CN116574908B - Process for jointly recycling zinc and indium by means of open-circuit impurity removal of electrolyte in zinc smelting process - Google Patents
Process for jointly recycling zinc and indium by means of open-circuit impurity removal of electrolyte in zinc smelting process Download PDFInfo
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- CN116574908B CN116574908B CN202310345862.7A CN202310345862A CN116574908B CN 116574908 B CN116574908 B CN 116574908B CN 202310345862 A CN202310345862 A CN 202310345862A CN 116574908 B CN116574908 B CN 116574908B
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- 239000011701 zinc Substances 0.000 title claims abstract description 99
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 87
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 87
- 229910052738 indium Inorganic materials 0.000 title claims abstract description 72
- APFVFJFRJDLVQX-UHFFFAOYSA-N indium atom Chemical compound [In] APFVFJFRJDLVQX-UHFFFAOYSA-N 0.000 title claims abstract description 70
- 238000000034 method Methods 0.000 title claims abstract description 60
- 238000003723 Smelting Methods 0.000 title claims abstract description 34
- 239000003792 electrolyte Substances 0.000 title claims abstract description 28
- 239000012535 impurity Substances 0.000 title claims abstract description 17
- 238000004064 recycling Methods 0.000 title claims abstract description 10
- 238000002386 leaching Methods 0.000 claims abstract description 112
- 239000007788 liquid Substances 0.000 claims abstract description 43
- 239000002893 slag Substances 0.000 claims abstract description 35
- 239000011777 magnesium Substances 0.000 claims abstract description 30
- 229910052749 magnesium Inorganic materials 0.000 claims abstract description 28
- FYYHWMGAXLPEAU-UHFFFAOYSA-N Magnesium Chemical compound [Mg] FYYHWMGAXLPEAU-UHFFFAOYSA-N 0.000 claims abstract description 26
- 230000003472 neutralizing effect Effects 0.000 claims abstract description 25
- 239000000460 chlorine Substances 0.000 claims abstract description 22
- 230000001376 precipitating effect Effects 0.000 claims abstract description 20
- 239000002699 waste material Substances 0.000 claims abstract description 20
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 claims abstract description 18
- 229910052801 chlorine Inorganic materials 0.000 claims abstract description 18
- 229910052731 fluorine Inorganic materials 0.000 claims abstract description 18
- 239000000779 smoke Substances 0.000 claims abstract description 18
- ZAMOUSCENKQFHK-UHFFFAOYSA-N Chlorine atom Chemical compound [Cl] ZAMOUSCENKQFHK-UHFFFAOYSA-N 0.000 claims abstract description 16
- 239000000428 dust Substances 0.000 claims abstract description 16
- 239000011737 fluorine Substances 0.000 claims abstract description 16
- 239000002253 acid Substances 0.000 claims abstract description 15
- 229910052787 antimony Inorganic materials 0.000 claims abstract description 15
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 claims abstract description 15
- 229910052785 arsenic Inorganic materials 0.000 claims abstract description 15
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 claims abstract description 15
- 238000000926 separation method Methods 0.000 claims abstract description 14
- 238000011084 recovery Methods 0.000 claims abstract description 13
- CSNNHWWHGAXBCP-UHFFFAOYSA-L Magnesium sulfate Chemical compound [Mg+2].[O-][S+2]([O-])([O-])[O-] CSNNHWWHGAXBCP-UHFFFAOYSA-L 0.000 claims abstract description 10
- 230000007935 neutral effect Effects 0.000 claims abstract description 10
- LWUVWAREOOAHDW-UHFFFAOYSA-N lead silver Chemical compound [Ag].[Pb] LWUVWAREOOAHDW-UHFFFAOYSA-N 0.000 claims abstract description 9
- 229910000029 sodium carbonate Inorganic materials 0.000 claims abstract description 9
- UOURRHZRLGCVDA-UHFFFAOYSA-D pentazinc;dicarbonate;hexahydroxide Chemical compound [OH-].[OH-].[OH-].[OH-].[OH-].[OH-].[Zn+2].[Zn+2].[Zn+2].[Zn+2].[Zn+2].[O-]C([O-])=O.[O-]C([O-])=O UOURRHZRLGCVDA-UHFFFAOYSA-D 0.000 claims abstract description 7
- 239000013078 crystal Substances 0.000 claims abstract description 5
- 238000001704 evaporation Methods 0.000 claims abstract description 5
- 229910052943 magnesium sulfate Inorganic materials 0.000 claims abstract description 5
- 235000019341 magnesium sulphate Nutrition 0.000 claims abstract description 5
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 claims description 20
- YCKRFDGAMUMZLT-UHFFFAOYSA-N Fluorine atom Chemical compound [F] YCKRFDGAMUMZLT-UHFFFAOYSA-N 0.000 claims description 15
- 238000001556 precipitation Methods 0.000 claims description 13
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 12
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 11
- 229910052760 oxygen Inorganic materials 0.000 claims description 11
- 239000001301 oxygen Substances 0.000 claims description 11
- 239000003795 chemical substances by application Substances 0.000 claims description 10
- 239000011787 zinc oxide Substances 0.000 claims description 8
- FMRLDPWIRHBCCC-UHFFFAOYSA-L Zinc carbonate Chemical compound [Zn+2].[O-]C([O-])=O FMRLDPWIRHBCCC-UHFFFAOYSA-L 0.000 claims description 6
- 239000007787 solid Substances 0.000 claims description 6
- 239000011667 zinc carbonate Substances 0.000 claims description 6
- 229910000010 zinc carbonate Inorganic materials 0.000 claims description 6
- 235000004416 zinc carbonate Nutrition 0.000 claims description 6
- 239000003350 kerosene Substances 0.000 claims description 4
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims description 3
- 238000001914 filtration Methods 0.000 claims description 3
- 239000011593 sulfur Substances 0.000 claims description 3
- 229910052717 sulfur Inorganic materials 0.000 claims description 3
- 238000005188 flotation Methods 0.000 claims description 2
- 239000000203 mixture Substances 0.000 claims description 2
- 239000012716 precipitator Substances 0.000 claims description 2
- PXGOKWXKJXAPGV-UHFFFAOYSA-N Fluorine Chemical compound FF PXGOKWXKJXAPGV-UHFFFAOYSA-N 0.000 abstract 1
- QSHDDOUJBYECFT-UHFFFAOYSA-N mercury Chemical compound [Hg] QSHDDOUJBYECFT-UHFFFAOYSA-N 0.000 abstract 1
- 238000000247 postprecipitation Methods 0.000 abstract 1
- 239000000243 solution Substances 0.000 description 24
- 239000004071 soot Substances 0.000 description 4
- JLVVSXFLKOJNIY-UHFFFAOYSA-N Magnesium ion Chemical compound [Mg+2] JLVVSXFLKOJNIY-UHFFFAOYSA-N 0.000 description 3
- 238000006243 chemical reaction Methods 0.000 description 3
- 238000005516 engineering process Methods 0.000 description 3
- 238000009854 hydrometallurgy Methods 0.000 description 3
- 229910001425 magnesium ion Inorganic materials 0.000 description 3
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 2
- 230000008901 benefit Effects 0.000 description 2
- 230000001276 controlling effect Effects 0.000 description 2
- 239000008151 electrolyte solution Substances 0.000 description 2
- 238000000605 extraction Methods 0.000 description 2
- 239000002920 hazardous waste Substances 0.000 description 2
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 2
- UCHHTHXUPVVOIA-UHFFFAOYSA-N [Mg][Zn][In] Chemical compound [Mg][Zn][In] UCHHTHXUPVVOIA-UHFFFAOYSA-N 0.000 description 1
- 230000002411 adverse Effects 0.000 description 1
- 239000012141 concentrate Substances 0.000 description 1
- 238000005137 deposition process Methods 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 238000004070 electrodeposition Methods 0.000 description 1
- 238000009776 industrial production Methods 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- 239000011133 lead Substances 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 229910052751 metal Inorganic materials 0.000 description 1
- 239000002184 metal Substances 0.000 description 1
- 235000010755 mineral Nutrition 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 238000006386 neutralization reaction Methods 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 230000001105 regulatory effect Effects 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 description 1
- 229960001763 zinc sulfate Drugs 0.000 description 1
- 229910000368 zinc sulfate Inorganic materials 0.000 description 1
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/007—Wet processes by acid leaching
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B11/00—Obtaining noble metals
- C22B11/04—Obtaining noble metals by wet processes
- C22B11/042—Recovery of noble metals from waste materials
- C22B11/044—Recovery of noble metals from waste materials from pyrometallurgical residues, e.g. from ashes, dross, flue dust, mud, skim, slag, sludge
-
- 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
- C22B13/00—Obtaining lead
- C22B13/04—Obtaining lead by wet processes
- C22B13/045—Recovery from waste materials
-
- 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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/26—Refining solutions containing zinc values, e.g. obtained by leaching zinc ores
-
- 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
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
-
- 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
- C22B26/00—Obtaining alkali, alkaline earth metals or magnesium
- C22B26/20—Obtaining alkaline earth metals or magnesium
- C22B26/22—Obtaining magnesium
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
- C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
- C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
- C22B3/3846—Phosphoric acid, e.g. (O)P(OH)3
-
- 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
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/02—Working-up flue dust
-
- 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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- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Metallurgy (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Organic Chemistry (AREA)
- Geology (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- Life Sciences & Earth Sciences (AREA)
- Geochemistry & Mineralogy (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a process for jointly recycling zinc and indium by opening and removing impurities from electrolyte in a zinc smelting process, which comprises the steps of directly adopting magnesium-containing waste electrolyte to perform primary neutral leaching on smoke dust generated in the zinc smelting process, and obtaining primary leaching liquid and primary leaching slag after solid-liquid separation; then neutralizing the first-stage leaching solution with sodium carbonate to precipitate zinc to obtain basic zinc carbonate and a post-precipitation zinc solution; evaporating and crystallizing the zinc-precipitated solution to obtain magnesium sulfate crystals; carrying out secondary acid leaching on the primary leaching slag, and controlling leaching conditions to obtain secondary leaching liquid and lead-silver slag; and extracting the second-stage leaching solution to obtain indium-rich liquid, neutralizing the indium-rich liquid, and precipitating indium to finally obtain indium-rich slag. The process has the characteristics of simple flow, novel process, high-efficiency comprehensive recovery of zinc, magnesium and indium, capability of effectively solving the problem of high magnesium content in a zinc smelting system, and capability of simultaneously opening the way of impurities such as fluorine, chlorine, arsenic, antimony and the like.
Description
Technical Field
The invention relates to the technical field of chemical smelting treatment, in particular to a process for jointly recycling zinc and indium by removing impurities from electrolyte in an open circuit in a zinc smelting process.
Background
Metal indium is one of important strategic mineral resources, which is an indispensable raw material for the development of strategic emerging industries, and recovery of indium in zinc smelting processes is an important raw indium source. At present, part of smeltery adopts oxygen pressure leaching technology to treat high-iron zinc concentrate, after floatation and hot filtration of leached slag, tailings and sulfur slag are subjected to oxygen-enriched side-blown smelting technology to recover lead and zinc, and smoke dust is produced at the same time, and part of smeltery returns the lead and zinc to a main system. Because the smoke contains indium and fluorine, chlorine, arsenic and antimony, on one hand, the added value of the smoke is reduced, and on the other hand, harmful impurities such as fluorine, chlorine, arsenic and antimony are circulated in the system, so that the subsequent process is not facilitated. Therefore, how to increase the added value of the smoke dust and simultaneously the harmful impurities such as open-circuit fluorine, chlorine, arsenic, antimony and the like become the problems to be solved urgently by zinc smelters. In addition, in the zinc smelting wet method system, the waste electrolyte contains excessive acid and contains a large amount of magnesium ions, if the waste electrolyte is returned to the main system, the magnesium ions can cause adverse effects on subsequent electrodeposition, and the problems of unbalanced acid and excessive magnesium content of the main system are also solved by a zinc smelting plant.
Patent CN 108411128A discloses a method for efficiently enriching indium from low-grade indium-containing smelting slag, which comprises the steps of carrying out concentrated leaching on zinc-smelting underflow slag by adopting high-temperature high acid, and then neutralizing indium to recover indium. The method is mainly aimed at zinc smelting underflow slag, and does not relate to open-circuit impurity removal of waste electrolyte magnesium and comprehensive recovery of zinc in zinc smelting. The CN 104141046A patent discloses a method for recovering indium and zinc from indium-containing zinc oxide soot, which adopts the processes of high-temperature volatilization enrichment, high-indium soot leaching, neutral leaching residue leaching and acid leaching liquid extraction and indium extraction to recover indium, and does not relate to open-circuit impurity removal of waste electrolyte magnesium in zinc smelting and comprehensive recovery of zinc. CN 109628745A patent discloses a method for removing magnesium ions from waste electrolytic solution of zinc hydrometallurgy, which adopts a low-temperature crystallization-centrifugal separation method to remove magnesium from waste electrolytic solution of zinc hydrometallurgy, but does not involve comprehensive recovery of indium and zinc. The CN 108624755A patent discloses a method for opening a circuit of impurities Mg and Cl in a zinc hydrometallurgy system, which utilizes zinc oxide materials (calcine, zinc oxide smoke dust and the like) in the flow of a zinc smelting enterprise to treat zinc sulfate solution with high fluorine, chlorine and magnesium, and simultaneously, the impurities Cl and Mg in the open circuit are not related to the comprehensive recovery of indium and zinc when zinc is recovered.
In view of the above, a plurality of problems existing in the zinc smelting process are necessary to be considered together, so that four problems of indium zinc magnesium recovery, excessive system acid, excessive solution magnesium content and open circuit of fluorine, chlorine, arsenic and antimony are comprehensively solved, and the economic benefit of zinc smelting enterprises is improved.
Disclosure of Invention
The invention aims to solve the technical problem of providing a process which has simpler treatment flow and can simultaneously solve four problems of comprehensive recovery of indium and zinc in smoke dust, excessive acid in a zinc smelting process by oxygen pressure leaching, excessive magnesium content in solution and open circuit of fluorine, chlorine, arsenic and antimony, and the like.
In order to solve the technical problems, the invention adopts the following technical scheme: a process for jointly recycling zinc and indium by using an electrolyte in an open-circuit impurity removal mode in a zinc smelting process is characterized by comprising the following steps of: the method comprises the following steps:
(1) And (3) leaching: directly adopting magnesium-containing waste electrolyte in a zinc-oxygen pressure leaching system to directly carry out primary neutral leaching on smoke dust, wherein the leaching temperature is 60-80 ℃, the leaching time is 1-2 h, and primary leaching liquid and primary leaching slag are obtained after solid-liquid separation; in the first leaching process, zinc oxide reacts with sulfuric acid to form Zn 2+ Is brought into solution. In the first neutral leaching process, a small amount of indium is added In 3+ The form goes into solution and mainly the following reactions occur: znO+H 2 SO 4 =ZnSO 4 +H 2 O，In 2 O 3 +3H 2 SO 4 =In 2 (SO4) 3 +3H2O。
(2) Neutralizing and precipitating zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the neutralizing and precipitating zinc temperature is 20-30 ℃, the neutralizing and precipitating zinc end point pH is 6.8-7.8, and the alkaline zinc carbonate and the precipitating zinc post-liquid are obtained after solid-liquid separation; the step is a neutralization zinc precipitation process, and mainly comprises the following reactions: 2Zn 2+ +CO 3 2- +2OH - =Zn 2 CO 3 (OH) 2 。
(3) Evaporating and crystallizing the zinc-precipitated liquid obtained in the step (2) to obtain magnesium sulfate crystals;
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 60-80 ℃, the leaching time is 4-5 h, and the solid ratio of leaching liquid is 4:1, obtaining a second-stage leaching solution and lead silver slag after solid-liquid separation, wherein the final pH value is 1.0-1.2; during the second-stage acid leaching, zinc and indium enter the solution in the same ionic form as each element during the first-stage leaching, and as most zinc enters the solution during the first-stage neutral leaching, indium enters the second-stage acid leaching along with the underflow, so that the leaching of indium is mainly performed during the second-stage acid leaching.
(5) Extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of a load phase to HCl of 1:1, so as to obtain indium-rich liquid and raffinate;
(6) Neutralizing and precipitating indium in the indium-rich liquid obtained in the step (5), wherein the indium precipitating process firstly adopts NaOH for preneutralization, and the pH value is adjusted to be 1; and then zinc oxide is utilized to carry out indium precipitation, the pH value of the indium precipitation is controlled to be 3.5-4.0, and the indium-rich slag is obtained. The indium deposition process mainly comprises the following reactions: 2In 3+ +6OH - =2In(OH) 3 。
The smoke dust in the step (1) is derived from oxygen pressure leaching slag generated in the process of oxygen pressure leaching zinc smelting, and tailings and sulfur slag obtained after flotation and hot filtration are generated in the process of lead and zinc recovery by adopting oxygen-enriched side-blown smelting and fuming processes; the components of the composition comprise: zn: 50.5-55%, pb:20.1 to 27.43%, in:0.3 to 0.5%, as 0.3 to 0.35%, sb: 0.25-0.28%, F: 0.03-0.06%, cl: 0.1-0.16%, fe: 4.1-6.2%, and 400-600 g/t of Ag.
The magnesium-containing waste electrolyte component in the step (1) comprises the following components: zn: 44.5-47.8 g/L, mg: 25-28 g/L, H 2 SO 4 ：160~170g/L。
In the step (1), the pH of the neutral leaching end point is controlled to be 4.2-4.4, the leaching rate of zinc in smoke dust is more than 80%, the leaching rate of indium is less than 8%, and the solid ratio of leaching liquid is 4.5:1.
in the step (2), sodium carbonate is adopted as a precipitator, and the dosage of the sodium carbonate is controlled to be 2.3t/t.Zn, so that the zinc precipitation rate reaches more than 98 percent, and the magnesium entrainment rate is less than 6 percent.
In the step (2), the obtained basic zinc carbonate has a zinc content of more than 45%, a magnesium content of less than 1.5%, a fluorine content of less than 0.009%, a chlorine content of less than 0.005%, an arsenic content of less than 0.007% and an antimony content of less than 0.002%.
In the step (2) and the step (4), most fluorine and antimony enter lead silver slag, part chlorine enters basic zinc carbonate, and arsenic is in a zinc precipitation solution.
The invention has the advantages that: 1. the invention adopts the magnesium-containing waste electrolyte in the zinc-oxygen pressure leaching system to directly leach the soot, and uses hazardous waste to treat hazardous waste, thereby avoiding the problem of separate treatment of the waste electrolyte and the soot, obviously reducing the cost, simplifying the treatment process, and being environment-friendly and efficient;
2. the invention develops a smoke dust and waste electrolyte cooperative treatment technology in the zinc smelting process of oxygen pressure leaching, can simultaneously solve four problems of comprehensive recovery of indium and zinc in the smoke dust, excessive acid content in a system in the zinc smelting process of oxygen pressure leaching, excessive magnesium content in a solution and open circuit of fluorine, chlorine, arsenic and antimony, not only improves the added value of the smoke dust, but also solves the difficult problem in the industrial production of zinc oxygen pressure leaching.
Drawings
FIG. 1 is a process flow diagram of the present invention.
Description of the embodiments
The invention is further illustrated by the following specific examples in connection with fig. 1:
example 1: the method is characterized in that the ash and the waste electrolyte of a certain zinc smelting plant are treated according to the process of the invention, and the ash component and the waste electrolyte component are respectively shown in the following table I and table II:
list one
| Element(s) | Zn,% | In,% | Fe,% | F,% | Cl,% | As,% | Sb,% |
| Content of | 54.1 | 0.49 | 4.4 | 0.06 | 0.16 | 0.35 | 0.28 |
Watch II
| Element(s) | Zn,% | Mg,% | H 2 SO 4 ,g/L |
| Content of | 46.79 | 27.88 | 165 |
The method comprises the following specific steps:
(1) Directly adopting magnesium-containing waste electrolyte to carry out one-stage neutral leaching on smoke dust, wherein the leaching temperature is 80 ℃, the leaching time is 1h, and the solid ratio of leaching liquid is 4.5:1, controlling the pH of a leaching end point at 4.3, and obtaining a first-stage leaching solution and a first-stage leaching slag after solid-liquid separation, wherein the leaching rate of Zn is 82.1%, the leaching rate of In is 6.8%, the leaching rate of F is 2.3%, and the leaching rate of Cl is as follows: 49.12%, and the leaching rate of As is As follows: 79.5 percent, the leaching rate of Sb is as follows: 4.58%.
(2) And (3) neutralizing and precipitating the zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the dosage of the neutralizing agent is 2.3t/t.Zn, the neutralizing and precipitating temperature is 20 ℃, the end point pH of the neutralizing and precipitating zinc is 7.4, and the alkaline zinc carbonate and the post-precipitating zinc solution are obtained after solid-liquid separation, wherein the precipitation rate of zinc is 98.9%, the entrainment rate of magnesium is 5.2%, the zinc content of the alkaline zinc carbonate is 45.6%, the magnesium content is 1.16%, the chlorine content is 0.0012%, the arsenic content is 0.0068%, the antimony content is 0.0019% and the fluorine content is 0.0089%.
(3) And (3) evaporating and crystallizing the zinc-precipitated solution obtained in the step (2) to obtain magnesium sulfate crystals.
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 80 ℃, the leaching time is 5h, and the solid ratio of leaching liquid is 4:1, the final pH value is 1.13, and the second-stage leaching solution and lead-silver slag are obtained after solid-liquid separation, wherein the lead-silver slag contains 0.0037 percent of chlorine, 0.11 percent of arsenic, 0.42 percent of antimony and 0.028 percent of fluorine.
(5) And (3) extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of the load phase to the HCl of 1:1, so as to obtain indium-rich liquid and raffinate.
(6) And (3) neutralizing the indium-rich liquid obtained in the step (5) to precipitate indium, wherein the indium precipitation process firstly adopts NaOH to perform preneutralization, the pH value is regulated to be about 1.0, then zinc oxide is utilized to precipitate indium, and the pH value of the indium precipitation is controlled to be 3.5, so that the indium-rich slag is obtained.
In the embodiment, the basic zinc carbonate with the grade of 45.6 percent and the indium-rich slag with the grade of 63kg/t are finally obtained, and the recovery rates of zinc and indium are respectively as follows: 90.1%, 82.3%.
Example 2: the method is characterized in that the ash and the waste electrolyte of a certain zinc smelting plant are treated according to the process of the invention, and the ash component and the waste electrolyte component are respectively shown in the following table III and table IV:
watch III
| Element(s) | Zn,% | In,% | Fe,% | F,% | Cl,% | As,% | Sb,% |
| Content of | 54.9 | 0.47 | 4.7 | 0.05 | 0.15 | 0.33 | 0.27 |
Table four
| Element(s) | Zn,% | Mg,% | H 2 SO 4 ,g/L |
| Content of | 45.52 | 26.99 | 168 |
The method comprises the following specific steps:
(1) Directly adopting magnesium-containing waste electrolyte to carry out one-stage neutral leaching on smoke dust, wherein the leaching temperature is 60 ℃, the leaching time is 1.5h, and the solid-to-liquid ratio of the leaching liquid is 4.5:1, controlling the pH of a leaching end point at 4.25, and obtaining a first-stage leaching solution and a first-stage leaching slag after solid-liquid separation, wherein the leaching rate of Zn is 81.5%, the leaching rate of In is 6.1%, the leaching rate of F is 2.0%, and the leaching rate of Cl is as follows: 47.12%, and the leaching rate of As is As follows: 78.5% and the leaching rate of Sb is 4.6%.
(2) And (3) neutralizing and precipitating the zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the dosage of the neutralizing agent is 2.3t/t.Zn, the neutralizing and precipitating temperature is 25 ℃, the end point pH of the neutralizing and precipitating zinc is 7.3, and the alkaline zinc carbonate and the post-precipitating zinc solution are obtained after solid-liquid separation, wherein the precipitation rate of zinc is 98.5%, the entrainment rate of magnesium is 5.0%, the zinc content of the alkaline zinc carbonate is 46.1%, the magnesium content is 1.10%, the chlorine content is 0.0011%, the arsenic content is 0.0065%, the antimony content is 0.0019% and the fluorine content is 0.0087%.
(3) And (3) evaporating and crystallizing the zinc-precipitated solution obtained in the step (2) to obtain magnesium sulfate crystals.
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 70 ℃, the leaching time is 4.5h, and the solid-to-liquid ratio of the leaching liquid is 4:1, the final pH value is 1.02, and a second-stage leaching solution and lead-silver slag are obtained after solid-liquid separation, wherein the lead-silver slag contains 0.0038% of chlorine, 0.12% of arsenic, 0.44% of antimony and 0.024% of fluorine.
(5) And (3) extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of the load phase to the HCl of 1:1, so as to obtain indium-rich liquid and raffinate.
(6) And (3) neutralizing the indium-rich liquid obtained in the step (5) to precipitate indium, wherein the indium precipitation process firstly adopts NaOH to perform preneutralization, the pH value is adjusted to be 1.1, then zinc oxide is used for precipitating indium, and the pH value of the precipitated indium is controlled to be 3.7, so that the indium-rich slag is obtained.
In the embodiment, the basic zinc carbonate with the grade of 45.3 percent and the indium-rich slag with the grade of 62.5kg/t are finally obtained, and the recovery rates of zinc and indium are respectively as follows: 89.98% and 81.12%.
The foregoing detailed description of the invention has been presented for purposes of illustration and description, but is not intended to limit the scope of the invention, i.e., the invention is not limited to the details shown and described.
Claims (5)
1. A process for jointly recycling zinc and indium by using an electrolyte in an open-circuit impurity removal mode in a zinc smelting process is characterized by comprising the following steps of: the method comprises the following steps:
(1) And (3) leaching: directly adopting magnesium-containing waste electrolyte in a zinc-oxygen pressure leaching system to directly carry out primary neutral leaching on smoke dust, wherein the leaching temperature is 60-80 ℃, the leaching time is 1-2 h, and primary leaching liquid and primary leaching slag are obtained after solid-liquid separation;
(2) Neutralizing and precipitating zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the neutralizing and precipitating zinc temperature is 20-30 ℃, the neutralizing and precipitating zinc end point pH is 6.8-7.8, and the alkaline zinc carbonate and the precipitating zinc post-liquid are obtained after solid-liquid separation;
(3) Evaporating and crystallizing the zinc-precipitated liquid obtained in the step (2) to obtain magnesium sulfate crystals;
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 60-80 ℃, the leaching time is 4-5 h, and the solid ratio of leaching liquid is 4:1, obtaining a second-stage leaching solution and lead silver slag after solid-liquid separation, wherein the final pH value is 1.0-1.2;
(5) Extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of a load phase to HCl of 1:1, so as to obtain indium-rich liquid and raffinate;
(6) Neutralizing and precipitating indium in the indium-rich liquid obtained in the step (5), wherein the indium precipitating process firstly adopts NaOH for preneutralization, and the pH value is adjusted to be 1; then zinc oxide is utilized to carry out indium precipitation, the pH value of the indium precipitation is controlled to be 3.5-4.0, and indium-rich slag is obtained;
the smoke dust in the step (1) is derived from oxygen pressure leaching slag generated in the process of oxygen pressure leaching zinc smelting, and tailings and sulfur slag obtained after flotation and hot filtration are generated in the process of lead and zinc recovery by adopting oxygen-enriched side-blown smelting and fuming processes; the components of the composition comprise: zn: 50.5-55%, pb:20.1 to 27.43%, in:0.3 to 0.5%, as 0.3 to 0.35%, sb: 0.25-0.28%, F: 0.03-0.06%, cl: 0.1-0.16%, fe: 4.1-6.2%, ag is 400-600 g/t;
the magnesium-containing waste electrolyte component in the step (1) comprises the following components: zn: 44.5-47.8 g/L, mg: 25-28 g/L, H 2 SO 4 ：160~170g/L。
2. The process for jointly recycling zinc and indium by using open-circuit impurity removal of electrolyte in zinc smelting process according to claim 1, which is characterized in that: in the step (1), the pH of the neutral leaching end point is controlled to be 4.2-4.4, the leaching rate of zinc in smoke dust is more than 80%, the leaching rate of indium is less than 8%, and the solid ratio of leaching liquid is 4.5:1.
3. the process for jointly recycling zinc and indium by using open-circuit impurity removal of electrolyte in zinc smelting process according to claim 1, which is characterized in that: in the step (2), sodium carbonate is adopted as a precipitator, and the dosage of the sodium carbonate is controlled to be 2.3t/t.Zn, so that the zinc precipitation rate reaches more than 98 percent, and the magnesium entrainment rate is less than 6 percent.
4. The process for jointly recycling zinc and indium by using open-circuit impurity removal of electrolyte in zinc smelting process according to claim 1, which is characterized in that: in the step (2), the obtained basic zinc carbonate has a zinc content of more than 45%, a magnesium content of less than 1.5%, a fluorine content of less than 0.009%, a chlorine content of less than 0.005%, an arsenic content of less than 0.007% and an antimony content of less than 0.002%.
5. The process for jointly recycling zinc and indium by using open-circuit impurity removal of electrolyte in zinc smelting process according to claim 1, which is characterized in that: in the step (2) and the step (4), most fluorine and antimony enter lead silver slag, part of chlorine enters basic zinc carbonate, and arsenic is in the zinc-precipitated liquid.
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