CN114438318B - Zinc hydrometallurgy start-up method - Google Patents
Zinc hydrometallurgy start-up method Download PDFInfo
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- CN114438318B CN114438318B CN202111652236.XA CN202111652236A CN114438318B CN 114438318 B CN114438318 B CN 114438318B CN 202111652236 A CN202111652236 A CN 202111652236A CN 114438318 B CN114438318 B CN 114438318B
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- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 135
- 239000011701 zinc Substances 0.000 title claims abstract description 124
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 123
- 238000000034 method Methods 0.000 title claims abstract description 86
- 238000009854 hydrometallurgy Methods 0.000 title claims abstract description 25
- 238000002386 leaching Methods 0.000 claims abstract description 184
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 95
- 239000007788 liquid Substances 0.000 claims abstract description 86
- 238000002156 mixing Methods 0.000 claims abstract description 72
- 239000006228 supernatant Substances 0.000 claims abstract description 62
- 239000002253 acid Substances 0.000 claims abstract description 50
- 230000008569 process Effects 0.000 claims abstract description 50
- 238000001556 precipitation Methods 0.000 claims abstract description 44
- 239000003792 electrolyte Substances 0.000 claims abstract description 38
- 239000002893 slag Substances 0.000 claims abstract description 35
- 229910052742 iron Inorganic materials 0.000 claims abstract description 32
- 230000009467 reduction Effects 0.000 claims abstract description 30
- 229910052760 oxygen Inorganic materials 0.000 claims abstract description 29
- 239000002699 waste material Substances 0.000 claims abstract description 29
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims abstract description 28
- 239000001301 oxygen Substances 0.000 claims abstract description 28
- 229910052738 indium Inorganic materials 0.000 claims abstract description 22
- APFVFJFRJDLVQX-UHFFFAOYSA-N indium atom Chemical compound [In] APFVFJFRJDLVQX-UHFFFAOYSA-N 0.000 claims abstract description 22
- 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 16
- 239000010949 copper Substances 0.000 claims abstract description 16
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 11
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 138
- 230000008719 thickening Effects 0.000 claims description 48
- 230000002829 reductive effect Effects 0.000 claims description 41
- RAHZWNYVWXNFOC-UHFFFAOYSA-N Sulphur dioxide Chemical compound O=S=O RAHZWNYVWXNFOC-UHFFFAOYSA-N 0.000 claims description 40
- 239000008394 flocculating agent Substances 0.000 claims description 35
- 238000000746 purification Methods 0.000 claims description 22
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 claims description 22
- 229960001763 zinc sulfate Drugs 0.000 claims description 22
- 229910000368 zinc sulfate Inorganic materials 0.000 claims description 22
- 239000012535 impurity Substances 0.000 claims description 21
- 235000019738 Limestone Nutrition 0.000 claims description 18
- 239000012530 fluid Substances 0.000 claims description 18
- 239000006028 limestone Substances 0.000 claims description 18
- 238000005868 electrolysis reaction Methods 0.000 claims description 14
- 150000002500 ions Chemical class 0.000 claims description 14
- 239000002002 slurry Substances 0.000 claims description 14
- LWUVWAREOOAHDW-UHFFFAOYSA-N lead silver Chemical compound [Ag].[Pb] LWUVWAREOOAHDW-UHFFFAOYSA-N 0.000 claims description 10
- 150000001462 antimony Chemical class 0.000 claims description 4
- 229910052793 cadmium Inorganic materials 0.000 claims description 3
- 238000004886 process control Methods 0.000 claims description 2
- 238000006386 neutralization reaction Methods 0.000 abstract description 11
- 238000000605 extraction Methods 0.000 abstract description 2
- 239000000243 solution Substances 0.000 description 155
- WGEATSXPYVGFCC-UHFFFAOYSA-N zinc ferrite Chemical compound O=[Zn].O=[Fe]O[Fe]=O WGEATSXPYVGFCC-UHFFFAOYSA-N 0.000 description 16
- 229910001308 Zinc ferrite Inorganic materials 0.000 description 15
- 238000000926 separation method Methods 0.000 description 13
- 239000002562 thickening agent Substances 0.000 description 12
- 238000006243 chemical reaction Methods 0.000 description 11
- 239000007787 solid Substances 0.000 description 10
- 238000005352 clarification Methods 0.000 description 9
- 230000000694 effects Effects 0.000 description 9
- 239000000084 colloidal system Substances 0.000 description 8
- 238000004090 dissolution Methods 0.000 description 7
- 230000007062 hydrolysis Effects 0.000 description 7
- 238000006460 hydrolysis reaction Methods 0.000 description 7
- 238000011084 recovery Methods 0.000 description 7
- 238000004062 sedimentation Methods 0.000 description 7
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 description 6
- 238000004519 manufacturing process Methods 0.000 description 6
- 229910052751 metal Inorganic materials 0.000 description 6
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 description 5
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 description 5
- 238000003723 Smelting Methods 0.000 description 5
- 229910001431 copper ion Inorganic materials 0.000 description 5
- 238000000151 deposition Methods 0.000 description 5
- 230000008021 deposition Effects 0.000 description 5
- 238000005189 flocculation Methods 0.000 description 5
- 230000036284 oxygen consumption Effects 0.000 description 5
- 239000000843 powder Substances 0.000 description 5
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 4
- UFHFLCQGNIYNRP-UHFFFAOYSA-N Hydrogen Chemical compound [H][H] UFHFLCQGNIYNRP-UHFFFAOYSA-N 0.000 description 4
- 235000011941 Tilia x europaea Nutrition 0.000 description 4
- 230000002411 adverse Effects 0.000 description 4
- 230000000903 blocking effect Effects 0.000 description 4
- 239000003795 chemical substances by application Substances 0.000 description 4
- 238000010586 diagram Methods 0.000 description 4
- 230000016615 flocculation Effects 0.000 description 4
- 239000004571 lime Substances 0.000 description 4
- 239000002184 metal Substances 0.000 description 4
- 239000002245 particle Substances 0.000 description 4
- 239000002994 raw material Substances 0.000 description 4
- 229910052709 silver Inorganic materials 0.000 description 4
- 239000004332 silver Substances 0.000 description 4
- 239000005083 Zinc sulfide Substances 0.000 description 3
- 230000009471 action Effects 0.000 description 3
- 229910052787 antimony Inorganic materials 0.000 description 3
- 229910052785 arsenic Inorganic materials 0.000 description 3
- 239000001257 hydrogen Substances 0.000 description 3
- 229910052739 hydrogen Inorganic materials 0.000 description 3
- 230000000670 limiting effect Effects 0.000 description 3
- 230000003472 neutralizing effect Effects 0.000 description 3
- 239000002244 precipitate Substances 0.000 description 3
- 229920006395 saturated elastomer Polymers 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
- 239000011787 zinc oxide Substances 0.000 description 3
- 229910052984 zinc sulfide 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
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 2
- PTFCDOFLOPIGGS-UHFFFAOYSA-N Zinc dication Chemical compound [Zn+2] PTFCDOFLOPIGGS-UHFFFAOYSA-N 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
- 230000009286 beneficial effect Effects 0.000 description 2
- 238000005266 casting Methods 0.000 description 2
- 238000002425 crystallisation Methods 0.000 description 2
- 230000008025 crystallization Effects 0.000 description 2
- 230000007423 decrease Effects 0.000 description 2
- 238000009792 diffusion process Methods 0.000 description 2
- 238000005516 engineering process Methods 0.000 description 2
- 230000002349 favourable effect Effects 0.000 description 2
- 229910001449 indium ion Inorganic materials 0.000 description 2
- 239000000463 material Substances 0.000 description 2
- 239000011259 mixed solution Substances 0.000 description 2
- 238000004064 recycling Methods 0.000 description 2
- 229910052717 sulfur Inorganic materials 0.000 description 2
- 239000011593 sulfur Substances 0.000 description 2
- 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
- 230000005540 biological transmission Effects 0.000 description 1
- 239000007795 chemical reaction product Substances 0.000 description 1
- 238000000354 decomposition reaction Methods 0.000 description 1
- 238000006073 displacement reaction Methods 0.000 description 1
- 238000010494 dissociation reaction Methods 0.000 description 1
- 230000005593 dissociations Effects 0.000 description 1
- 238000004070 electrodeposition Methods 0.000 description 1
- 238000005265 energy consumption Methods 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
- 150000004679 hydroxides Chemical class 0.000 description 1
- 230000033444 hydroxylation Effects 0.000 description 1
- 238000005805 hydroxylation reaction Methods 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
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- 230000036961 partial effect Effects 0.000 description 1
- 229920002401 polyacrylamide Polymers 0.000 description 1
- 239000000047 product Substances 0.000 description 1
- 230000001698 pyrogenic effect Effects 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
- 239000002351 wastewater Substances 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 1
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- 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
Abstract
The invention discloses a method for starting zinc hydrometallurgy, which comprises the following steps: mixing and leaching a part of zinc calcine with dilute acid, and delivering the mixed leaching bottom flow to SO 2 The reduction leaching system is used for respectively carrying out copper precipitation, preneutralization, neutralization indium precipitation and hematite iron precipitation on the reduction leaching supernatant to obtain hematite slag and iron precipitation post-liquid, and returning the iron precipitation post-liquid to mixed leaching; and (5) sending the reduction leaching bottom stream to a high-acid leaching system to finish the starting of the slag treatment system. Mixing the other part of the zinc calcine with dilute acid, mixed leaching supernatant and oxygen to perform neutral leaching, and returning neutral leaching underflow to mixed leaching; and the neutral leaching supernatant is sent to be purified and then electrolyzed to obtain zinc and waste electrolyte, and part of the waste electrolyte is returned to mixed leaching and/or neutral leaching, thereby completing the starting of the main process system. The parallel driving mode of the main process and the slag treatment double systems has the advantages of short driving time, quick zinc extraction of the system, high total leaching rate of zinc, low driving cost and the like.
Description
Technical Field
The invention belongs to the field of nonferrous metal smelting, and particularly relates to a method for starting zinc hydrometallurgy.
Background
Zinc smelting processes are divided into pyrogenic and wet zinc smelting, wherein zinc smelting by wet method accounts for more than 85%; the wet zinc smelting process mainly comprises a conventional leaching method, a hot acid leaching method and an oxygen pressure leaching method; according to the treatment process of acid leaching slag, the acid leaching slag is mainly divided into conventional leaching-volatilizing kiln/fuming furnace, hot acid leaching-goethite method, hot acid leaching-jarosite method, hot acid leaching-hematite method and other processes. The traditional zinc hydrometallurgy process cannot realize the direct utilization of iron slag, and the problems of low recovery rate, high energy consumption, difficult stacking and difficult utilization of residues and the like caused by the fire treatment of acid leaching slag in the conventional zinc hydrometallurgy process are also solved, and the problems of poor raw material adaptability, difficult utilization of produced sulfur slag and difficult stacking in the oxygen pressure leaching zinc hydrometallurgy process are also outstanding.
Unlike the domestic zinc hydrometallurgy process, the acid leaching slag adopts reduction leaching, and zinc and iron enter solution at high temperature and under a reducing atmosphere; iron in solution sinks as hematite into the slag under high temperature, high acid and high pressure conditions in the autoclave.
In industrialized application, SO is adopted for wet zinc hydrometallurgy acid leaching residues at present in China 2 The technology of the reduction leaching and hematite iron removal process needs to be further studied.
Disclosure of Invention
The present invention aims to solve at least one of the technical problems in the related art to some extent. Therefore, one purpose of the invention is to provide a zinc hydrometallurgy starting method, and the zinc hydrometallurgy starting method 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 zinc hydrometallurgy. 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 zinc calcine and a first sulfuric acid solution so as to obtain a first mixed leaching solution;
(2) Mixing the first mixed leaching solution and a first flocculating agent, and then carrying out first thickening treatment so as to obtain a first supernatant fluid containing zinc sulfate and a first thickening bottom fluid;
(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 and a second flocculating agent, and then carrying out second thickening treatment so as to obtain a second supernatant fluid and a second thickening bottom flow, and returning the second thickening bottom flow to the step (1) for carrying out the mixed leaching;
(5) Purifying and impurity-removing the second supernatant to obtain a purified liquid;
(6) Mixing the purified solution with a third sulfuric acid solution for electrolysis to obtain zinc and a waste electrolyte, and returning a part of the waste electrolyte to step (1) and/or step (3).
According to the method for starting zinc hydrometallurgy, disclosed by the embodiment of the invention, the first mixed leaching solution is obtained by mixing and leaching a part of zinc calcine and the first sulfuric acid solution, and the first mixed leaching solution is mixed with the first flocculating agent and then subjected to first thickening treatment, so that a first supernatant fluid and a first thickening underflow containing zinc sulfate are obtained. And 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 leaching solution, the second mixed leaching solution is mixed with a second flocculating agent and then subjected to second thickening treatment to obtain a second supernatant and a second thickening bottom flow, and the second thickening bottom flow is returned to the mixed leaching, so that the mixed leaching and the neutral leaching can be started at the same time, the starting 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 electrolysis waste liquid, and returning part of the waste electrolyte to mixed leaching or neutral leaching, thereby realizing the recycling of the waste electrolyte and saving the production cost.
In addition, the method for starting the zinc hydrometallurgy 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 a further portion of the spent electrolyte with the first dense underflow, sulfur dioxide and a fourth sulfuric acid solution for reduction leaching to obtain a reduced liquor; (8) Mixing the reduced liquid with a third flocculant and then carrying out third thickening treatment so as to obtain a third supernatant and a third thickening bottom flow; (9) Mixing the third supernatant with iron powder for copper precipitation so as to obtain copper precipitation post-liquid; (10) Mixing the copper-precipitated liquid with first limestone ore pulp for preneutralization so as to obtain a neutralized liquid; (11) Mixing the neutralized solution with second limestone ore pulp to precipitate indium so as to obtain indium-precipitated solution; (12) Mixing the indium-precipitated liquid with oxygen to precipitate iron so as to obtain hematite slag and iron-precipitated liquid, and supplying the iron-precipitated liquid to the step (1) to perform mixed leaching.
In some embodiments of the invention, further comprising: (13) Mixing the third dense underflow with a fifth sulfuric acid solution for high acid leaching so as to obtain a high acid leaching solution; (14) Mixing the high acid leaching solution with a fourth flocculant 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 the 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 150-180 g/L. Thus, on the one hand, the zinc in the zinc calcine can be ensured to be more completely dissolved, and on the other hand, the influence of the excessive content of harmful impurities in the first supernatant liquid on neutral leaching can be avoided.
In some embodiments of the invention, in step (1), the mixed leaching process has a solids to liquid ratio of 1t: (3-6) m 3 . Thereby, the sufficiency of the mixed leaching can be ensured.
In some embodiments of the present invention, in the step (2), the mass concentration of the first flocculant is 0.5 to 2%o, and the amount of the first flocculant added is 7 to 15ml based on 1L of the first mixed leaching solution. Thus, the sedimentation of solids 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 150-180 g/L. Therefore, on one hand, zinc in the zinc calcine can be dissolved in the solution as quickly and completely as possible, and high leaching rate is obtained; on the other hand, part of impurities (such as iron, arsenic, antimony, etc.) can be removed by utilizing a hydrolysis precipitation method so as to lighten the burden of solution purification.
In some embodiments of the invention, in step (3), the zinc calcine is combined with the second sulfuric acid solution and the sulfur-containing componentThe ratio of the first supernatant liquid of zinc acid is 1t (2-5) m 3 ：（2~5）m 3 . Thereby, the sufficiency of neutral leaching can be ensured.
In some embodiments of the invention, in step (3), the amount of oxygen is 4-8 Nm based on 1t of the zinc calcine 3 . Thus, part of impurities can be removed by the hydrolysis precipitation method to the maximum extent, so that the burden of subsequent solution purification is reduced.
In some embodiments of the present invention, in the step (4), the mass concentration of the second flocculant is 0.5 to 2%o, and the amount of the second flocculant added is 7 to 15ml based on 1L of the second mixed leaching solution. Thus, the sedimentation of solids is facilitated, and the solid-liquid separation speed of the second mixed leaching solution is increased.
In some embodiments of the present invention, in step (5), the purifying and impurity removing process includes a first-stage purification, a second-stage purification, and a third-stage purification, wherein the first-stage purification is to add zinc powder to remove Cu and Cd, the second-stage purification is to add zinc powder and antimony salt to remove Co, the third-stage purification is to add zinc powder to remove the rest Cd, and the purified solution ensures Cd < 1mg/L and Co < 0.5mg/L. Thus, the smooth progress of the electrolysis process is ensured, and high-quality zinc precipitation is obtained.
In some embodiments of the present invention, in the step (6), the concentration of the third sulfuric acid solution is 150-180 g/L. Thus, on the one hand, the hydrolysis of zinc sulfate into zinc hydroxide can be avoided; on the other hand, the current efficiency can be ensured.
In some embodiments of the present invention, in step (6), the volume ratio of the purified solution to the third sulfuric acid solution is 1 (20-30). Thus, on the one hand, the hydrolysis of zinc sulfate into zinc hydroxide can be avoided; on the other hand, the current efficiency can be ensured.
In some embodiments of the invention, in step (6), the electrolysis process controls the current density to be 360-420A/m 2 The acid zinc ratio is 3.0-3.5. Thus, high quality zinc precipitation can be obtained.
In some embodiments of the present invention, in the step (7), the concentration of the fourth sulfuric acid solution is 150-180 g/L. Thus, the sedimentation of solids is facilitated, and the solid-liquid separation speed of the second mixed leaching solution is increased.
In some embodiments of the present invention, in step (7), the volume ratio of the waste electrolyte to the first dense underflow to the fourth sulfuric acid solution is 1 (2-5): 1. Thereby facilitating the dissolution of zinc in zinc ferrite and the reduction behavior of Fe (III) under the action of sulfur dioxide.
In some embodiments of the present invention, in step (7), the sulfur dioxide is used in an amount of 30 to 50L based on 1L of the first dense underflow. Thereby facilitating the dissolution of zinc in zinc ferrite and the reduction behavior of Fe (III) under the action of sulfur dioxide.
In some embodiments of the present invention, in the step (8), the mass concentration of the third flocculant is 0.5 to 2%o, and the amount of the third flocculant added is 10 to 15ml based on 1L of the reduced solution. Thus, the sedimentation of solids is facilitated, and the speed of liquid-solid-liquid separation after reduction is accelerated.
In some embodiments of the present invention, in step (9), the iron powder is required to be added in an amount of 2 to 5g based on 1L of the third supernatant. Thereby, the sufficiency of replacement copper deposition can be ensured.
In some embodiments of the present invention, in step (10), the preneutralization endpoint acidity is 5-15 g/L. Thereby being beneficial to subsequent neutralization and indium precipitation.
In some embodiments of the present invention, in the step (11), the volume ratio of the neutralized solution to the second limestone slurry is (20-30): 1, and the indium precipitation endpoint pH is 4.5-5.2. Therefore, the method is beneficial to ensuring the sufficiency of neutralization indium precipitation.
In some embodiments of the present invention, in step (12), the oxygen is used in an amount of 5 to 8L based on the 1L indium-precipitating solution. Thus, ferrous sulfate in the solution can be hydrolyzed and precipitated to the maximum extent, and iron in the solution can be removed.
In some embodiments of the present invention, in the step (13), the concentration of the fifth sulfuric acid solution is 150-180 g/L. Therefore, zinc ferrite, zinc sulfide and the like which are not dissolved in the third dense underflow can be dissolved, so that the zinc leaching rate is improved.
In some embodiments of the present invention, in step (13), the volume ratio of the third dense 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 dense underflow can be dissolved, so that the zinc leaching rate is improved.
In some embodiments of the present invention, in step (14), the mass concentration of the fourth flocculant is 0.5 to 2 per mill, and the amount of the fourth flocculant added is 5 to 15ml based on 1L of the high acid leaching solution. Thus, the sedimentation of solids is facilitated, and the solid-liquid separation speed of the high-acid leaching solution is accelerated.
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 foregoing and/or additional aspects and advantages of the invention will become apparent and may be better understood from the following description of embodiments taken in conjunction with the accompanying drawings in which:
FIG. 1 is a schematic flow diagram of a zinc hydrometallurgy start-up process according to an embodiment of the invention;
FIG. 2 is a schematic flow diagram of a method of operating a zinc hydrometallurgy in accordance with a further embodiment of the invention;
FIG. 3 is a schematic flow diagram of a zinc hydrometallurgy start-up process according to a further embodiment of the invention;
fig. 4 is a schematic diagram of a process flow of a zinc hydrometallurgy start-up according to an embodiment of the invention.
Detailed Description
Embodiments of the present invention are described in detail below, examples of which are illustrated in the accompanying drawings, wherein like or similar reference numerals refer to like or similar elements or elements having like or similar functions throughout. The embodiments described below by referring to the drawings are illustrative and intended to explain the present invention and should not be construed as limiting the invention.
Furthermore, the terms "first," "second," and the like, are used for descriptive purposes only and are not to be construed as indicating or implying a relative importance or implicitly indicating the number of technical features indicated. Thus, a feature defining "a first" or "a second" may explicitly or implicitly include at least one such feature. In the description of the present invention, the meaning of "plurality" means at least two, for example, two, three, etc., unless specifically defined otherwise.
The application provides a method for starting zinc hydrometallurgy. Referring to fig. 1-3, the method, according to an embodiment of the present application, includes:
s100: mixing and leaching a part of zinc calcine with a first sulfuric acid solution
In the step, part of the zinc calcine is mixed and leached with the first sulfuric acid solution, so that the soluble zinc in the zinc calcine is reacted to the maximum extent: znO+H 2 SO 4 = ZnSO 4 + H 2 O enters the solution, thereby obtaining a first mixed leaching solution containing zinc sulfate. Further, the concentration of the first sulfuric acid solution is 150 to 180g/L, preferably 160 to 180g/L, and more preferably 165 to 170g/L. The inventor finds that the higher the concentration of sulfuric acid is, the more favorable the leaching of zinc is, but the excessive acidity can lead a large amount of impurity elements such as iron and the like to enter the process, and simultaneously can corrode equipment, cause crystallization and block a pipeline. In addition, the solid-liquid ratio in the mixed leaching process is 1t: (3-6) m 3 Preferably 1t: (3-5) m 3 More preferably 1t: (3-4) m 3 The 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; if the solid-liquid ratio is too low, the load on the leaching and solid-liquid separation equipment or the loss of the leaching agent increases. Therefore, the leaching rate can be improved by adopting the solid-liquid ratio of the application; on the other hand, the increase of the load of leaching and solid-liquid separation equipment or the loss of leaching agent can be avoided.
S200: mixing the first mixed leaching solution and the first flocculating agent, and then carrying out first thickening treatment
In the step, the purpose of solid-liquid separation is achieved by mixing the first mixed leaching solution and the first flocculating agent and then performing first thickening treatment. Preferably, the process is carried out in a thickener, sedimentation is carried out by gravity of the solid particles themselves, and the addition of a flocculant can agglomerate the tiny suspended particles into larger particles, thereby increasing the sedimentation rate, and thereby obtaining a first supernatant containing zinc sulfate and a first dense underflow. It should be noted that, a person skilled in the art may select a specific type of the first flocculant according to actual needs, so long as the first mixed solution concentrating and clarifying speed can be increased, for example, the first flocculant is polyacrylamide. Meanwhile, a person skilled in the art can select the specific type of the thickener according to actual needs, so long as the functions can be realized. Further, the mass concentration of the first flocculant is 0.5 to 2 per mill, based on 1L of the first mixed leaching solution, the amount of the first flocculant to be added is 7 to 15mL, preferably 7 to 12mL, more preferably 10mL, and the inventor finds that if the concentration of the first flocculant is too low or the addition amount is too low, the flocculation effect is not ideal, so that the time for concentration and clarification is increased; if the concentration of the first flocculating agent is too high or the adding amount is too high, the first mixed leaching solution becomes very viscous if the adding amount is too large because the flocculating agent is also a transparent colloid, and at the moment, the flocculating agent in the first mixed leaching solution cannot be gathered together to precipitate, and can block equipment such as a pipeline and the like. Therefore, by adopting the concentration and the input amount of the first flocculating agent, the time for concentration and clarification can be shortened on one hand; on the other hand, the blocking of the pipeline can be avoided.
S300: mixing the other part of zinc calcine with second sulfuric acid solution, first supernatant containing zinc sulfate and oxygen for neutral leaching
In this step, the zinc oxide is mostly dissolved by leaching by mixing the other part of the zinc calcine with the second sulfuric acid solution, the first supernatant liquid containing zinc sulfate. In the leaching process, in addition to zinc going into solution, metallic impurities dissolve to varying degrees and go into solution with zinc. These impurities adversely affect the zinc electrodeposition process and therefore must be present before the electrodepositionHarmful impurities are removed as much as possible. In the leaching process, partial impurities (such As Fe, as, sb and the like) are removed by utilizing a hydrolysis precipitation method As much As possible so As to lighten the burden of solution purification. The pH value of the leaching end point is controlled to be close to neutral by adjusting the addition amount of acid, so that Fe can be obtained 3+ In the form of Fe (OH) 3 Hydrolysis and precipitation, the reaction is: fe (Fe) 3+ + 3H 2 O ↔ Fe(OH) 3 + 3H + . When ph=5.2 to 5.4, fe 3+ Can be completely removed, but Fe 2+ To this end, fe must be added 2+ Oxidation to Fe 3+ Can be removed, and thus oxygen is required to be introduced into the process. As and Sb in solution can be co-precipitated with Fe, fe (OH) 3 Is a colloid, in Fe (OH) 3 In the flocculation process of the colloid, the colloid has strong adsorption capacity, at the moment, the hydroxides of As and Sb can be adsorbed and coprecipitated, and Fe (OH) is formed in the leaching liquid 3 Colloid and As 3+ The following reactions occur: 4Fe (OH) 3 + H 3 AsO 3 = Fe 4 O 5 (OH) 5 As↓ + 5H 2 O, sb react similarly. By neutral leaching, the second mixed leaching solution with better quality can be ensured to be obtained, the leaching rate of zinc is improved, and the zinc content of the intermediate leaching slag is reduced.
Further, the concentration of the second sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, more preferably 165-170 g/L, and the inventor finds that if the concentration is too high, impurities such as iron and the like are caused to enter the leaching solution in a large amount, so that difficulty is brought to clarification of ore pulp, the quality of zinc sulfate solution is reduced, the technical and economic indexes of zinc hydrometallurgy are affected, and in addition, equipment is corroded, crystallization is caused, and a pipeline is 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 ensured by adopting the second sulfuric acid concentration. Meanwhile, the ratio of the zinc calcine to the second sulfuric acid solution to the first supernatant fluid containing zinc sulfate is 1t (2-5) m 3 ：（2~5）m 3 . The inventors found that, by adopting the ratio of the scope of the application, on one hand, the sufficiency of neutral leaching can be ensured, and on the other hand, the need for subsequent thickening treatment can be avoided The load is borne greatly, and the loss of raw materials is reduced. In addition, based on 1t of zinc calcine, the amount of oxygen used is 4-8 Nm 3 Preferably 5 to 8Nm 3 More preferably 5 to 6Nm 3 The inventors found that too low an amount of oxygen would result in a large amount of Fe still in the solution 2+ Is not oxidized, so that the waste water cannot be removed from the solution by a hydrolysis precipitation method, further increases the burden of subsequent solution purification, and can have a certain influence on the product quality; if the oxygen consumption is too high, unnecessary waste of oxygen is caused. Therefore, the burden of subsequent solution purification can be reduced by adopting the oxygen consumption of the application; on the other hand, the cost can be saved.
S400: mixing the second mixed leaching solution and the second flocculant, performing second thickening treatment, and returning second thickening 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 fluid and a second dense underflow are obtained after the thickening treatment, and the second dense underflow is returned to the step S100 for mixed leaching, so that zinc oxide which is not dissolved in the neutral leaching process is further dissolved out to the maximum extent, and the leaching rate of zinc is improved. It should be noted that the thickener and the second flocculant are the same as described above, and are not described here again. Further, the mass concentration of the second flocculant is 0.5 to 2 per mill, based on 1L of the second mixed leaching solution, the amount of the second flocculant to be added is 7 to 15mL, preferably 10 to 15mL, more preferably 10 to 13mL, and the inventor finds that if the concentration of the second flocculant is too low or the addition amount is too low, the flocculation effect is not ideal, so that the time for concentration and clarification is increased; if the concentration of the second flocculating agent is too high or the adding amount is too high, the second mixed leaching solution becomes very viscous if the adding amount is too large because the flocculating agent is also a transparent colloid, and at the moment, the flocculating agent in the second mixed leaching solution can not be gathered together to precipitate, and can also block equipment such as a pipeline. Therefore, by adopting the concentration and the input amount of the second flocculant, the time for concentration and clarification can be shortened on one hand; on the other hand, the blocking 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 carrying out three-stage continuous purification on the second supernatant. The three-stage purification principle is to replace Cu, cd, co, ni and other impurities in the second supernatant by zinc powder by utilizing the principle of potential difference of metal electrodes, and the specific process is that the first stage of zinc powder is added for replacing Cu and Cd, the second stage of zinc powder and antimony salt are added for Co removal, and the three-stage purification is added with a small amount of zinc powder for removing the rest Cd, so that the content of Cu, cd, co, ni and other impurities can be reduced to the allowable range of 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 purified slag. The main impurity ion concentration of the purified liquid meets the following requirements: cd is less than 1mg/L, co is less than 0.5mg/L. The amounts of zinc powder and antimony salt added in the above three-stage purification process are not particularly limited, and may be selected by those skilled in the art 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 liquid is continuously supplied to an electrolytic tank, a third sulfuric acid solution is added in a certain proportion for mixing to be used as an electrolyte, a lead-silver alloy plate is used as an anode, a pure casting aluminum plate is used as a cathode, and when direct current passes through, positive ions and negative ions in the solution respectively start to move in opposite directions: positive ions move towards the cathode and negative ions move towards the anode. The metal zinc is precipitated on the cathode, and the reaction is: zn (zinc) 2+ +2e=zn; oxygen is discharged from the anode, and the reaction is as follows: 2OH - - 2e = H 2 O + 1/2O 2 ∈, the general reaction formula is: znSO (ZnSO) 4 + H 2 O = Zn + H 2 SO 4 + 1/2O 2 And ≡. As the process continues, the zinc content in the electrolyte decreases,the sulfuric acid content is increasing. The electrolyte is subjected to electrolytic deposition to become waste electrolyte, 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 zinc calcine, so that the recycling of the waste electrolyte is realized, and the production cost is greatly saved. It should be noted that, those skilled in the art may select the electrolytic cell, the cathode and the anode and the specific type according to actual needs, so long as the above functions can be achieved. Further, the concentration of the third sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, 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 that hydrogen is precipitated on the cathode is increased, and thus 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, so that cathode zinc is in a sponge state, and the conductivity of the electrolyte is reduced. Therefore, the concentration of the third sulfuric acid solution can ensure the current efficiency and simultaneously avoid the cathode zinc in a sponge state and the reduction of the conductivity of the electrolyte. Meanwhile, the mixing volume ratio of the purified solution to the third sulfuric acid solution is 1 (20-30), preferably 1 (20-28), more preferably 1 (25-28), and the inventor finds that if the mixing volume ratio is too high, the acidity of the electrolyte is too low; if the mixing volume ratio is too low, the acidity of the electrolyte is too high. Thus, the adoption of the mixing volume ratio of the application can ensure that the electrolyte has proper acidity. In addition, the current density in the electrolysis process is controlled to be 360-420A/m according to precipitation conditions 2 The acid zinc ratio is 3.0-3.5. The inventor finds that if the current density is too high, the temperature of the electrolyte is increased to further precipitate zinc for dissolution; if the current density is too low, the hydrogen is easy to be separated out on the cathode, and the electrical efficiency is reduced. Meanwhile, if the acid zinc ratio is too low, the concentration of zinc ions is high, so that the cell voltage is increased; if the acid zinc ratio is too high, the concentration of zinc ions is reduced, so that the precipitation of hydrogen on the cathode is aggravated, the precipitated zinc is re-dissolved, and the current efficiency is reduced. Thus, with the current density and acid zinc ratio of the present application, a decrease in current efficiency and an increase in cell voltage can be avoided.
Further, referring to fig. 2, the method for starting the zinc hydrometallurgy includes:
s700: mixing a part of the waste electrolyte with the first dense underflow, sulfur dioxide and a fourth sulfuric acid solution for reduction leaching
In this step, a further part of the waste electrolyte is mixed with the first dense underflow and the fourth sulfuric acid solution obtained in S200, and sulfur dioxide gas is introduced into the mixed solution, so that the first dense underflow is subjected to reduction leaching, and a reduced solution is obtained. Zinc ferrite is inevitably generated in the roasting process, and the zinc ferrite structure has quite large stability, so that the zinc ferrite is difficult to dissolve out if only mixed leaching is adopted. The zinc ferrite is dissolved through the following three steps: 1) Hydroxylation of water molecules on the surface of zinc ferrite solid to form complex molecules; 2) Dissociation and diffusion of the complex molecules occur on the surface of the zinc ferrite solid; 3) The reaction product was dissociated. In these three steps, fe 3+ The transfer process of ions from the surface of the solid zinc ferrite particles to the main body of the solution is a limiting link in the whole dissolution process. When the solution contains a large amount of Fe 3+ When ions are carried out, the transmission process is slower; when the solution contains a large amount of Fe 2+ When ions are transferred, the transfer process is faster, so that the zinc ferrite is treated by adopting the sulfur dioxide reduction method, and Fe can be completed while the high-efficiency decomposition of the zinc ferrite is realized 3+ Ion orientation Fe 2+ Ion conversion. The zinc ferrite is reduced and decomposed, and the iron in the reduced liquid exists in the form of Fe (II), so that the recovery rate of zinc can be improved, and favorable conditions can be created for the subsequent separation of zinc, indium and iron. Further, the volume ratio of the waste electrolyte to the first dense underflow to the fourth sulfuric acid solution is 1 (2-5): 1, preferably 1 (3-4): 1, more preferably 1:3:1. The inventors found that the volume ratio within the scope of the application is advantageous for the dissolution of zinc in zinc ferrite and the reduction behaviour of Fe (III) under the action of sulfur dioxide. Meanwhile, based on 1L of the first dense underflow, the dosage of sulfur dioxide is 30-50L, preferably 30-40L, more preferably 35-40L, and the inventor finds that if the dosage is too low, the leaching rate of zinc and the reduction rate of Fe (III) are lower; if the dosage is too high, the solution is already dissolved The decomposed sulfur dioxide tends to be saturated, so that the promotion effect of the continuous increase of the sulfur dioxide introduction amount on the reduction leaching is small, and the resource waste is caused. Therefore, by adopting the sulfur dioxide consumption of the application, on one hand, higher zinc dissolution rate and Fe (III) reduction rate can be ensured; on the other hand, the cost can be saved. In addition, sulfur dioxide reduction leaching is carried out at the temperature of 100-140 ℃. The inventors found that using a reaction temperature within the scope of the present application facilitates diffusion of the species to increase the leaching rate.
S800: mixing the reduced liquid with a third flocculant, and performing third thickening treatment
In the step, the purpose of solid-liquid separation is achieved by mixing the reduced liquid with a third flocculant and then performing third thickening treatment. Preferably, the process is carried out in a thickener, after which a third supernatant and a third thickened underflow are obtained. It should be noted that the thickener and the third flocculant are the same as described above, and will not be described here again. 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 liquid. 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 thus the time for concentration and clarification is increased; if the concentration of the third flocculating agent is too high or the adding amount is too high, the flocculating agent is also a transparent colloid, and if the adding amount is too large, the liquid becomes very viscous after reduction if the adding amount reaches a saturated state, and at the moment, the flocculating agent can not be gathered together and deposited in the liquid after reduction, and can also block equipment such as pipelines. Therefore, by adopting the concentration and the input amount of the third flocculant, the time for concentration and clarification can be shortened; on the other hand, the blocking of the pipeline can be avoided.
S900: mixing the third supernatant with iron powder for copper deposition
In this step, the third supernatant is mixed with the iron powder to displace the copper ions in the solution with the iron powder, and the copper ions precipitate out of the solution to become Fe 2+ Ions enter the solution, so that adverse effects on the subsequent electrolysis process caused by the existence of copper ions in the solution are avoided. Further toIn the above-mentioned third supernatant liquid of 1L, add the quality of iron powder to be 2-5 g, preferably 2-4 g, more preferably 3-4 g, the inventors found, if add too little, still there is a large amount of copper ions in the third supernatant liquid to be unable to precipitate out, copper ions in the electrolyte can precipitate out at the cathode in the subsequent electrolytic process, thus can accelerate the hydrogen gas to precipitate out, reduce the current efficiency, and make the zinc precipitated appear loose black, can appear the hole seriously; if too much is added, the iron powder is wasted. 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, the cost can be saved.
S1000: mixing the copper-precipitating solution with first limestone slurry for pre-neutralization
In the step, as the acidity of the solution after copper precipitation is too high, if the indium precipitation is directly neutralized, the consumption of the neutralizer is too large, so that the taste of indium slag is reduced, and the pre-neutralization process is carried out before the indium precipitation is neutralized. The process is based on the principle that the first limestone slurry reacts with sulfuric acid in the copper-precipitating solution, so that the acidity of the copper-precipitating solution is reduced to a certain degree, and the slag amount in the subsequent indium-precipitating process is reduced. After the pre-neutralization, a post-neutralization solution can be obtained. Specifically, the first lime slurry is lime slurry with a mass concentration of 20% (limestone powder contains CaCO) 3 > 93 wt%). Further, the above-mentioned preneutralization terminal acidity is 5 to 15g/L, preferably 8 to 15g/L, more preferably 8 to 12g/L, and the inventors have found that too low terminal acidity results in too much neutralizing agent consumption in the subsequent steps and reduced indium slag taste; if the final acidity is too high, a small amount of indium is lost to the slag. Therefore, the adoption of the pre-neutralization terminal acidity of the method can ensure that the subsequent indium slag has higher taste and recovery rate; on the other hand, the dosage of the neutralizing agent can be saved.
S1100: mixing the neutralized solution with second limestone slurry for indium precipitation
In the step, the pH of the neutralized liquid can be adjusted to 4.5-5.2 by mixing the neutralized liquid with the second limestone slurry, and at this time, indium ions in the neutralized liquid are hydrolyzed and precipitated, so that indium-precipitated liquid is obtained, and the chemical reaction equation is as follows: in (In) 3+ + 3H 2 O = In(OH) 3 + 3H + . Specifically, the second lime slurry is lime slurry with mass concentration of 20% (limestone powder contains CaCO) 3 > 93 wt%). Further, the volume ratio of the neutralized solution to the second limestone slurry is (20-30): 1, preferably (22-28): 1, and more preferably (24-26): 1, and the inventors found that if the volume ratio is too high, the neutralized indium precipitation may be insufficient, a large amount of indium ions still in the neutralized solution may remain in the solution, cannot be enriched, and may have adverse effects on the subsequent process; if the volume ratio is too low, the taste of indium in the slag is low. Therefore, the volume ratio of the application can ensure that the taste and the recovery rate of the indium slag are higher.
S1200: mixing the indium-precipitated liquid with oxygen to precipitate iron, and feeding the iron-precipitated liquid into the step (1) to perform mixed leaching
In the step, the indium-precipitated liquid is mixed with oxygen to carry out iron precipitation, and the mixture reacts at 170-200 ℃, wherein the chemical reaction equation is as follows: 2FeSO 4 +0.5O 2 +2H 2 O=Fe 2 O 3 +2H 2 SO 4 And (3) obtaining hematite slag and iron-precipitating liquid, and supplying the iron-precipitating liquid to S100 for mixed leaching, so that the recovery rate of zinc is improved. Preferably, the hematite iron removal process is performed in a hematite iron removal reactor. It should be noted that, a person skilled in the art may select a specific type of the hematite iron removal reactor according to actual needs, so long as the above functions can be achieved. Further, based on 1L of the above indium-deposited liquid, the amount of oxygen is 5 to 8L, preferably 6 to 8L, and the inventors found that if the amount of oxygen is too low, fe in the indium-deposited liquid cannot be ensured 2+ The ions are fully oxidized, hydrolyzed and precipitated, so that adverse effects can be generated on the subsequent process; 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 intake on the iron precipitation of the hematite is small, and the resource waste is caused. Therefore, the oxygen consumption of the application ensures that the iron deposition can be thoroughly performed on one hand; on the other hand, the resource waste can be avoided. In addition, the hematite iron deposition is carried out at the temperature of 170-200 ℃. Invention of the application It was found that the reaction temperature within the scope of the present application was useful for increasing the rate of hematite iron precipitation.
Further, referring to fig. 3, the method for starting the zinc hydrometallurgy includes:
s1300: mixing the third dense underflow with a fifth sulfuric acid solution for high acid leaching
In the step, zinc sulfide in the third dense underflow and zinc ferrite which is not decomposed in the sulfur dioxide reduction leaching process can be dissolved by carrying out high-acid leaching on the third dense underflow and a fifth sulfuric acid solution, so that high-acid leaching solution is obtained, the zinc-iron leaching rate of the third dense underflow is improved, the leaching slag amount is greatly reduced, and the chemical reaction equations generated by the high-acid leaching solution are as follows: znS+H 2 SO 4 = ZnSO 4 + H 2 S and ZnO.Fe 2 O 3 + 4H 2 SO 4 = ZnSO 4 + Fe(SO 4 ) 3 + 4H 2 O. Further, the concentration of the fifth sulfuric acid solution is 150-180 g/L, preferably 160-180 g/L, more preferably 165-170 g/L, and the inventor finds that if the concentration is too high, a part of lead and silver in the third dense underflow can be leached, so that the enrichment degree of lead and silver is reduced; while if the concentration is too low, the rate of high acid leaching is reduced. Therefore, the concentration of the fifth sulfuric acid solution can ensure the high acid leaching rate and improve the recovery rate of lead and silver. In addition, the volume ratio of the third dense underflow to the fifth sulfuric acid solution is 1 (2-5), preferably 1 (3-5), more preferably 1 (3-4), and the inventor finds that if the volume ratio is too high, the high acid leaching is possibly insufficient, so that the total leaching rate of zinc is reduced, and the leaching slag amount is increased; if the volume ratio is too low, the load of the subsequent thickening treatment increases. Therefore, by adopting the volume ratio of the application, the total leaching rate of zinc can be improved on one hand; on the other hand, the load of the subsequent thickening treatment can be prevented from being increased.
S1400: mixing the high acid leaching solution with a fourth flocculant for fourth thickening treatment, and returning the 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 high-acid leaching solution and the fourth flocculating agent. Preferably, the process is carried out in a thickener, and the fourth supernatant and the lead-silver slag are obtained after the thickening treatment, so that the lead-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 thickener and the fourth flocculant are the same as described above, and will not be described here again. Further, the mass concentration of the fourth flocculant is 0.5 to 2 per mill, based on 1L of the high acid leaching solution, the amount of the fourth flocculant to be added is 5 to 15mL, preferably 5 to 10mL, more preferably 6 to 8mL, and the inventor finds that if the concentration of the fourth flocculant is too low or the addition amount is too low, the flocculation effect is not ideal, so that the time for concentration and clarification is increased; if the concentration of the fourth flocculant is too high or the addition amount is too high, the high acid leaching solution becomes very viscous if the addition amount is too large because the flocculant is also a transparent colloid, and at the moment, the flocculant in the high acid leaching solution cannot be gathered together to precipitate, and can block equipment such as pipelines. Therefore, by adopting the concentration and the input amount of the fourth flocculating agent, the time for concentration and clarification can be shortened on one hand; on the other hand, the blocking of the pipeline can be avoided.
The inventors found that a first mixed leachate is obtained by mixing and leaching a part of zinc calcine with a first sulfuric acid solution, and a first supernatant fluid and a first dense underflow fluid containing zinc sulfate are obtained by mixing the first mixed leachate with a first flocculant and then performing a first thickening treatment. Then the first dense bottom flow is sent to a sulfur dioxide reduction leaching system to obtain reduced liquid, the reduced liquid is mixed with a third flocculating agent and then subjected to third thickening treatment to obtain third supernatant fluid and third dense bottom flow, wherein the third supernatant fluid is subjected to iron powder copper precipitation, preneutralization, neutralization indium precipitation and hematite iron precipitation to obtain hematite slag and iron precipitation liquid, and the iron precipitation liquid is returned to mixed leaching; and the third dense bottom is subjected to high acid leaching to obtain high acid leaching liquid, and the high acid leaching liquid is mixed with a fourth flocculating agent to carry out fourth thickening treatment to obtain a fourth supernatant and lead-silver slag, so that the starting of a slag treatment system is completed, and the lead-silver can be enriched and recovered.
Mixing the other part of the zinc calcine with a second sulfuric acid solution, a first supernatant containing zinc sulfate and oxygen to perform neutral leaching to obtain a second mixed leaching solution, mixing the second mixed leaching solution with a second flocculant, and performing 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 sent to be purified and then is subjected to electrolysis to obtain zinc and waste electrolyte, and part of the waste electrolyte is returned to be mixed and leached and/or neutral leached, so that the starting of a main process system is completed, the waste electrolyte can be recycled, and the production cost is saved.
By adopting the main process and slag treatment double-system parallel driving mode, the driving time is greatly shortened, the production efficiency is improved, and the production cost is reduced. The invention provides a novel starting scheme of a zinc hydrometallurgy system, which takes a novel technology of sulfur dioxide reduction leaching, neutralization indium precipitation and hematite iron precipitation as a core, so that the total leaching rate and recovery rate of zinc are improved, and the zinc content of leaching residues is reduced.
Examples
Step 1: mixing and leaching a part of zinc calcine with zinc content of 55wt% and a first sulfuric acid solution with 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 (Aisen 6000S) in a thickener, and then carrying out first thickening treatment, wherein the amount of the first flocculating agent (mass concentration is 1 per mill) required to be added is 10mL based on 1L of the first mixed leaching solution, and a first supernatant fluid and a first thickening bottom fluid containing zinc sulfate are obtained after the thickening treatment;
step 3: mixing the other part of the zinc calcine with a second sulfuric acid solution with the concentration of 165g/L, a first supernatant fluid containing zinc sulfate and oxygen for neutral leaching, wherein the ratio of the zinc calcine to the second sulfuric acid solution to the first supernatant fluid containing zinc sulfate is 1t:3 m: 2m, controlling the end-point pH of neutral leaching at 5.0, and oxygen content at 6Nm based on 1g zinc calcine 3 Thereby (a)Obtaining a second mixed leaching solution;
step 4: mixing the second mixed leaching solution and a second flocculating agent (Aisen 655S) in a thickener, performing second thickening treatment, adding 13mL of the second flocculating agent (mass concentration is 1 per mill) based on 1L of the second mixed leaching solution, obtaining a second supernatant and a second thickened bottom flow after thickening treatment, and returning the second thickened bottom flow to the step 1 for mixed leaching;
step 5: purifying and removing impurities from the second supernatant to obtain purified liquid, wherein the purified liquid adopts a continuous three-section zinc powder displacement precipitation method, and the main impurity ion concentration of the purified liquid meets the following requirements: cd is less than 1mg/L, co is less than 0.5mg/L;
step 6: mixing the purified solution with a third sulfuric acid solution with the concentration of 165g/L in an electrolytic tank for electrolysis, taking a pure casting aluminum plate as a cathode, taking 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 electrolytic process 2 The acid zinc ratio is 3.0-3.5, so that zinc and waste electrolyte are obtained, the zinc content in the obtained precipitated zinc is 99.995wt%, and part of the waste electrolyte is returned to step 1 and/or step 3;
Step 7: mixing a part of waste electrolyte with a first dense underflow, sulfur dioxide and a fourth sulfuric acid solution with the concentration of 165g/L, and carrying out reduction leaching at 110 ℃ to obtain a reduced solution, wherein the volume ratio of the waste electrolyte to the first dense underflow to the fourth sulfuric acid solution is 1:3:1, and the dosage of the sulfur dioxide is 38L based on 1L of the first dense underflow;
step 8: mixing the reduced liquid and a third flocculating agent (Aisen 6000S) in a thickener, and performing third thickening treatment to obtain a third supernatant and a third thickening bottom flow, wherein the amount of the third flocculating agent (mass concentration is 1 per mill) required to be added is 10mL based on 1L of the reduced liquid;
step 9: mixing the third supernatant with iron powder for copper precipitation to obtain copper precipitation solution, adding iron powder (reduced iron powder, fe-containing) into 1L of the third supernatant 0 > 95 wt%) of 3g;
step 10: will sink toCopper post-solution and limestone slurry with 20% mass concentration (limestone powder contains CaCO) 3 Mixing more than 93wt percent for preneutralization, wherein the terminal acidity of preneutralization is 12g/L, thereby obtaining a neutralized solution;
step 11: mixing the neutralized solution with limestone slurry (limestone powder contains CaCO) with 20% mass concentration 3 Mixing more than 93wt percent to precipitate indium, and neutralizing the end point pH of the precipitated indium to be 4.8 to obtain a post-indium-precipitation solution, wherein the volume ratio of the post-neutralization solution to limestone slurry is 25:1;
Step 12: feeding the indium-precipitated liquid and oxygen into a hematite iron removal reactor to carry out mixed iron precipitation at 180 ℃ to obtain hematite slag and iron-precipitated liquid, wherein the zinc content of the hematite slag is 58-60 wt%, and feeding the iron-precipitated liquid into the step 1 to carry out mixed leaching, and the oxygen consumption is 8L based on 1L of the indium-precipitated liquid;
step 13: mixing the third dense underflow with a fifth sulfuric acid solution with the concentration of 165g/L for high-acid leaching to obtain a high-acid leaching solution, wherein the volume ratio of the third dense underflow to the fifth sulfuric acid solution is 1:3;
step 14: mixing the high-acid leaching solution and a fourth flocculating agent (Aisen 6000S) in a thickener for fourth thickening treatment, wherein the amount of the fourth flocculating agent (mass concentration is 1 per mill) required to be added is 7mL based on 1L of the high-acid leaching solution, obtaining a fourth supernatant and lead silver slag after thickening treatment, wherein the zinc content of the lead silver slag is 3-5 wt%, and returning the fourth supernatant to the step 7 for reduction leaching.
In the description of the present specification, a description referring to terms "one embodiment," "some embodiments," "examples," "specific examples," or "some examples," etc., means 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 present invention. In this specification, schematic representations of the above terms are not necessarily directed 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, the different embodiments or examples described in this specification and the features of the different embodiments or examples may be combined and combined by those skilled in the art without contradiction.
While embodiments of the present invention have been shown and described above, it will be understood that the above embodiments are illustrative and not to be construed as limiting the invention, and that variations, modifications, alternatives and variations may be made to the above embodiments by one of ordinary skill in the art within the scope of the invention.
Claims (6)
1. A method of operating a zinc hydrometallurgy system, comprising:
(1) Mixing and leaching a part of zinc calcine and a first sulfuric acid solution so as to obtain a first mixed leaching solution;
(2) Mixing the first mixed leaching solution and a first flocculating agent, and then carrying out first thickening treatment so as to obtain a first supernatant fluid containing zinc sulfate and a first thickening bottom fluid;
(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 and a second flocculating agent, and then carrying out second thickening treatment so as to obtain a second supernatant fluid and a second thickening bottom flow, and returning the second thickening bottom flow to the step (1) for carrying out the mixed leaching;
(5) Purifying and impurity-removing the second supernatant to obtain a purified liquid;
(6) Mixing the purified solution with a third sulfuric acid solution for electrolysis so as to obtain zinc and waste electrolyte, and returning a part of the waste electrolyte to the step (1) and/or the step (3);
(7) Mixing a further portion of the spent electrolyte with the first dense underflow, sulfur dioxide and a fourth sulfuric acid solution for reduction leaching to obtain a reduced liquor;
(8) Mixing the reduced liquid with a third flocculant and then carrying out third thickening treatment so as to obtain a third supernatant and a third thickening bottom flow;
(9) Mixing the third supernatant with iron powder for copper precipitation so as to obtain copper precipitation post-liquid;
(10) Mixing the copper-precipitated liquid with first limestone ore pulp for preneutralization so as to obtain a neutralized liquid;
(11) Mixing the neutralized solution with second limestone ore pulp to precipitate indium so as to obtain indium-precipitated solution;
(12) Mixing the indium-precipitated liquid with oxygen to precipitate iron so as to obtain hematite slag and iron-precipitated liquid, and supplying the iron-precipitated liquid to the step (1) to perform mixed leaching;
(13) Mixing the third dense underflow with a fifth sulfuric acid solution for high acid leaching so as to obtain a high acid leaching solution;
(14) 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, returning the fourth supernatant to the step (7) for the reduction leaching,
in the step (1), the concentration of the first sulfuric acid solution is 150-180 g/L;
in the step (1), the solid-to-liquid ratio of the mixed leaching process is 1t: (3-6) m 3 ；
In the step (3), the concentration of the second sulfuric acid solution is 150-180 g/L;
in the step (3), the ratio of the zinc calcine to the second sulfuric acid solution to the first supernatant fluid containing zinc sulfate is 1t (2-5) m 3 ：（2~5）m 3 ；
In the step (3), the oxygen is used in an amount of 4-8 Nm based on 1t of zinc calcine 3 ；
In the step (6), the concentration of the third sulfuric acid solution is 150-180 g/L;
in the step (6), the mixing volume ratio of the purified liquid and the third sulfuric acid solution is 1 (20-30);
in the step (6), the electrolysis process controls the current density to be 360-420A/m 2 The acid zinc ratio is 3.0-3.5;
in the step (7), the concentration of the fourth sulfuric acid solution is 150-180 g/L;
in the step (7), the volume ratio of the waste electrolyte to the first dense underflow to the fourth sulfuric acid solution is 1 (2-5): 1;
In the step (7), based on 1L of the first dense underflow, the dosage of the sulfur dioxide is 30-50L;
in the step (9), based on 1L of the third supernatant, the mass of the iron powder to be added is 2-5 g;
in the step (10), the content of the pre-neutralized end point sulfuric acid is 5-15 g/L;
in the step (11), the volume ratio of the neutralized liquid to the second limestone slurry is (20-30) 1, and the final pH of indium precipitation is 4.5-5.2;
in the step (12), based on 1L of the indium-precipitated liquid, the using amount of the oxygen is 5-8L;
in the step (13), the concentration of the fifth sulfuric acid solution is 150-180 g/L;
in the step (13), the volume ratio of the third dense underflow to the fifth sulfuric acid solution is 1 (2-5).
2. The method according to claim 1, wherein in the step (2), the mass concentration of the first flocculant is 0.5 to 2 per mill, and the amount of the first flocculant added is 7 to 15ml based on 1L of the first mixed leachate.
3. The method according to claim 1, wherein in the step (4), the mass concentration of the second flocculant is 0.5 to 2 per mill, and the amount of the second flocculant added is 7 to 15ml based on 1L of the second mixed leachate.
4. The method according to claim 1, wherein in the step (5), the purifying and impurity removing process comprises a first-stage purification, a second-stage purification and a third-stage purification, wherein the first-stage purification is to add zinc powder to remove Cu and Cd, the second-stage purification is to add zinc powder and antimony salt to remove Co, the third-stage purification is to add zinc powder to remove the balance Cd, and the main impurity ion concentration of the purified liquid meets the following requirements: cd is less than 1mg/L, co is less than 0.5mg/L.
5. The method according to claim 1, wherein in the step (8), the mass concentration of the third flocculant is 0.5 to 2 per mill, and the amount of the third flocculant added is 10 to 15ml based on 1L of the reduced liquid.
6. The method according to claim 1, wherein in the step (14), the mass concentration of the fourth flocculant is 0.5-2 per mill, and the amount of the fourth flocculant added is 5-15 ml based on 1L of the high acid leaching solution.
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