) for mixed leaching.
3. The method of claim 2, further comprising:
(13) mixing the third concentrated underflow with a fifth sulfuric acid solution for high-acid leaching so as to obtain a high-acid leaching solution;
(14) and (4) mixing the high-acid leachate and a fourth flocculating agent for fourth thickening treatment so as to obtain a fourth supernatant and lead silver slag, and returning the fourth supernatant to the step (7) for reduction leaching.
4. The method according to claim 1, wherein in the step (1), the concentration of the first sulfuric acid solution is 150-180 g/L;
optionally, in step (1), the solid-to-liquid ratio of the mixed leaching process is 1t: (3 to 6) m3。
5. The method according to claim 1, wherein in the step (2), the mass concentration of the first flocculating agent is 0.5-2 per thousand, and the amount of the first flocculating agent added is 7-15 mL based on 1L of the first mixed leaching solution;
optionally, in the step (3), the concentration of the second sulfuric acid solution is 150-180 g/L;
optionally, in the step (3), the ratio of the zinc calcine to the second sulfuric acid solution and the first supernatant containing zinc sulfate is 1t (2-5) m3：(2～5)m3；
Optionally, in the step (3), the amount of the oxygen is 4 to 8Nm based on 1t of the zinc calcine3；
Optionally, in the step (4), the mass concentration of the second flocculating agent is 0.5-2 per mill, and the amount of the second flocculating agent added is 7-15 mL based on 1L of the second mixed leaching solution.
6. The method as claimed in claim 1, wherein in the step (5), the purifying and impurity removing process comprises a first-stage purifying process, a second-stage purifying process and a third-stage purifying process, wherein the first-stage purifying process is to add zinc powder to remove Cu and Cd, the second-stage purifying process is to add zinc powder and antimonate to remove Co, the third-stage purifying process is to add zinc powder to remove the balance Cd, and the concentration of main impurity ions in the purified liquid meets the following requirements: cd is less than 1mg/L, Co is less than 0.5 mg/L;
optionally, in the step (6), the concentration of the third sulfuric acid solution is 150-180 g/L;
optionally, in the step (6), the mixing volume ratio of the purified liquid to the third sulfuric acid solution is 1 (20-30);
optionally, in the step (6), the electrolysis process control current density is 360-420A/m2The ratio of zinc to zinc is 3.0 to 3.5.
7. The method according to claim 2, wherein in the step (7), the concentration of the fourth sulfuric acid solution is 150 to 180 g/L;
optionally, in the step (7), the volume ratio of the waste electrolyte to the first concentrated underflow and the fourth sulfuric acid solution is 1 (2-5): 1;
optionally, in the step (7), the amount of the sulfur dioxide is 30-50L based on 1L of the first concentrated bottom flow.
8. The method according to claim 2, characterized in that in the step (8), the mass concentration of the third flocculating agent is 0.5-2 per thousand, and the amount of the third flocculating agent added is 10-15 mL based on 1L of the reduced solution;
optionally, in the step (9), the weight of the iron powder is added to be 2-5 g based on 1L of the third supernatant;
optionally, in the step (10), the pre-neutralized end point sulfuric acid content is 5-15 g/L.
9. The method according to claim 2, characterized in that in the step (11), the volume ratio of the neutralized liquid to the second limestone slurry is (20-30): 1, and the indium precipitation end point pH is 4.5-5.2;
optionally, in the step (12), the amount of the oxygen is 5-8L based on 1L of the indium precipitation solution.
10. The method according to claim 3, wherein in the step (13), the concentration of the fifth sulfuric acid solution is 150 to 180 g/L;
optionally, in the step (13), the volume ratio of the third concentrated bottom flow to the fifth sulfuric acid solution is 1 (2-5);
optionally, in the step (14), the mass concentration of the fourth flocculating agent is 0.5-2 per mill, and the amount of the fourth flocculating agent added is 5-15 mL based on 1L of the high-acid leaching solution.
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