CN119464730A - A method for coordinating copper smelting slag and zinc leaching slag - Google Patents
A method for coordinating copper smelting slag and zinc leaching slag Download PDFInfo
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- CN119464730A CN119464730A CN202411644994.0A CN202411644994A CN119464730A CN 119464730 A CN119464730 A CN 119464730A CN 202411644994 A CN202411644994 A CN 202411644994A CN 119464730 A CN119464730 A CN 119464730A
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- slag
- copper
- zinc
- copper smelting
- smelting slag
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- 239000010949 copper Substances 0.000 title claims abstract description 279
- 239000002893 slag Substances 0.000 title claims abstract description 238
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 208
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 197
- 239000011701 zinc Substances 0.000 title claims abstract description 141
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 111
- 238000003723 Smelting Methods 0.000 title claims abstract description 100
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 97
- 238000000034 method Methods 0.000 title claims abstract description 70
- 238000002386 leaching Methods 0.000 title claims abstract description 39
- 239000011133 lead Substances 0.000 claims abstract description 63
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 claims abstract description 56
- 239000000428 dust Substances 0.000 claims abstract description 39
- 238000005188 flotation Methods 0.000 claims abstract description 37
- 239000000203 mixture Substances 0.000 claims abstract description 36
- 239000011787 zinc oxide Substances 0.000 claims abstract description 31
- 239000003245 coal Substances 0.000 claims abstract description 27
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 26
- 239000012141 concentrate Substances 0.000 claims abstract description 26
- 238000000227 grinding Methods 0.000 claims abstract description 24
- 238000010583 slow cooling Methods 0.000 claims abstract description 21
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims abstract description 9
- 229910052709 silver Inorganic materials 0.000 claims abstract description 9
- 239000004332 silver Substances 0.000 claims abstract description 9
- 238000002156 mixing Methods 0.000 claims abstract description 7
- 230000004907 flux Effects 0.000 claims abstract description 6
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 claims abstract description 5
- 229910052737 gold Inorganic materials 0.000 claims abstract description 5
- 239000010931 gold Substances 0.000 claims abstract description 5
- JQJCSZOEVBFDKO-UHFFFAOYSA-N lead zinc Chemical compound [Zn].[Pb] JQJCSZOEVBFDKO-UHFFFAOYSA-N 0.000 claims abstract 2
- WPSAZVDYGAPZSD-UHFFFAOYSA-N silver;zinc Chemical compound [Zn+2].[Ag+] WPSAZVDYGAPZSD-UHFFFAOYSA-N 0.000 claims abstract 2
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 57
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 claims description 36
- 229910052742 iron Inorganic materials 0.000 claims description 24
- 229910052745 lead Inorganic materials 0.000 claims description 17
- 235000012239 silicon dioxide Nutrition 0.000 claims description 17
- 239000000377 silicon dioxide Substances 0.000 claims description 13
- 229910052717 sulfur Inorganic materials 0.000 claims description 12
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims description 11
- 239000011593 sulfur Substances 0.000 claims description 11
- 229910052683 pyrite Inorganic materials 0.000 claims description 9
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 claims description 9
- 239000011028 pyrite Substances 0.000 claims description 9
- 238000001816 cooling Methods 0.000 claims description 8
- 229910052906 cristobalite Inorganic materials 0.000 claims description 5
- 229910052500 inorganic mineral Inorganic materials 0.000 claims description 5
- 239000011707 mineral Substances 0.000 claims description 5
- 229910052905 tridymite Inorganic materials 0.000 claims description 5
- 229910052681 coesite Inorganic materials 0.000 claims description 4
- 229910052682 stishovite Inorganic materials 0.000 claims description 4
- OMZSGWSJDCOLKM-UHFFFAOYSA-N copper(II) sulfide Chemical compound [S-2].[Cu+2] OMZSGWSJDCOLKM-UHFFFAOYSA-N 0.000 claims description 3
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 claims description 2
- TXKMVPPZCYKFAC-UHFFFAOYSA-N disulfur monoxide Inorganic materials O=S=S TXKMVPPZCYKFAC-UHFFFAOYSA-N 0.000 claims description 2
- XTQHKBHJIVJGKJ-UHFFFAOYSA-N sulfur monoxide Chemical compound S=O XTQHKBHJIVJGKJ-UHFFFAOYSA-N 0.000 claims description 2
- XOCUXOWLYLLJLV-UHFFFAOYSA-N [O].[S] Chemical compound [O].[S] XOCUXOWLYLLJLV-UHFFFAOYSA-N 0.000 claims 1
- 238000011084 recovery Methods 0.000 abstract description 41
- 239000000779 smoke Substances 0.000 abstract description 25
- 238000005265 energy consumption Methods 0.000 abstract description 9
- 238000005272 metallurgy Methods 0.000 abstract description 2
- 230000008569 process Effects 0.000 description 42
- 229910004298 SiO 2 Inorganic materials 0.000 description 32
- 238000006243 chemical reaction Methods 0.000 description 19
- 238000009826 distribution Methods 0.000 description 18
- 239000002994 raw material Substances 0.000 description 17
- 238000009854 hydrometallurgy Methods 0.000 description 16
- 230000000052 comparative effect Effects 0.000 description 13
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 11
- 229910052760 oxygen Inorganic materials 0.000 description 11
- 239000001301 oxygen Substances 0.000 description 11
- 238000010790 dilution Methods 0.000 description 9
- 239000012895 dilution Substances 0.000 description 9
- 230000009467 reduction Effects 0.000 description 9
- 239000002245 particle Substances 0.000 description 8
- 239000010453 quartz Substances 0.000 description 8
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 7
- 239000000463 material Substances 0.000 description 7
- 238000002844 melting Methods 0.000 description 7
- 230000008018 melting Effects 0.000 description 7
- 239000004575 stone Substances 0.000 description 7
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N Iron oxide Chemical compound [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 description 6
- 239000007789 gas Substances 0.000 description 6
- 239000007788 liquid Substances 0.000 description 6
- 238000002360 preparation method Methods 0.000 description 6
- 239000000155 melt Substances 0.000 description 5
- 239000002184 metal Substances 0.000 description 5
- 229910052751 metal Inorganic materials 0.000 description 5
- 230000001698 pyrogenic effect Effects 0.000 description 5
- 239000000243 solution Substances 0.000 description 5
- 239000002918 waste heat Substances 0.000 description 5
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 4
- 238000002485 combustion reaction Methods 0.000 description 4
- AQMRBJNRFUQADD-UHFFFAOYSA-N copper(I) sulfide Chemical compound [S-2].[Cu+].[Cu+] AQMRBJNRFUQADD-UHFFFAOYSA-N 0.000 description 4
- 239000013078 crystal Substances 0.000 description 4
- 239000003546 flue gas Substances 0.000 description 4
- 239000000446 fuel Substances 0.000 description 4
- 150000002739 metals Chemical class 0.000 description 4
- 239000002253 acid Substances 0.000 description 3
- 238000006477 desulfuration reaction Methods 0.000 description 3
- 230000023556 desulfurization Effects 0.000 description 3
- 238000007667 floating Methods 0.000 description 3
- 239000003500 flue dust Substances 0.000 description 3
- 238000004519 manufacturing process Methods 0.000 description 3
- 230000004048 modification Effects 0.000 description 3
- 238000012986 modification Methods 0.000 description 3
- 238000009853 pyrometallurgy Methods 0.000 description 3
- 238000000926 separation method Methods 0.000 description 3
- 239000002699 waste material Substances 0.000 description 3
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 3
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 description 2
- QPLDLSVMHZLSFG-UHFFFAOYSA-N Copper oxide Chemical compound [Cu]=O QPLDLSVMHZLSFG-UHFFFAOYSA-N 0.000 description 2
- 239000005751 Copper oxide Substances 0.000 description 2
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 2
- 230000009286 beneficial effect Effects 0.000 description 2
- 229910052799 carbon Inorganic materials 0.000 description 2
- 229910001779 copper mineral Inorganic materials 0.000 description 2
- BWFPGXWASODCHM-UHFFFAOYSA-N copper monosulfide Chemical compound [Cu]=S BWFPGXWASODCHM-UHFFFAOYSA-N 0.000 description 2
- 229910000431 copper oxide Inorganic materials 0.000 description 2
- 238000007599 discharging Methods 0.000 description 2
- 230000000694 effects Effects 0.000 description 2
- 239000002923 metal particle Substances 0.000 description 2
- 238000004064 recycling Methods 0.000 description 2
- 238000005507 spraying Methods 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- 238000012546 transfer Methods 0.000 description 2
- 239000005749 Copper compound Substances 0.000 description 1
- 229910000640 Fe alloy Inorganic materials 0.000 description 1
- MBMLMWLHJBBADN-UHFFFAOYSA-N Ferrous sulfide Chemical compound [Fe]=S MBMLMWLHJBBADN-UHFFFAOYSA-N 0.000 description 1
- 241001424392 Lucia limbaria Species 0.000 description 1
- BPQQTUXANYXVAA-UHFFFAOYSA-N Orthosilicate Chemical compound [O-][Si]([O-])([O-])[O-] BPQQTUXANYXVAA-UHFFFAOYSA-N 0.000 description 1
- 229910000831 Steel Inorganic materials 0.000 description 1
- 238000007664 blowing Methods 0.000 description 1
- 239000004566 building material Substances 0.000 description 1
- 239000003638 chemical reducing agent Substances 0.000 description 1
- 150000001875 compounds Chemical class 0.000 description 1
- 150000001880 copper compounds Chemical class 0.000 description 1
- TVZPLCNGKSPOJA-UHFFFAOYSA-N copper zinc Chemical compound [Cu].[Zn] TVZPLCNGKSPOJA-UHFFFAOYSA-N 0.000 description 1
- BERDEBHAJNAUOM-UHFFFAOYSA-N copper(I) oxide Inorganic materials [Cu]O[Cu] BERDEBHAJNAUOM-UHFFFAOYSA-N 0.000 description 1
- KRFJLUBVMFXRPN-UHFFFAOYSA-N cuprous oxide Chemical compound [O-2].[Cu+].[Cu+] KRFJLUBVMFXRPN-UHFFFAOYSA-N 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 238000003912 environmental pollution Methods 0.000 description 1
- 230000005484 gravity Effects 0.000 description 1
- 238000010438 heat treatment Methods 0.000 description 1
- 229910052595 hematite Inorganic materials 0.000 description 1
- 239000011019 hematite Substances 0.000 description 1
- 229910052738 indium Inorganic materials 0.000 description 1
- APFVFJFRJDLVQX-UHFFFAOYSA-N indium atom Chemical compound [In] APFVFJFRJDLVQX-UHFFFAOYSA-N 0.000 description 1
- 239000004615 ingredient Substances 0.000 description 1
- 150000002506 iron compounds Chemical class 0.000 description 1
- LIKBJVNGSGBSGK-UHFFFAOYSA-N iron(3+);oxygen(2-) Chemical compound [O-2].[O-2].[O-2].[Fe+3].[Fe+3] LIKBJVNGSGBSGK-UHFFFAOYSA-N 0.000 description 1
- SZVJSHCCFOBDDC-UHFFFAOYSA-N iron(II,III) oxide Inorganic materials O=[Fe]O[Fe]O[Fe]=O SZVJSHCCFOBDDC-UHFFFAOYSA-N 0.000 description 1
- JEIPFZHSYJVQDO-UHFFFAOYSA-N iron(III) oxide Inorganic materials O=[Fe]O[Fe]=O JEIPFZHSYJVQDO-UHFFFAOYSA-N 0.000 description 1
- 238000012544 monitoring process Methods 0.000 description 1
- 239000010665 pine oil Substances 0.000 description 1
- 239000000843 powder Substances 0.000 description 1
- 238000010791 quenching Methods 0.000 description 1
- 230000000171 quenching effect Effects 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 239000007787 solid Substances 0.000 description 1
- 238000007711 solidification Methods 0.000 description 1
- 230000008023 solidification Effects 0.000 description 1
- 239000004071 soot Substances 0.000 description 1
- 239000010959 steel Substances 0.000 description 1
- 238000003756 stirring Methods 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 229910052569 sulfide mineral Inorganic materials 0.000 description 1
- YOXKVLXOLWOQBK-UHFFFAOYSA-N sulfur monoxide zinc Chemical compound [Zn].S=O YOXKVLXOLWOQBK-UHFFFAOYSA-N 0.000 description 1
Classifications
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- 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
- C22B7/04—Working-up slag
-
- 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
- C22B11/02—Obtaining noble metals by dry processes
- C22B11/021—Recovery of noble metals 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
- C22B11/00—Obtaining noble metals
- C22B11/02—Obtaining noble metals by dry processes
- C22B11/021—Recovery of noble metals from waste materials
- C22B11/023—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
- C22B15/00—Obtaining copper
- C22B15/0026—Pyrometallurgy
- C22B15/0028—Smelting or converting
- C22B15/0052—Reduction smelting or converting
-
- 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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/34—Obtaining zinc oxide
-
- 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
- C22B7/001—Dry processes
-
- 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
-
- 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)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
本发明属于熔渣冶金技术领域，涉及一种铜熔炼渣和锌浸出渣的协同处理方法，包括以下步骤：S1、将铜熔炼渣、锌浸出渣、硫化剂和熔剂混合，得到混合料；再通入粉煤和空气的混合物，在1150‑1300℃下反应2‑4h，得到气相、渣相和冰铜；S2、气相进行除尘，得到的富含锌、铅、银的氧化锌烟尘；氧化锌烟尘送湿法浸出锌；S3、渣相经缓冷、磨矿、浮选选铜，得到铜精矿，回收金、银、铜元素。本发明利用高锌热态铜熔炼渣的热量搭配处理高铜常温的湿法炼锌渣，降低了湿法炼锌渣的渣处理能耗。通过混合处理，混合渣中的铅、锌进入次氧化锌烟尘得到回收，铅、锌的回收率均达80％以上；混合渣中的铜进入冰铜回收总回收率大于85％。The present invention belongs to the technical field of slag metallurgy, and relates to a method for the coordinated treatment of copper smelting slag and zinc leaching slag, comprising the following steps: S1, mixing copper smelting slag, zinc leaching slag, a sulfiding agent and a flux to obtain a mixture; then introducing a mixture of pulverized coal and air, reacting at 1150-1300°C for 2-4h, and obtaining a gas phase, a slag phase and matte; S2, dust removal of the gas phase, and zinc oxide smoke dust rich in zinc, lead and silver; zinc oxide smoke dust is sent to wet leaching of zinc; S3, the slag phase is subjected to slow cooling, grinding, and flotation to select copper, and copper concentrate is obtained, and gold, silver and copper elements are recovered. The present invention utilizes the heat of high-zinc hot copper smelting slag to treat high-copper normal-temperature wet zinc smelting slag, thereby reducing the slag treatment energy consumption of wet zinc smelting slag. Through mixed treatment, lead and zinc in the mixed slag enter the secondary zinc oxide smoke dust for recovery, and the recovery rates of lead and zinc are both above 80%; the copper in the mixed slag enters the matte for recovery, and the total recovery rate is greater than 85%.
Description
Technical Field
The invention belongs to the technical field of slag metallurgy, and relates to a cooperative treatment method of copper smelting slag and zinc leaching slag.
Background
The pyrometallurgy copper smelting is a main method of copper smelting, and the pyrometallurgy copper smelting process can be divided into two main types of molten pool smelting and flash smelting. Regardless of the fire copper smelting process, the copper smelting slag contains recoverable copper (0.5-5%) and other associated valuable metals such as zinc, lead, gold and silver. At present, the most main treatment method of copper smelting slag is a hot slag slow cooling-crushing-grinding floating process, has a good recovery effect on copper, gold and silver, and copper in floating tailings can be reduced to below 0.3%, but the recovery rate of zinc and lead cannot be effectively improved. In the pyrometallurgy process, about 70% of zinc in the copper concentrate raw material enters smelting slag, about 50% of lead enters smelting slag, the slag content of the smelting slag is large, the zinc and lead grades are low, and the zinc and lead are mainly in the form of oxides and difficult to recover, so that the zinc recovery rate is generally less than 35% and the lead recovery rate is generally less than 50% in the copper smelting process. For copper concentrate raw materials with high zinc and lead content, such as certain copper concentrates in Arabian countries, the zinc content is up to 4% -8% and the lead content is up to 4%, and how to improve the recovery rate of zinc and lead in the raw materials is an important factor affecting project economy.
Zinc hydrometallurgy is a main method of zinc smelting, and the zinc hydrometallurgy process can be divided into two main types of roasting-leaching and direct leaching. All zinc hydrometallurgy processes produce significant amounts of zinc hydrometallurgy slag. On one hand, the wet zinc smelting slag belongs to dangerous waste and needs to be subjected to pyrogenic harmless treatment, and on the other hand, the wet zinc smelting slag also contains valuable metals such as zinc, lead, copper, silver and the like, has recovery value and can be recovered in the pyrogenic treatment process. At present, the pyrogenic process treatment process of the wet zinc smelting slag comprises a rotary kiln volatilization method, a fuming method, a side-blowing smelting method and the like, and all the pyrogenic process treatment process needs higher energy consumption. In the cremation treatment process, metals such as zinc, lead, silver and the like in the wet zinc smelting slag are volatilized and enriched in the secondary zinc oxide smoke dust, the recovery rate can reach more than 80 percent, more than 90 percent of copper which is not volatilized is lost in the slag, and the recovery rate of copper is less than 10 percent. The copper content of a certain zinc concentrate in Saudi Arabia is 0.8% -1.96%, the copper leaching rate of conventional zinc hydrometallurgy is about 35%, the copper leaching rate of high-temperature high-acid leaching is about 80%, and copper in zinc slag cannot be recovered. Therefore, the comprehensive recovery of the associated valuable metals in the concentrate must be enhanced, and the comprehensive utilization rate of resources is improved.
The prior art CN 106367605B relates to a production method of side-blown depleted copper smelting slag, which comprises the following steps of adding a vulcanizing agent and copper smelting slag into a side-blown depleted furnace from a feed inlet, wherein the mass ratio of the copper smelting slag to the vulcanizing agent is 100:0-15, the copper smelting slag is molten copper smelting slag or solid copper smelting slag, injecting fuel into a molten pool through a nozzle arranged on the side part of the depleted furnace, wherein the fuel is a mixture of pulverized coal and compressed air or a mixture of pulverized coal and oxygen-enriched air, stirring, heat transfer and mass transfer are carried out in the depleted furnace through the fuel, the temperature in the furnace is 1100-1300 ℃, the pressure is 0.1-0.4 Mpa, the depth of the molten pool is 800-2000 mm, copper matte is placed through a copper matte opening, the depleted slag is placed through a slag outlet, flue gas enters a desulfurization system after being cooled, and is discharged after desulfurization standard, and the process has the advantages of low investment, low energy consumption, low cost, high automation degree, good environment and the like. However, this technique does not involve the recovery of zinc from the copper smelting slag, and a large amount of Zn remains in the slag during the treatment, which is not suitable for treating high zinc copper smelting slag.
CN 107723470A relates to a method for producing mixed slag containing copper and iron, which comprises the following steps of S1, mixing slag, adding copper slag into a smelting reaction device, adding one or more of lead smelting slag, blast furnace slag, steel slag and iron alloy slag to form mixed slag, simultaneously adding one or more of copper oxide mineral, copper sulfide mineral and copper-containing material, uniformly mixing, heating the mixed slag to a molten state to serve as reaction slag, monitoring the reaction slag in real time, and obtaining the reacted slag through regulation and control, and S2, separating and recycling. The method can treat hot slag, fully utilize molten copper slag, molten metallurgical slag, physical heat resources and hot metallurgical flux, treat cold slag, realize metallurgical modification of slag through slag mixing or cold mixing, and effectively solve the problems of high-efficiency recycling of metallurgical resources and heat energy and environmental pollution. However, the technology aims at high-iron and high-copper materials to obtain an iron-rich phase and a copper-rich phase, iron enters the iron-rich phase to cause less iron in slag, and a CaO-SiO 2 system is mainly adopted, so that the CaO-SiO 2 binary slag system is adopted, the alkalinity CaO/SiO 2 ratio of reaction slag is controlled, but the smelting temperature of the slag is higher than FeO-SiO 2, and the energy consumption is higher.
In summary, in the traditional copper smelting slag, zinc and lead are basically lost in slag and cannot be recycled in the slow cooling-crushing-grinding floating process of hot slag, while copper is basically lost in slag and cannot be recycled in the single pyrogenic process of traditional zinc hydrometallurgy slag. For the copper smelting slag produced by the high-zinc copper concentrate and the wet zinc smelting slag produced by the high-copper zinc concentrate, the problem of low copper recovery rate in the zinc in the copper smelting slag and the wet zinc smelting slag is needed to be solved.
Disclosure of Invention
The invention aims to provide a cooperative treatment method of copper smelting slag and zinc leaching slag, so as to effectively extract lead and zinc elements in the copper smelting slag and copper elements in the zinc hydrometallurgy slag simultaneously.
In order to achieve the above purpose, the main technical scheme adopted by the invention is as follows:
A cooperative treatment method of copper smelting slag and zinc leaching slag comprises the following steps:
S1, mixing copper smelting slag, zinc leaching slag, a vulcanizing agent and a flux to obtain a mixture, and then introducing a mixture of pulverized coal and air to react for 2-4 hours at 1150-1300 ℃ to obtain gas phase, slag phase and matte;
s2, dust removal is carried out on the gas phase, and zinc oxide smoke dust rich in zinc, lead and silver is obtained;
s3, carrying out slow cooling, ore grinding and copper flotation on the slag phase to obtain copper concentrate, and recovering gold, silver and copper elements.
And slowly cooling, namely separating out cuprous sulfide crystals and metal copper particles in the process of fully cooling sulfide phases in the slag, wherein the slow cooling speed has great influence on the size of copper mineral grains separated out of the slag, for example, most copper-containing grains are less than 5 mu m in water quenching, and the particles are difficult to separate from the slag body, so that the cooling speed needs to be controlled to be beneficial to the growth of copper particles.
Grinding, namely separating sulfide and metal particles from other components by adopting crushing and fine grinding modes of the slag after slow cooling solidification. Typically, it is necessary to grind to a particle size of-0.048 mm to 90% to allow adequate separation.
Copper flotation, namely, copper in slag basically belongs to the property of sulphide ores, and yellow powder, pine oil and the like are selected to float copper-containing substances upwards into flotation copper concentrate.
The process and parameters of slow cooling, ore grinding and copper flotation are referred to the prior art CN202111358752.1.
During the smelting of copper matte (a melt of Cu 2 S and FeS), the stable copper compounds are Cu 2 S and Cu 2 O, the iron compounds are FeO and FeS, and these stable compounds react further with each other or with other components in the feed stock, which has the formula SiO 2+Cu2O+FeS→Cu2S+FeO·SiO2.
The interfacial tension between FeS-Cu 2 S copper sulfur and 2 FeO.SiO 2 melt is about 0.02-0.06N/m, so that the copper sulfur is easy to suspend in slag, and the specific gravity of copper matte and slag phase are greatly different, so that copper matte and slag can be layered in the furnace.
The dust removal comprises flue gas, waste heat utilization of a waste heat boiler, flue gas at an outlet of the waste heat boiler, a surface cooler, a bag type dust remover, a fan and tail gas treatment (refer to the prior art Zhang Fubing. The process research on purifying waste acid and comprehensive utilization of zinc smelting flue gas [ J ]. Sulfuric acid industry 2016, (05): 27-29).
The wet zinc leaching comprises an atmospheric leaching process, a pressure leaching process or an atmospheric-pressure combined leaching process (refer to the prior art 202210614330.4, an oxygen pressure leaching method of indium-containing zinc oxide smoke dust).
In one preferred embodiment, the zinc leaching slag comprises 3-20% of Zn, 0.3-3% of Cu, 1-5% of Pb, 25-35% of Fe and 3-10% of SiO 2.
In one preferred embodiment, the copper smelting slag comprises 0.76% -4.58% of Cu, 40.92% -45.99% of Fe, 25% -35% of SiO 2%, 4% -8% of Zn and 0.5% -3% of Pb.
In one preferred embodiment, the mass ratio of the copper smelting slag to the zinc leaching slag is 4:1-4:3.
In one preferred embodiment, cu is present in the copper smelting slag primarily in the form of Cu 2 S, and secondarily copper oxide and metallic copper.
In one preferred embodiment, the Fe is present in the copper smelting slag primarily in the form of iron silicate, magnetite, and secondarily hematite, iron sulfide.
In one preferred embodiment, the water content of the zinc leaching residue is below 8%.
The zinc leaching slag has high water content, unstable temperature of combustion tools (such as a depletion furnace) and high energy consumption.
In one preferred embodiment, the vulcanizing agent is one or more of sulfur, sulfate, sulfur oxide or sulfur-containing minerals.
In one preferred embodiment, the sulfur-containing mineral is one or more of pyrite, copper sulfide concentrate or oxysulfide zinc ore.
In one preferred embodiment, the mass ratio of Cu to S in the mixture is 3:1-4:1.
If the mass ratio of Cu to S in the mixture is more than 4:1 and the sulfur is insufficient, elemental copper can be generated to enter the matte, so that metallized matte is obtained, even a matte-containing blister copper phase appears, and if the mass ratio of Cu to S in the mixture is less than 3:1, the sulfur is excessive, feS can be generated to enter the matte, the matte grade is low, and the slag quantity is also increased.
In one preferred embodiment, the mass ratio of the copper smelting slag to the vulcanizing agent is 100:4-12.
The mass ratio of the copper smelting slag to the vulcanizing agent is controlled to meet the requirement that the mass ratio of Cu to S of the mixture is 3:1-4:1.
In one preferred embodiment, the flux is a high SiO 2 content material such as quartz, tridymite, cristobalite.
In one preferred embodiment, the iron to silica mass ratio Fe/SiO 2 = 1.6-2.0 in the mix is controlled.
If Fe/SiO 2 <1.6 or >2.0, the temperature in the furnace is required to be higher, and if the temperature is not reached, the slag viscosity increases, and the fluidity of the melt is poor.
The iron is from copper smelting slag and zinc smelting slag, which are not needed to be added, and SiO 2 and SiO 2 are usually needed to be added.
In one of the preferred embodiments, the molar ratio of pulverized coal to oxygen in the mixture of pulverized coal and air is 1:0.5-0.7.
The pulverized coal is mainly C, the reaction is C+O 2＝CO2, and if all the C is reacted, the excess coefficient of oxygen is 1. If O 2 is insufficient, c+co 2 =2co will occur, and a reducing atmosphere will be formed to reduce Zn, pb, and the like.
Pulverized coal is used as fuel and reducing agent, and the following reaction occurs:
2C+O2→2CO;
ZnO·Fe2O3( Liquid and its preparation method )+CO→ZnO( Liquid and its preparation method )+2FeO( Liquid and its preparation method )+CO2;
ZnO( Liquid and its preparation method )+CO→Zn( Air flow )+CO2;
PbO( Liquid and its preparation method )+CO→Pb( Air flow 、 Liquid and its preparation method )+CO2;
In copper smelting slag, part of the lead and zinc are present in the form of oxides and thus reduced.
In one preferred embodiment, the excess air ratio is 0.5 to 0.7.
When the air excess coefficient is 0.5-0.7, the reactions respectively occur as follows:
the excess coefficient is 0.5:C+0.5O 2 →CO
Coefficient of excess 0.6:C+0.6O 2→0.8CO+0.2CO2
Coefficient of excess 0.7:C+0.7O 2→0.6CO+0.4CO2
When the air excess coefficient is less than 0.5, the carbon combustion is insufficient, and as the excess coefficient is increased, the amount of CO generated by the carbon combustion is reduced, the reducing atmosphere is weakened, and the reduction and volatilization of valuable elements such as lead, zinc and the like are not facilitated.
If the oxygen is excessive, sulfur will be oxidized and burned to generate SO 2, after desulfurization, copper matte cannot be produced, and if the oxygen is too little, iron will be reduced to elemental iron, the viscosity is high, the fluidity is poor, and the furnace may be dead when serious.
The reaction temperature in S1 can obviously influence the fluidity, the consumption of pulverized coal and the like is increased due to the excessively high temperature, and the melt fluidity is poor due to the excessively low temperature;
The influence of the reaction time in S1 is that the reaction is incomplete due to insufficient time, and the energy consumption is high due to overlong time.
In one preferred embodiment, the slow cooling is at a cooling rate of no more than 10 ℃ per minute.
Sulfide phases in slag can be separated out into cuprous sulfide crystals and metallic copper particles in a fully slow cooling process, then the cuprous sulfide crystals and the metallic copper particles can be mechanically separated through crushing and fine grinding, sulfide slag concentrate can be produced through flotation by means of the difference of the surface physical and chemical properties of the cuprous sulfide crystals and the metallic copper particles and other slag forming components in slag, and the sulfide slag concentrate is returned to a smelting process, and the copper-containing level of the produced flotation slag tailings is less than 0.3% and is used as waste slag treatment.
The slow cooling rate of the slag has a great influence on the size of the copper mineral grains precipitated in the slag, preferably at <10 ℃ per minute. The cooled slag is ground to separate the sulphide and metal particles from the other components, typically by fine grinding to a particle size of-0.048 mm up to 90%.
The invention is further explained as follows:
The hot copper smelting slag is added into a depletion furnace, meanwhile, a zinc hydrometallurgy slag cold material, a proper amount of vulcanizing agent and flux are added, a mixture of pulverized coal and air is sprayed into a molten pool of the depletion furnace, the depletion furnace is controlled to be in a reducing atmosphere and the temperature of slag, and the slag stays in the furnace for a certain time. And after waste heat is recovered from the depleted furnace gas through a waste heat boiler, the depleted furnace gas enters a dust removing device, zinc, lead and silver volatilized in the depleted furnace gas are enriched in zinc oxide smoke dust products obtained by dust removal, and copper matte generated in the depleted furnace is intermittently discharged through a copper matte discharge port. And (3) slowly cooling and grinding the depleted slag, performing floatation to obtain copper concentrate, returning the obtained floatation copper concentrate to smelting ingredients, and enabling copper content of floatation tailings to be less than 0.3%, wherein the copper concentrate is used as paving materials and building materials for sale.
The beneficial effects of the invention are as follows:
In the copper smelting process, the recovery rate of zinc and lead in copper concentrate raw materials is generally less than 50%, while in the zinc smelting process, the recovery rate of copper is only about 35%, the recovery rate of copper by a high-temperature high-acid method can reach 80%, and the residual copper cannot be recovered even if the residual copper enters slag treatment. The invention utilizes the heat of the high-zinc hot copper smelting slag to match with the high-copper normal-temperature wet zinc smelting slag, and reduces the slag treatment energy consumption of the wet zinc smelting slag. Reducing and volatilizing the mixed slag by a depletion furnace, recovering lead and zinc in the mixed slag into secondary zinc oxide smoke dust, wherein the recovery rate of the lead and the zinc is more than 80%, and the recovery rate of copper in the mixed slag, which is more than 65%, is more than 85% after copper in the mixed slag enters copper matte, and the copper remained in the slag is recovered by a slow cooling-grinding-floatation process.
Drawings
FIG. 1 is a schematic illustration of the treatment process of the present invention.
Detailed Description
In order to make the objects, technical solutions and advantages of the embodiments of the present invention more clear, the technical solutions of the embodiments of the present invention will be clearly and completely described below. It will be apparent that the described embodiments are some, but not all, embodiments of the invention. All other embodiments, which can be made by those skilled in the art based on the embodiments of the invention without making any inventive effort, are intended to be within the scope of the invention.
Example 1:
As shown in fig. 1, 10t/h of hot copper smelting slag (containing Cu 3%, zn 5%, pb 1%, fe 41 and SiO 2%) and 5t/h of zinc hydrometallurgy slag (containing Cu 1%, zn 18%, pb 3%, fe 26% and SiO 2%) are fed into a depletion furnace to be subjected to reduction and volatilization treatment, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:6, the mass ratio of Cu to S in the mixture is controlled to be 3.2:1, quartz stone is added, and the mass ratio of iron to silicon dioxide is controlled to be Fe/SiO 2 =1.6. Spraying a mixture of pulverized coal and air into a melting pool of the depletion furnace, wherein the molar ratio of the pulverized coal to the oxygen is 1:0.55, controlling the depletion furnace to be in a reducing atmosphere, maintaining the temperature at 1250 ℃, and keeping slag in the furnace for 2 hours.
After the reaction, 2.6t of zinc suboxide (containing 0.25% of Cu, 48.5% of Zn and 8.5% of Pb) of the flue dust product of the lean furnace, 0.4t of copper matte at the bottom of the lean furnace (containing 62.12% of Cu, 1.75% of Zn and 0.65% of Pb) and 13.2t of lean slag (containing 0.72% of Cu, 1.0% of Zn and 0.2% of Pb) are obtained. The distribution rates of Cu in the secondary zinc oxide smoke dust, copper matte and slag are respectively 1.86%, 70.99% and 27.15%, the distribution rates of Zn in the secondary zinc oxide smoke dust, copper matte and slag are respectively 90.07%, 0.50% and 9.43%, and the distribution rates of Pb in the secondary zinc oxide smoke dust, copper matte and slag are respectively 88.40%, 1.04% and 10.56%.
The depleted slag is subjected to slow cooling, ore grinding and copper flotation (the process and parameters of slow cooling, ore grinding and copper flotation are referred to the prior art CN 202111358752.1) to obtain copper concentrate 0.28t (Cu 21.02%) and flotation tailings 12.92t (Cu 0.28%), and the recovery rate of copper in flotation is 61.94%.
In the treatment flow, the total recovery rate of Cu, zn and Pb is 88.96%, 89.51% and 88.40%.
Example 2
10T/h copper smelting slag (containing Cu 3%, zn 5%, pb 1%, fe 40.92% and SiO 2%) and 5t/h zinc hydrometallurgy slag (containing Cu 0.9%, zn 5%, pb 1%, fe 25% and SiO 2%) are fed into a depletion furnace for reduction and volatilization treatment, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:10, cu: S=3.5:1 in the mixture is controlled, quartz stone is added, and the mass ratio of iron to silicon dioxide is controlled to Fe/SiO 2 =1.8:1. Spraying a mixture of pulverized coal and air into a melting pool of the depletion furnace, wherein the molar ratio of the pulverized coal to the oxygen is 1:0.65, controlling the depletion furnace to be in a reducing atmosphere, maintaining the temperature at 1280 ℃, and keeping slag in the furnace for 2.5h.
After the reaction, 1.34t of zinc suboxide (containing 0.22% of Cu, 46.1% of Zn and 9.1% of Pb) of the flue dust product of the lean furnace, 0.38t of copper matte at the bottom of the lean furnace (containing 61.41% of Cu, 2.11% of Zn and 1.25% of Pb) and 15.53t of lean slag (containing 0.70% of Cu, 0.80% of Zn and 0.15% of Pb) are obtained. The distribution rates of Cu in the secondary zinc oxide smoke dust, copper matte and slag are respectively 0.85%, 67.64% and 31.51%, the distribution rates of Zn in the secondary zinc oxide smoke dust, copper matte and slag are respectively 82.37%, 1.07% and 16.57%, and the distribution rates of Pb in the secondary zinc oxide smoke dust, copper matte and slag are respectively 81.29%, 3.18% and 15.53%.
The depleted slag is subjected to slow cooling, ore grinding and copper flotation (the process and parameters of slow cooling, ore grinding and copper flotation are referred to the prior art CN 202111358752.1), and 0.29t (Cu 21.20%) of copper concentrate, 15.24t (Cu 0.31%) of flotation tailing slag and 56.54% of copper recovery rate of flotation are obtained.
In the treatment flow, the total recovery rate of Cu, zn and Pb is 85.93%, 82.37% and 81.29%.
Example 3
10T/h copper smelting slag (containing 0.76 percent of Cu, 8 percent of Zn, 3 percent of Pb, 42 percent of Fe and 2 percent of SiO) and 2.5t/h zinc hydrometallurgy slag (containing 3 percent of Cu, 20 percent of Zn, 1 percent of Pb, 35 percent of Fe and 2 percent of SiO) are sent into a depletion furnace for reduction and volatilization treatment, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:12, the mass ratio of Cu to S in the mixture is controlled to be 3.6:1, quartz stone is added, and the mass ratio of iron to silicon dioxide is controlled to be Fe/SiO 2 =1.6. And (3) injecting a mixture of pulverized coal and air into a melting pool of the depletion furnace, wherein the molar ratio of the pulverized coal to the oxygen is 1:0.50, controlling the depletion furnace to be in a reducing atmosphere, maintaining the temperature at 1300 ℃, and keeping slag in the furnace for 2h.
After the reaction, 2.4t of zinc suboxide (containing 0.08% of Cu, 44.90% of Zn and 12.10% of Pb12.10%) of furnace dust products, 0.18t of copper matte (containing 57.34% of Cu, 2.22% of Zn and 1.63% of Pb) of furnace bottom and 10.92t of slag (containing 0.42% of Cu, 2.0% of Zn and 0.29% of Pb) of furnace dust products are obtained. The distribution rates of Cu in the secondary zinc oxide smoke dust, copper matte and slag are respectively 1.27%, 68.35% and 30.37%, the distribution rates of Zn in the secondary zinc oxide smoke dust, copper matte and slag are respectively 82.89%, 0.31% and 16.80%, and the distribution rates of Pb in the secondary zinc oxide smoke dust, copper matte and slag are respectively 89.35%, 0.90% and 9.74%.
The depleted slag is subjected to slow cooling, ore grinding and copper flotation (the process and parameters of slow cooling, ore grinding and copper flotation are referred to the prior art CN 202111358752.1) to obtain copper concentrate 0.1t (Cu 24.22%) and flotation tailings 10.82t (Cu 0.20%), and the recovery rate of copper in flotation is 52.82%.
In the treatment flow, the total recovery rate of Cu, zn and Pb is 85.70%, 82.89% and 89.35%.
Example 4
10T/h copper smelting slag (containing Cu 4.58%, zn 4%, pb 0.5%, fe 41 and SiO 2%) and 5t/h zinc hydrometallurgy slag (containing Cu 0.5%, zn 15%, pb 5%, fe 30% and SiO 2%) are fed into a depletion furnace for reduction and volatilization treatment, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:4, the mass ratio of Cu to S in the mixture is controlled to be 3.5:1, quartz stone is added, and the mass ratio of iron to silicon dioxide is controlled to be Fe/SiO 2 =2. And (3) injecting a mixture of pulverized coal and air into a melting pool of the depletion furnace, wherein the molar ratio of the pulverized coal to the oxygen is 1:0.7, controlling the depletion furnace to be in a reducing atmosphere, maintaining the temperature at 1200 ℃, and keeping slag in the furnace for 4 hours.
After the reaction, 2.2t of zinc suboxide (containing 0.30% of Cu, 45.90% of Zn and 11.70% of Pb11.70%) of furnace dust products of the lean furnace, 0.52t of copper matte at the bottom of the lean furnace (containing 68.28% of Cu, 0.52% of Zn and 0.67% of Pb) and 13.48t of slag (containing 0.90% of Cu, 1.02% of Zn and 0.29% of Pb) are obtained. The distribution rates of Cu in the secondary zinc oxide smoke dust, copper matte and slag are respectively 1.37%, 73.52% and 25.12%, the distribution rates of Zn in the secondary zinc oxide smoke dust, copper matte and slag are respectively 87.81%, 0.24% and 11.96%, and the distribution rates of Pb in the secondary zinc oxide smoke dust, copper matte and slag are respectively 85.80%, 1.17% and 13.03%.
The depleted slag is subjected to slow cooling, ore grinding and copper flotation (the processes and parameters of slow cooling, ore grinding and copper flotation are referred to the prior art CN 202111358752.1) to obtain copper concentrate 0.34t (Cu 24.09%) and flotation tailings 13.14 (Cu 0.30%), and the recovery rate of copper in flotation is 67.51%.
In the treatment flow, the total recovery rate of Cu, zn and Pb is 91.39%, 87.81% and 85.80%.
Example 5
8T/h copper smelting slag (containing Cu 4.58%, zn 4%, pb 0.5%, fe 41 and SiO 2%) and 6t/h zinc hydrometallurgy slag (containing Cu 0.5%, zn 15%, pb 5%, fe 30% and SiO 2%) are fed into a depletion furnace for reduction and volatilization treatment, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:4, the mass ratio of Cu to S in the mixture is controlled to be 3.7:1, quartz stone is added, and the mass ratio of iron to silicon dioxide is controlled to be Fe/SiO 2 =2. And (3) injecting a mixture of pulverized coal and air into a melting pool of the depletion furnace, wherein the molar ratio of the pulverized coal to the oxygen is 1:0.7, controlling the depletion furnace to be in a reducing atmosphere, maintaining the temperature at 1200 ℃, and keeping slag in the furnace for 4 hours.
After the reaction, 2.25t of zinc suboxide (containing 0.28% of Cu, 48.12% of Zn and 13.85% of Pb13) of flue dust products of the dilution furnace, 0.48t of copper matte at the bottom of the dilution furnace (containing 60.10% of Cu, 0.73% of Zn and 0.75% of Pb) and 12.39t of slag (containing 0.82% of Cu, 1.08% of Zn and 0.20% of Pb) of the dilution furnace are obtained. The distribution rates of Cu in the secondary zinc oxide smoke dust, copper matte and slag are respectively 1.59%, 72.78% and 25.63%, the distribution rates of Zn in the secondary zinc oxide smoke dust, copper matte and slag are respectively 88.75%, 0.29% and 10.97%, and the distribution rates of Pb in the secondary zinc oxide smoke dust, copper matte and slag are respectively 91.65%, 1.06% and 7.29%.
The depleted slag is subjected to slow cooling, ore grinding and copper flotation (the process and parameters of slow cooling, ore grinding and copper flotation are referred to the prior art CN 202111358752.1), and 0.34t (Cu 19.25%) of copper concentrate and 12.05t (Cu 0.30%) of flotation tailings are obtained, and the recovery rate of copper flotation is 64.42%.
In the treatment flow, the total recovery rate of Cu, zn and Pb is 90.31%, 88.75% and 91.65%.
Comparative example 1
On the basis of the embodiment 1, copper smelting slag and wet smelting zinc slag are not adopted for cooperative treatment, but are directly and respectively and independently fed into a depletion furnace for reduction and volatilization treatment, and other raw materials and steps are the same as the embodiment 1.
As a result, it was found that the total recovery rates of Zn and Pb were 75.68% and 72.30% in the treatment process and the recovery rates of Pb and Zn were all <80% in the treatment process, and that the copper matte layer could not be obtained at the bottom of the dilution furnace after the reaction due to the small copper content of the material in the wet-process zinc smelting slag, and the copper was dispersed in the slag and recovered in the processes of slow cooling, grinding and copper flotation, and the recovery rate of copper in the whole process was only 64.38%. In addition, the zinc hydrometallurgy slag is treated independently, the temperature is raised and melted by completely relying on external energy, the energy consumption is high, the two materials are separately treated, the equipment quantity is large, and the production cost is high.
Comparative example 2
Based on the embodiment 1, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:2, and the Cu: S=4.2:1 in the mixture is controlled, and other raw materials and steps are the same as those in the embodiment 1.
As a result, it was found that some of the copper in the mix was not bound to sulfur, but was present as elemental copper, most of which entered the matte to obtain metallized matte, and a small amount was entrained in the depleted slag, and the elemental copper could not be recovered during flotation. The recovery rate of copper in the flotation process is 41.54%, the copper content in tailings is 0.43%, and the recovery rate of copper in the whole process is 83.04%.
Comparative example 3
Based on the embodiment 1, pyrite is added as a vulcanizing agent, the mass ratio of the hot copper smelting slag to the vulcanizing agent is 100:14, and the Cu: S=2.8:1 in the mixture is controlled, and other raw materials and steps are the same as those in the embodiment 1. The result shows that the vulcanizing agent is consumed greatly, the smelting slag amount is large, part of iron and sulfur in the mixture are combined to generate FeS which enters the matte, the copper grade in the matte is reduced to 58.40%, and the slag amount in the subsequent converting process is large.
Comparative example 4
On the basis of example 1, quartz stone was added, and the mass ratio Fe/SiO 2 =1.4 of iron to silica was controlled, and other raw materials and steps were the same as in example 1.
As a result, it was found that the mixed slag smelting temperature was controlled at 1410 ℃ and that good fluidity of the slag could be obtained.
Comparative example 5
On the basis of example 1, quartz stone was added, and the mass ratio Fe/SiO 2 =2.1 of iron to silica was controlled, and other raw materials and steps were the same as in example 1.
As a result, it was found that the mixed slag smelting temperature was controlled to 1420 ℃ and that good fluidity of the slag could be obtained.
Comparative example 6
On the basis of example 1, a mixture of pulverized coal and air was injected into the melting bath of the dilution furnace, and the dilution furnace was controlled to have a reducing atmosphere excess air ratio of 0.45, and other raw materials and steps were the same as those in example 1.
As a result, it was found that the pulverized coal was insufficiently combusted and the reducing atmosphere was weak, and zinc-containing sub-zinc oxide fume was 2.5t (Zn44.2%, pb7.76%) and depleted slag 13.3t (Zn1.8%, pb0.35%) were obtained. The distribution rates of Zn and Pb in the secondary zinc oxide smoke dust are 78.93 percent and 77.60 percent respectively, and the volatilization rates of Zn and Pb are low.
Comparative example 7
On the basis of example 1, a mixture of pulverized coal and air was injected into the melting bath of the dilution furnace, and the dilution furnace was controlled to have a reducing atmosphere excess air ratio of 0.75, and other raw materials and steps were the same as those in example 1.
As a result, it was found that the amount of CO produced by the pulverized coal combustion was reduced, the reducing atmosphere was weakened, and zinc-containing secondary zinc oxide soot was 2.45t (Zn43.8%, pb7.5%) and depleted slag 13.35t (Zn1.9%, pb0.4%) were obtained. The distribution rates of Zn and Pb in the secondary zinc oxide smoke dust are 76.65% and 73.5%, respectively, and the volatilization rates of Zn and Pb are reduced.
Comparative example 8
The reaction temperature after passing the mixture of pulverized coal and air was set to 1100℃on the basis of example 1, and other raw materials and steps were the same as in example 1.
As a result, the furnace condition is poor, the melt fluidity is poor, and slag discharging and matte discharging operations are affected.
Comparative example 9
The difference on the basis of example 1 is that the temperature is maintained at 1520 ℃, and other raw materials and steps are the same as in example 1.
As a result, it was found that the melt flowability was good, but the coal rate was increased by 8%, and the energy consumption was large.
Comparative example 10
On the basis of the embodiment 1, the components of the zinc leaching slag are changed, wherein the zinc leaching slag contains 18% of Zn, 3% of Pb, 0.2% of Cu, 26% of Fe, 2% of SiO and other raw materials and steps in the same manner as in the embodiment 1.
As a result, it was found that the copper content in the mixture was reduced, the copper matte layer formed in the dilution furnace was thin, the layering effect of the copper matte and the depleted slag was poor, and the copper distribution ratio in the copper matte was 62.88%. And (3) slowly cooling, grinding and flotation copper separation are carried out on the depleted slag, so that copper concentrate 0.28t (Cu 21.65%) and flotation tailings 12.96t (Cu 0.37%) are obtained, and the recovery rate of copper in flotation is 55.83%. In the treatment process, the total recovery rate of Cu was 83.61%.
Comparative example 11
The composition of the copper smelting slag containing Cu 3%, zn 2%, pb 1%, fe 41%, siO 2% was changed on the basis of example 1, and the other raw materials and steps were the same as in example 1.
As a result, it was found that after the reaction, 1.93t (containing Zn 47.50%, pb 10.75%) of a lean furnace dust product and 13.87t (containing Zn 1.3%, pb 0.22%) of a lean slag were obtained. In the treatment flow, the total recovery rate of Zn and Pb is 83.34 percent and 82.99 percent.
Comparative example 12:
25t/h copper smelting slag (containing Cu 3%, zn 5%, pb 1%, fe 41% and SiO 2%) and 5t/h zinc hydrometallurgy slag (containing Cu 1%, zn 18%, pb 3%, fe 26% and SiO 2% are fed into a depletion furnace for reduction and volatilization treatment, and other raw materials and steps are the same as in example 1.
As a result, it was found that after the reaction, 3.38t of a lean furnace dust product (containing 48.3% of Zn and 9.4% of Pb) and 28.12t of a lean slag (containing 1.80% of Zn and 0.20% of Pb) were obtained. In the treatment flow, the total recovery rate of Zn and Pb is 75.93 percent and 79.43 percent.
Comparative example 13:
10t/h copper smelting slag (containing Cu 3%, zn 5%, pb 1%, fe 41% and SiO 2%) and 10t/h zinc hydrometallurgy slag (containing Cu 1%, zn 18%, pb 3%, fe 26% and SiO 2% are fed into a depletion furnace for reduction and volatilization treatment, and other raw materials and steps are the same as in example 1.
As a result, it was found that 0.42t of matte (wherein Cu was 62.0%) and 17.33t of depleted slag (wherein Cu was 0.75%) were obtained after the reaction. And (3) slowly cooling, grinding and flotation copper separation are carried out on the depleted slag, so that copper concentrate 0.33t (Cu 21.87%) and flotation tailings 17.00t (Cu 0.34%) are obtained, and the recovery rate of copper in flotation is 55.53%. In the treatment process, the total recovery rate of Cu was 84.48%.
The foregoing embodiments are merely for illustrating the technical solution of the present invention, but not for limiting the same, and although the present invention has been described in detail with reference to the foregoing embodiments, it will be understood by those skilled in the art that modifications may be made to the technical solution described in the foregoing embodiments or equivalents may be substituted for parts of the technical features thereof, and such modifications or substitutions may be made without departing from the spirit and scope of the technical solution of the embodiments of the present invention.
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