CN113403480A - Recovery method and application of high-alkali high-zinc blast furnace gas ash - Google Patents
Recovery method and application of high-alkali high-zinc blast furnace gas ash Download PDFInfo
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- CN113403480A CN113403480A CN202110530346.2A CN202110530346A CN113403480A CN 113403480 A CN113403480 A CN 113403480A CN 202110530346 A CN202110530346 A CN 202110530346A CN 113403480 A CN113403480 A CN 113403480A
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- zinc
- blast furnace
- furnace gas
- alkali
- gas ash
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- 239000003513 alkali Substances 0.000 title claims abstract description 55
- 238000000034 method Methods 0.000 title claims abstract description 52
- 239000011701 zinc Substances 0.000 title claims abstract description 51
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 47
- 238000011084 recovery Methods 0.000 title claims abstract description 11
- 238000002386 leaching Methods 0.000 claims abstract description 98
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 claims abstract description 94
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 72
- 239000000843 powder Substances 0.000 claims abstract description 43
- 229910052742 iron Inorganic materials 0.000 claims abstract description 36
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims abstract description 30
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 24
- 239000000919 ceramic Substances 0.000 claims abstract description 18
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 17
- 229910001868 water Inorganic materials 0.000 claims abstract description 16
- 238000007885 magnetic separation Methods 0.000 claims abstract description 11
- 239000012141 concentrate Substances 0.000 claims abstract description 8
- 239000011734 sodium Substances 0.000 claims abstract description 8
- DGAQECJNVWCQMB-PUAWFVPOSA-M Ilexoside XXIX Chemical compound C[C@@H]1CC[C@@]2(CC[C@@]3(C(=CC[C@H]4[C@]3(CC[C@@H]5[C@@]4(CC[C@@H](C5(C)C)OS(=O)(=O)[O-])C)C)[C@@H]2[C@]1(C)O)C)C(=O)O[C@H]6[C@@H]([C@H]([C@@H]([C@H](O6)CO)O)O)O.[Na+] DGAQECJNVWCQMB-PUAWFVPOSA-M 0.000 claims abstract description 7
- ZLMJMSJWJFRBEC-UHFFFAOYSA-N Potassium Chemical compound [K] ZLMJMSJWJFRBEC-UHFFFAOYSA-N 0.000 claims abstract description 7
- 229910052700 potassium Inorganic materials 0.000 claims abstract description 7
- 239000011591 potassium Substances 0.000 claims abstract description 7
- 238000002360 preparation method Methods 0.000 claims abstract description 7
- 229910052708 sodium Inorganic materials 0.000 claims abstract description 7
- 238000001953 recrystallisation Methods 0.000 claims abstract description 6
- 238000000498 ball milling Methods 0.000 claims abstract description 4
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 claims description 72
- 239000000243 solution Substances 0.000 claims description 72
- WCUXLLCKKVVCTQ-UHFFFAOYSA-M Potassium chloride Chemical compound [Cl-].[K+] WCUXLLCKKVVCTQ-UHFFFAOYSA-M 0.000 claims description 60
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 claims description 57
- 239000011787 zinc oxide Substances 0.000 claims description 43
- 238000001704 evaporation Methods 0.000 claims description 41
- 239000011780 sodium chloride Substances 0.000 claims description 36
- 239000001103 potassium chloride Substances 0.000 claims description 30
- 235000011164 potassium chloride Nutrition 0.000 claims description 30
- 229910021529 ammonia Inorganic materials 0.000 claims description 28
- 239000000463 material Substances 0.000 claims description 26
- 230000008020 evaporation Effects 0.000 claims description 24
- 238000006722 reduction reaction Methods 0.000 claims description 24
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims description 21
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 claims description 21
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 21
- 239000001099 ammonium carbonate Substances 0.000 claims description 18
- 150000003839 salts Chemical class 0.000 claims description 18
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical group [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 claims description 17
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 claims description 17
- 239000002904 solvent Substances 0.000 claims description 15
- 238000001238 wet grinding Methods 0.000 claims description 14
- CURLTUGMZLYLDI-UHFFFAOYSA-N Carbon dioxide Chemical compound O=C=O CURLTUGMZLYLDI-UHFFFAOYSA-N 0.000 claims description 13
- 238000001914 filtration Methods 0.000 claims description 13
- 238000002156 mixing Methods 0.000 claims description 13
- IWLXWEWGQZEKGZ-UHFFFAOYSA-N azane;zinc Chemical compound N.[Zn] IWLXWEWGQZEKGZ-UHFFFAOYSA-N 0.000 claims description 12
- 239000013078 crystal Substances 0.000 claims description 12
- LQBJWKCYZGMFEV-UHFFFAOYSA-N lead tin Chemical compound [Sn].[Pb] LQBJWKCYZGMFEV-UHFFFAOYSA-N 0.000 claims description 12
- ATRRKUHOCOJYRX-UHFFFAOYSA-N Ammonium bicarbonate Chemical compound [NH4+].OC([O-])=O ATRRKUHOCOJYRX-UHFFFAOYSA-N 0.000 claims description 11
- 238000010438 heat treatment Methods 0.000 claims description 11
- 239000002893 slag Substances 0.000 claims description 11
- 238000000926 separation method Methods 0.000 claims description 10
- 238000001816 cooling Methods 0.000 claims description 9
- 238000001035 drying Methods 0.000 claims description 9
- 229920006395 saturated elastomer Polymers 0.000 claims description 9
- 239000003638 chemical reducing agent Substances 0.000 claims description 8
- 238000005187 foaming Methods 0.000 claims description 8
- 239000011259 mixed solution Substances 0.000 claims description 8
- 239000000126 substance Substances 0.000 claims description 8
- 238000005406 washing Methods 0.000 claims description 8
- 239000003463 adsorbent Substances 0.000 claims description 7
- 235000012501 ammonium carbonate Nutrition 0.000 claims description 7
- 229910021538 borax Inorganic materials 0.000 claims description 7
- VTYYLEPIZMXCLO-UHFFFAOYSA-L calcium carbonate Substances [Ca+2].[O-]C([O-])=O VTYYLEPIZMXCLO-UHFFFAOYSA-L 0.000 claims description 7
- 229910002092 carbon dioxide Inorganic materials 0.000 claims description 7
- 239000004927 clay Substances 0.000 claims description 7
- 239000000395 magnesium oxide Substances 0.000 claims description 7
- CPLXHLVBOLITMK-UHFFFAOYSA-N magnesium oxide Inorganic materials [Mg]=O CPLXHLVBOLITMK-UHFFFAOYSA-N 0.000 claims description 7
- 229910000069 nitrogen hydride Inorganic materials 0.000 claims description 7
- 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 description 7
- 239000004576 sand Substances 0.000 claims description 7
- 239000000377 silicon dioxide Substances 0.000 claims description 7
- 229910000029 sodium carbonate Inorganic materials 0.000 claims description 7
- 229910001415 sodium ion Inorganic materials 0.000 claims description 7
- 239000004328 sodium tetraborate Substances 0.000 claims description 7
- 235000010339 sodium tetraborate Nutrition 0.000 claims description 7
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 6
- 239000001569 carbon dioxide Substances 0.000 claims description 6
- 238000002425 crystallisation Methods 0.000 claims description 6
- 230000008025 crystallization Effects 0.000 claims description 6
- 229910000013 Ammonium bicarbonate Inorganic materials 0.000 claims description 5
- 239000000654 additive Substances 0.000 claims description 5
- 235000012538 ammonium bicarbonate Nutrition 0.000 claims description 5
- 229910000019 calcium carbonate Inorganic materials 0.000 claims description 5
- 229910010293 ceramic material Inorganic materials 0.000 claims description 5
- 239000007788 liquid Substances 0.000 claims description 5
- 230000000996 additive effect Effects 0.000 claims description 3
- 239000011363 dried mixture Substances 0.000 claims description 3
- 238000011049 filling Methods 0.000 claims description 3
- 238000001179 sorption measurement Methods 0.000 claims description 3
- 239000003245 coal Substances 0.000 claims description 2
- 238000011161 development Methods 0.000 claims description 2
- 239000013589 supplement Substances 0.000 claims description 2
- 230000008569 process Effects 0.000 abstract description 17
- 230000009467 reduction Effects 0.000 abstract description 14
- 229910052751 metal Inorganic materials 0.000 abstract description 7
- 239000002184 metal Substances 0.000 abstract description 6
- 150000002739 metals Chemical class 0.000 abstract description 6
- 229910052783 alkali metal Inorganic materials 0.000 abstract description 4
- 150000001340 alkali metals Chemical class 0.000 abstract description 4
- 230000008901 benefit Effects 0.000 abstract description 4
- 239000011133 lead Substances 0.000 abstract description 3
- 239000007789 gas Substances 0.000 description 54
- 239000002994 raw material Substances 0.000 description 11
- 238000004064 recycling Methods 0.000 description 10
- 239000012535 impurity Substances 0.000 description 9
- 229910000831 Steel Inorganic materials 0.000 description 7
- 239000007864 aqueous solution Substances 0.000 description 7
- 239000000428 dust Substances 0.000 description 7
- 239000010959 steel Substances 0.000 description 7
- 239000005695 Ammonium acetate Substances 0.000 description 6
- 229940043376 ammonium acetate Drugs 0.000 description 6
- 239000002245 particle Substances 0.000 description 6
- 239000000047 product Substances 0.000 description 5
- USFZMSVCRYTOJT-UHFFFAOYSA-N Ammonium acetate Chemical compound N.CC(O)=O USFZMSVCRYTOJT-UHFFFAOYSA-N 0.000 description 4
- ATJFFYVFTNAWJD-UHFFFAOYSA-N Tin Chemical compound [Sn] ATJFFYVFTNAWJD-UHFFFAOYSA-N 0.000 description 4
- 238000010521 absorption reaction Methods 0.000 description 4
- 235000019257 ammonium acetate Nutrition 0.000 description 4
- 238000006243 chemical reaction Methods 0.000 description 4
- 238000000227 grinding Methods 0.000 description 4
- 238000003756 stirring Methods 0.000 description 4
- 239000007832 Na2SO4 Substances 0.000 description 3
- XOLBLPGZBRYERU-UHFFFAOYSA-N SnO2 Inorganic materials O=[Sn]=O XOLBLPGZBRYERU-UHFFFAOYSA-N 0.000 description 3
- PMZURENOXWZQFD-UHFFFAOYSA-L Sodium Sulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=O PMZURENOXWZQFD-UHFFFAOYSA-L 0.000 description 3
- 229910052799 carbon Inorganic materials 0.000 description 3
- 238000001514 detection method Methods 0.000 description 3
- 239000002002 slurry Substances 0.000 description 3
- 229910052938 sodium sulfate Inorganic materials 0.000 description 3
- 239000008399 tap water Substances 0.000 description 3
- 235000020679 tap water Nutrition 0.000 description 3
- LFQSCWFLJHTTHZ-UHFFFAOYSA-N Ethanol Chemical compound CCO LFQSCWFLJHTTHZ-UHFFFAOYSA-N 0.000 description 2
- FMRLDPWIRHBCCC-UHFFFAOYSA-L Zinc carbonate Chemical compound [Zn+2].[O-]C([O-])=O FMRLDPWIRHBCCC-UHFFFAOYSA-L 0.000 description 2
- 238000009835 boiling Methods 0.000 description 2
- 239000000571 coke Substances 0.000 description 2
- 238000004821 distillation Methods 0.000 description 2
- 230000000694 effects Effects 0.000 description 2
- 230000007613 environmental effect Effects 0.000 description 2
- 238000005188 flotation Methods 0.000 description 2
- 239000004615 ingredient Substances 0.000 description 2
- -1 ore Substances 0.000 description 2
- 239000011667 zinc carbonate Substances 0.000 description 2
- QTBSBXVTEAMEQO-UHFFFAOYSA-M Acetate Chemical compound CC([O-])=O QTBSBXVTEAMEQO-UHFFFAOYSA-M 0.000 description 1
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 1
- BVKZGUZCCUSVTD-UHFFFAOYSA-L Carbonate Chemical compound [O-]C([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-L 0.000 description 1
- ZAMOUSCENKQFHK-UHFFFAOYSA-N Chlorine atom Chemical compound [Cl] ZAMOUSCENKQFHK-UHFFFAOYSA-N 0.000 description 1
- 238000003723 Smelting Methods 0.000 description 1
- 229910001128 Sn alloy Inorganic materials 0.000 description 1
- 229910001308 Zinc ferrite Inorganic materials 0.000 description 1
- 230000002411 adverse Effects 0.000 description 1
- 238000005054 agglomeration Methods 0.000 description 1
- 230000002776 aggregation Effects 0.000 description 1
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 1
- KKTCWAXMXADOBB-UHFFFAOYSA-N azanium;hydrogen carbonate;hydrate Chemical compound [NH4+].O.OC([O-])=O KKTCWAXMXADOBB-UHFFFAOYSA-N 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000011449 brick Substances 0.000 description 1
- 238000001354 calcination Methods 0.000 description 1
- 235000010216 calcium carbonate Nutrition 0.000 description 1
- 230000008859 change Effects 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 239000000460 chlorine Substances 0.000 description 1
- 229910052801 chlorine Inorganic materials 0.000 description 1
- 229910052593 corundum Inorganic materials 0.000 description 1
- 239000010431 corundum Substances 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 239000008367 deionised water Substances 0.000 description 1
- 229910021641 deionized water Inorganic materials 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 230000002349 favourable effect Effects 0.000 description 1
- 239000003546 flue gas Substances 0.000 description 1
- 230000004907 flux Effects 0.000 description 1
- 238000009776 industrial production Methods 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 238000005259 measurement Methods 0.000 description 1
- 238000002844 melting Methods 0.000 description 1
- 230000008018 melting Effects 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 230000007935 neutral effect Effects 0.000 description 1
- 239000001301 oxygen Substances 0.000 description 1
- 229910052760 oxygen Inorganic materials 0.000 description 1
- 238000003825 pressing Methods 0.000 description 1
- 238000012545 processing Methods 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 238000004062 sedimentation Methods 0.000 description 1
- 235000011152 sodium sulphate Nutrition 0.000 description 1
- 239000007787 solid Substances 0.000 description 1
- 239000002910 solid waste Substances 0.000 description 1
- 238000000935 solvent evaporation Methods 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 239000011135 tin Substances 0.000 description 1
- WGEATSXPYVGFCC-UHFFFAOYSA-N zinc ferrite Chemical compound O=[Zn].O=[Fe]O[Fe]=O WGEATSXPYVGFCC-UHFFFAOYSA-N 0.000 description 1
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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
- C22B7/02—Working-up flue dust
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03C—MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03C1/00—Magnetic separation
- B03C1/02—Magnetic separation acting directly on the substance being separated
- B03C1/30—Combinations with other devices, not otherwise provided for
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01C—AMMONIA; CYANOGEN; COMPOUNDS THEREOF
- C01C1/00—Ammonia; Compounds thereof
- C01C1/26—Carbonates or bicarbonates of ammonium
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
- C01G9/00—Compounds of zinc
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
- C01G9/00—Compounds of zinc
- C01G9/02—Oxides; Hydroxides
-
- 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
- C22B19/34—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
- C22B25/00—Obtaining tin
- C22B25/04—Obtaining tin by wet 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
- C22B25/00—Obtaining tin
- C22B25/06—Obtaining tin from scrap, especially tin scrap
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- 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
-
- 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
- C22B7/008—Wet processes by an alkaline or ammoniacal leaching
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Abstract
The invention relates to a method for recovering high-alkali high-zinc blast furnace gas ash and application thereof. The method for recovering the high-alkali high-zinc blast furnace gas ash comprises the following steps: the method comprises the steps of dealkalizing blast furnace gas ash by adopting water, separating potassium and sodium elements by recrystallization, dezincifying the blast furnace gas ash by carbothermic reduction, separating zinc oxide powder by adopting a leaching agent, and enriching dezincified ash by ball milling and magnetic separation to obtain iron concentrate powder and iron tailings. The invention uses the recycled iron tailings in the preparation of the foamed ceramics. The recovery method of the high-alkali high-zinc blast furnace gas ash provided by the invention is used for performing targeted treatment on alkali metals, zinc, lead and other low-boiling-point metals in the blast furnace gas ash and simultaneously recovering the alkali metals, the zinc, the lead and other low-boiling-point metals as secondary resources. The method provided by the invention realizes the removal of harmful components in the high-alkali high-zinc blast furnace gas ash and the comprehensive utilization of valuable resources, and has the advantages of simple process, high resource utilization rate, low treatment cost and high comprehensive benefits.
Description
Technical Field
The invention relates to the technical field of resource recycling, in particular to a method for recycling high-alkali high-zinc blast furnace gas ash and application thereof.
Background
The blast furnace gas ash is one of main solid wastes in the steel industry, mainly contains elements such as iron, carbon, zinc and the like, and is an iron-containing zinc-containing secondary resource with extremely high recycling value. In the blast furnace smelting process, impurities of lead, zinc and other metals in iron ore, coke and the like are reduced into steam to volatilize, and salt substances with low melting points are volatilized. The substances, ore, coke, flux and other fine dust are carried out of the blast furnace together with blast furnace gas and collected by a gas dust removal and purification system.
As the largest steel producing country in the world, the amount of blast furnace gas ash produced by the steel industry in China is about 1000 million tons per year. However, since the raw materials used by the respective iron and steel companies have different impurity components, the impurity components contained in the produced blast furnace gas ash are also greatly different. The existing technologies for processing the blast furnace gas ash in large scale mostly adopt the methods of leaching, magnetic separation and flotation or directly using the blast furnace gas ash as a roasting ingredient to process the blast furnace gas ash. However, some steel enterprises such as steelworks produce blast furnace gas ash with high contents of zinc, alkali metals and impurities. The method of magnetic separation, flotation or direct use as roasting ingredients can not recover the high-alkali high-zinc blast furnace gas ash, so that the high-alkali high-zinc blast furnace gas ash is stored in a large amount, and when a large amount of environmental cost is brought to enterprises, important secondary resources are wasted and the environment is polluted. In the existing leaching method, ammonia water-ammonium acetate and the like are mostly used as leaching agents to directly leach blast furnace gas ash. The ammonia water-ammonium acetate is used as a leaching agent, and the zinc-ammonia complex solution obtained by leaching the ammonium acetate is subjected to ammonia distillation, crystallization, washing and roasting to prepare the zinc oxide, so that acetate is decomposed, the generated ammonium acetate cannot be recycled in the leaching process, and the recovery cost of blast furnace gas ash is greatly increased. Further, the direct leaching treatment of blast furnace gas ash with ammonium acetate or the like makes it difficult to separate the bottom low boiling point metals such as lead, tin, etc. from the blast furnace gas ash, which have an extremely adverse effect on the subsequent recovery of iron from the blast furnace gas ash. Furthermore, leaching of blast furnace dust directly with ammonium acetate is difficult to remove the zinc present in the zinc ferrite phase, and some of the zinc is still present in the blast furnace dust.
Disclosure of Invention
Technical problem to be solved
In view of the defects and shortcomings of the prior art, the invention provides a recovery method and application of high-alkali high-zinc blast furnace gas ash, which solve the technical problem that the conventional recovery method of blast furnace gas ash cannot completely recover metal elements such as zinc, potassium, sodium, lead, tin and the like in the blast furnace gas ash.
(II) technical scheme
In order to achieve the purpose, the invention adopts the main technical scheme that:
in a first aspect, an embodiment of the present invention provides a method for recovering blast furnace gas ash containing high alkali and high zinc, including the following steps:
(1) dealkalizing blast furnace gas ash: leaching the blast furnace gas ash by using water, and continuously carrying out solid-liquid separation to obtain an alkali leaching solution and dealkalized ash;
(2) separation of potassium and sodium elements: evaporating and recrystallizing the alkali leaching solution to obtain industrial-grade potassium chloride and crude salt with the main component of sodium chloride;
(3) dezincification treatment of blast furnace gas ash: mixing the dealkalized ash with a carbonaceous reducing agent to carry out carbothermic reduction reaction, and separating a zinc-containing substance from the dealkalized ash to obtain secondary ZnO powder and dezincized ash;
(4) leaching and purifying the secondary ZnO powder in the step (3) by adopting an ammonia-containing leaching agent to dissolve zinc in the secondary ZnO powder in a zinc ammonia complex form into a solution, filtering and separating to obtain a zinc ammonia complex solution and lead-tin slag, performing ammonia evaporation crystallization on the zinc ammonia complex solution to obtain basic zinc carbonate crystals, washing and roasting the basic zinc carbonate crystals to obtain zinc oxide powder;
(5) and (4) ball-milling and magnetic separation enriching the dezincification ash obtained in the step (3) to obtain iron concentrate powder and iron tailings.
In the step (1), deionized water or tap water is adopted to leach out the blast furnace gas ash. Tap water is used in large-scale industrial production. The solid-liquid separation mode comprises centrifugal filtration, sedimentation filtration or plate-and-frame filter pressing. The alkaline leaching solution in the step (1) mainly comprises potassium chloride and sodium chloride.
And (3) carrying out reduction roasting in a rotary hearth furnace or a rotary kiln by using the process equipment used in the carbon reduction dezincification process, and collecting secondary ZnO powder by using a dust collecting chamber. The raw materials can be reduced by powder or briquetted or pelletized.
The carbothermic dezincification process does not need to control the atmosphere and only needs to be carried out in the air environment.
The carbothermic reduction reaction in step (2) mainly takes place as follows:
ZnO+C＝Zn(g)+CO(g)
ZnO+CO(g)＝Zn(g)+CO2(g)
Zn(g)+O2(g)＝ZnO
and (4) separating the zinc-containing substance from the dealkalized ash through the reaction in the step (3), and collecting secondary ZnO powder from the cooling flue gas.
The following reaction mainly occurs in the step (4):
ZnO+nNH3+H2O＝[Zn(NH3)n]2++2OH-
the reaction causes ZnO powder to be dissolved in the solution in the form of a zinc-ammonia complex so as to obtain high-purity zinc oxide. The use of ammonium carbonate or ammonium bicarbonate facilitates the distillation of ammonia to obtainCrystallizing basic zinc carbonate, collecting evaporated ammonia to obtain ammonia water, and calcining CO generated by basic zinc carbonate2Or can be collected and reacted with ammonia water to prepare ammonium carbonate, thereby realizing the recycling of the leaching agent.
And (5) after ball milling of the dezincification ash, selecting iron by using a magnetic separation tube.
Preferably, the field strength used in step (5) is 1000-5000Oe, preferably 1000-2500 Oe.
Preferably, the temperature of the leaching treatment in the step (1) is 20-90 ℃, and preferably 30-70 ℃;
the leaching time is 30-90min, preferably 45-60 min.
Preferably, in step (1): the mass ratio of the water to the blast furnace gas ash is 1-100: 1, preferably 5 to 50: 1, more preferably 10 to 25: 1.
preferably, in the step (2), the alkali leaching solution is subjected to evaporation recrystallization, and the method specifically comprises the following steps:
heating the alkali leaching solution, performing color development adsorption on the alkali leaching solution by using an adsorbent, filtering the adsorbent, and adjusting the pH value of the alkali leaching solution by using dilute hydrochloric acid;
measuring the content of sodium ions in the alkali leaching solution, calculating the total content of sodium chloride in the alkali leaching solution according to the content of the sodium ions, calculating the saturated solvent amount of the sodium chloride in the solution by combining the sodium chloride solubility, and taking the saturated solvent amount of the sodium chloride as an evaporation critical point;
evaporating the alkali leaching solution after the pH is adjusted, stopping evaporation until the evaporation amount of the solvent reaches an evaporation critical point, cooling to separate out potassium chloride crystals, and filtering potassium chloride; and continuously evaporating and crystallizing the residual leaching solution to obtain crude salt with the main component of NaCl.
The solubility of KCl is obviously improved along with the increase of the temperature, the solubility of potassium chloride is reduced when the temperature is reduced, and the influence of the temperature on the solubility of NaCl is small. When the solvent is boiled and evaporated, the solubility of NaCl in the solution is basically the same as that at normal temperature, the KCl solubility is greatly improved, KCl which cannot be dissolved due to the reduction of the solubility can be separated out from the solution by cooling the hot solution, and the solubility change of NaCl is very small due to the reduction of the temperature. Therefore, the total content of sodium chloride in the alkali leaching solution is calculated by measuring the content of sodium ions in the alkali leaching solution, the saturated solvent amount of sodium chloride in the solution is calculated by combining the sodium chloride solubility, and the saturated solvent amount is used as an evaporation critical point, so that the phenomenon that a large amount of NaCl is separated out due to excessive evaporation of the solvent to influence the quality of a KCl product is prevented.
Preferably, the adsorbent is activated carbon;
the addition amount of the adsorbent is 0.5-1.5 g/L
The heating temperature of the alkali leaching solution is 70-95 ℃, and preferably 85-90 ℃;
the time for adsorbing the color substance by the adsorbent is 5-10 min.
Preferably, the KCl crystals are washed with anhydrous ethanol.
Preferably, the separation of the potassium and sodium elements adopts a single-effect or multi-effect evaporator.
And (3) separating KCl and NaCl by adopting a recrystallization method on the alkali leaching solution after the pH is adjusted, wherein the evaporation time depends on the concentration of NaCl in the leaching solution and is 1-8h, and preferably 2-4 h. After recrystallization, purer KCl crystal and crude salt with NaCl as the main component are obtained.
Preferably, in step (2):
adjusting the pH value of the leachate to 6.0-8.5, preferably 6.5-7.5;
the concentration of the dilute hydrochloric acid is 1-3.8mol/L, and preferably 1-2 mol/L.
The pH is adjusted to remove carbonate ions and hydroxide ions, so that the solution is neutral, and the main components of the solution are potassium chloride and sodium chloride.
Preferably, in step (3): the mass ratio of the carbonaceous reducing agent to the dealkalized ash is 1-10: 1, preferably 2-5: 1;
the carbonaceous reducing agent is coke powder or coal powder.
The carbonaceous reducing agent adopted by the invention can ensure that the carbothermic dezincification reduction is smoothly carried out and other harmful impurities are not introduced.
Preferably, in step (3):
the temperature of the reduction reaction is 900-1300 ℃, and preferably 1000-1200 ℃;
the time of the reduction reaction is 10-120 min, preferably 30-90 min.
Preferably, in step (4):
the leaching time of the secondary ZnO powder is 0.5-2h, preferably 1-1.5 h;
the leaching temperature is 25-65 ℃, preferably 25-40 ℃.
Preferably, in step (4):
the leaching agent is a mixed solution of ammonia water and ammonium carbonate or a mixed solution of ammonia water and ammonium bicarbonate;
the concentration of ammonia water in the mixed solution is 4-8mol/L, preferably 6-8 mol/L;
NH in the leaching agent4+：NH3The molar ratio of (A) to (B) is 2-6: 1;
the molar ratio of zinc oxide in the secondary ZnO powder to ammonia water in the leaching agent is 1: 4-6.
And (4) recovering ammonia and carbon dioxide obtained by evaporating gas during ammonia evaporation crystallization in the step (4), and returning the recovered ammonia and carbon dioxide to supplement a leaching agent so as to leach and purify the ZnO powder.
In another aspect, the present invention provides an application of a recycling material of blast furnace gas ash with high alkali and high zinc, wherein iron tailings recycled by the recycling method of blast furnace gas ash with high alkali and high zinc are used for preparing foamed ceramics, and the preparation method of the foamed ceramics comprises the following steps:
takes clay and silica sand as auxiliary materials, and takes MgO, borax and Na2CO3、CaCO3Mixing an additive with the iron tailings to prepare a mixed material;
wet grinding the mixed material, and drying the mixed material after the wet grinding is finished;
and filling the dried mixture into a mold, placing the mold in a heating furnace for roasting, foaming and pore-forming, and preserving the heat of the roasted, foamed and pore-formed material along with the furnace to obtain the foamed ceramic material.
(III) advantageous effects
The invention has the beneficial effects that: the invention relates to a method for preparing high-alkali high-zinc blast furnace gas ashThe recovery method is used for performing targeted treatment on low-boiling-point metals such as alkali metals, zinc, lead and the like in the blast furnace gas ash and simultaneously recovering the metals as secondary resources. The invention adopts carbothermic reduction reaction to separate iron-containing substances in blast furnace gas ash, and takes ammonia water-ammonium carbonate or ammonia water ammonium bicarbonate as a leaching agent to ensure that ZnO powder is completely dissolved in solution in a zinc-ammonia complex form, and zinc oxide is separated from lead and tin to obtain high-purity zinc oxide. The use of ammonium carbonate or ammonium bicarbonate is favorable for obtaining basic zinc carbonate crystals after ammonia evaporation, and simultaneously, the evaporated ammonia can be collected to prepare ammonia water and CO generated by roasting the basic zinc carbonate2Or can be collected and reacted with ammonia water to prepare ammonium carbonate, thereby realizing the recycling of the leaching agent.
The utilization rate of the blast furnace gas ash is high, and the harmless utilization of the blast furnace gas ash is basically completed; the recovery method of the invention enables the iron in the blast furnace gas ash to smoothly return to the blast furnace flow for recycling, reduces the environmental cost and the production cost of iron and steel enterprises, and prepares the foamed ceramic product with high added value by utilizing the residual iron-selecting tail ash.
The method provided by the invention realizes the removal of harmful components in the high-alkali high-zinc blast furnace gas ash and the comprehensive utilization of valuable resources, and has the advantages of simple process, high resource utilization rate, low treatment cost and high comprehensive benefits.
Drawings
FIG. 1 is a flow chart of the method for recovering blast furnace gas ash containing high alkali and high zinc according to the present invention.
FIG. 2 is a flow chart of the present invention for preparing foamed ceramics.
Detailed Description
For a better understanding of the present invention, reference will now be made in detail to the present invention by way of specific embodiments thereof.
The materials, reagents and the like used in the present invention are commercially available unless otherwise specified.
The invention simply crushes the blast furnace gas ash before recycling so as to eliminate the influence of partial moisture absorption agglomeration on the dealkalization process.
In the invention, the blast furnace gas ash is preferably a dust removal product of blast furnace top gas generated in a blast furnace process in a steel enterprise; in the examples of the present invention, the blast furnace gas ash contains about 17.02% of iron, about 9.9% of potassium, about 5.10% of sodium, about 8.63% of zinc, about 17.8% of carbon, about 6.96% of chlorine, and the balance of oxygen as a main component, on a mass basis.
The invention provides a method for recovering high-alkali high-zinc blast furnace gas ash for the first time, which comprises the following steps as shown in figure 1:
(1) dealkalizing blast furnace gas ash: leaching the blast furnace gas ash by using water, and continuously carrying out solid-liquid separation to obtain an alkali leaching solution and dealkalized ash; the leaching treatment specifically comprises: mixing water and blast furnace gas ash according to a mass ratio of 1-100: 1, mixing, stirring for 30-90min at 20-90 ℃, and dissolving; the preferable mass ratio of the water to the blast furnace gas ash is 5-50: 1, more preferably 10 to 25: 1; the stirring temperature and time are further preferably 30-70 ℃ and 45-60 min;
(2) separation of potassium and sodium elements: evaporating and recrystallizing the alkali leaching solution to obtain industrial-grade potassium chloride and crude salt with sodium chloride as a main component; the alkali leaching solution evaporation recrystallization specifically comprises the following steps: heating the alkali leaching solution to 70-95 ℃, preferably 80-90 ℃, adding activated carbon particles into the alkali leaching solution for adsorption for 5-10min, wherein the adding amount of the activated carbon particles is 0.5-1.5 g/L, and adjusting the pH value of the alkali leaching solution to 6.0-8.5 (preferably 6.5-7.5) by using dilute hydrochloric acid with the concentration of 1-3.8mol/L (preferably 1-2 mol/L); measuring the content of sodium ions in the alkali leaching solution, calculating the total content of sodium chloride in the alkali leaching solution according to the content of the sodium ions, calculating the saturated solvent amount of the sodium chloride in the solution by combining the sodium chloride solubility, and taking the saturated solvent amount of the sodium chloride as an evaporation critical point;
evaporating the alkali leaching solution after the pH is adjusted, stopping evaporation until the evaporation amount of the solvent reaches an evaporation critical point, cooling to separate out potassium chloride crystals, and filtering potassium chloride; and continuously evaporating and crystallizing the residual leaching solution to obtain crude salt with the main component of NaCl.
(3) Dezincification treatment of blast furnace gas ash: washing and drying the dealkalized ash obtained in the step (1); mixing a carbonaceous reducing agent and the dried dealkalized ash according to the mass ratio of 1-10: 1 (preferably 2-5: 1), and mixing the prepared materials by using a mixer; carrying out carbothermic reduction reaction by adopting heating equipment such as a muffle furnace and the like at 900-1300 ℃ (preferably 1000-1200 ℃), wherein the reduction reaction time is 10-120 min (preferably 30-90 min), cooling along with the furnace after the reaction is finished, and taking out the dezincification ash and the sublimed sub-ZnO powder at the bottom; the carbothermic reduction reaction can also use a rotary hearth furnace, a rotary kiln and the like to carry out reduction roasting, secondary ZnO powder is collected by a dust collecting chamber, and the raw material can be powder to carry out reduction and can also be briquetted or pelletized to carry out reduction;
(4) adding secondary ZnO powder into an ammonia-containing leaching agent, leaching for 0.5-2h (preferably 1-1.5h) at 25-65 ℃ (preferably 25-40 ℃), continuously adding an ammonium carbonate aqueous solution for multiple times, and performing multiple leaching treatment on ZnO powder to dissolve zinc elements in the secondary ZnO powder in the solution in a zinc ammonia complex form, filtering the solution after leaching is finished, separating to obtain insoluble lead-tin slag and a solution containing a zinc ammonia complex, wherein the lead-tin slag can be used as a secondary resource to prepare a lead-tin alloy; evaporating and crystallizing the solution containing the zinc-ammonia complex to obtain basic zinc carbonate, washing and roasting to obtain zinc oxide powder with higher purity, recovering ammonia and carbon dioxide obtained by evaporating gas during ammonia evaporation and crystallization, and returning the recovered ammonia and carbon dioxide to a leaching agent to obtain secondary ZnO powder.
The leaching agent is a mixed solution of aqueous solution of ammonia water-ammonium carbonate or aqueous solution of ammonia water-ammonium bicarbonate; the concentration of ammonia water in the mixed solution is 4-8mol/L, preferably 6-8 mol/L; NH in leaching agent4+：NH3The molar ratio of (A) to (B) is 2-6: 1; the molar ratio of zinc oxide in the secondary ZnO powder to ammonia water in the leaching agent is 1: 4-6;
(5) wet grinding the dezincification gas ash in the step (3) by using a ball mill, wherein the mass ratio of water to dezincification ash to the grinding balls is 0.6-0.9:1: 2-4; after wet grinding, selecting iron by adopting a magnetic separation tube, wherein the magnetic separation field intensity is 1000-5000Oe, preferably 1000-2500Oe, drying to obtain iron concentrate powder, and the residual substance after iron selection by the magnetic separation tube is iron tailing.
On the other hand, another embodiment of the present invention provides an application of the recovered high-alkali high-zinc blast furnace gas ash, wherein the iron tailings recovered by the above-mentioned recovery method of high-alkali high-zinc blast furnace gas ash are used for preparing foamed ceramics, as shown in fig. 2, the preparation method of the foamed ceramics comprises the following steps:
takes clay and silica sand as auxiliary materials, and takes MgO, borax and Na2CO3、CaCO3Mixing the additive with the iron tailings to prepare a mixed material;
wet grinding the mixed material by using a ball mill, and drying the mixed material after the wet grinding is finished; the mass ratio of water, all raw materials and grinding balls in wet grinding is 0.6-0.9:1: 2-4;
filling the dried mixture into a mold, placing the mold in a heating furnace for roasting, foaming and pore-forming, keeping the temperature of the roasted, foamed and pore-formed material along with the furnace, cooling the material to room temperature, and obtaining high-strength porous ceramic to obtain a foamed ceramic material, wherein the temperature system in the foaming and pore-forming process is as follows: the furnace temperature is increased from room temperature to 800-900 ℃ at the rate of 8-13 ℃/min and is kept for at least 10min, then the temperature is increased to 1000-1200 ℃ at the rate of 3.5-7.5 ℃/min and is kept for 40-90 min, and the roasting, foaming and pore-forming are completed.
Example 1
(1) Adding a blast furnace gas ash raw material into tap water according to the liquid-solid mass ratio of 10:1, leaching for 90 minutes at room temperature, and stirring in the leaching process; and after leaching, carrying out solid-liquid separation to obtain an alkali leaching solution and dealkalized ash.
(2) Heating the alkaline leaching solution to 90 deg.C in water bath, adding activated carbon particles at a ratio of 1g/L, adsorbing color-developing impurities for 10min, filtering, and taking out the activated carbon particles. And (3) adjusting the pH value of the alkaline leaching solution to 7.2 by using dilute hydrochloric acid with the concentration of 2.0 mol/L. Measurement of Na in solution+Concentration, the amount of solvent evaporation of the solution was calculated from the saturated solubility of NaCl. Evaporating and heating the solvent by adopting a boiling evaporation method, stopping heating when the solvent is close to a NaCl saturation point, waiting for the solution to be cooled to room temperature, filtering and separating KCl crystals after cooling, washing the crystals by using absolute ethyl alcohol to obtain a high-purity KCl product, and evaporating the residual solution to dryness to obtain crude salt with the main component of NaCl.
(3) Washing and drying the dealkalized ash, mixing the dried dealkalized ash and the coke powder according to the mass ratio of the coke powder to the dealkalized ash of 10:1, and mixing the materials for 6 hours by using a mixer. And putting the corundum crucible into a muffle furnace for reduction roasting, wherein the thickness of a material layer of the mixture is less than 2 cm. The temperature schedule during the reduction roasting is: the furnace temperature is increased to 1100 ℃ at the speed of 10 ℃/min and is kept for 80min, then the furnace temperature is cooled to 900 ℃ at the speed of 5 ℃/min, and then the furnace temperature is slowly cooled, thus finishing the reduction roasting. And taking out the secondary ZnO powder condensed on the upper edge of the crucible and the dezincification ash at the bottom after cooling.
(4) Putting the sub-ZnO powder generated in the dezincification process into an aqueous solution of ammonia water-ammonium carbonate, and leaching for 1.5h at room temperature, wherein the concentration of the ammonia water in the aqueous solution of the ammonia water-ammonium carbonate is 4mol/L, and NH is contained in the system4+：NH3The molar ratio of the zinc oxide in the secondary ZnO powder to the ammonia water in the leaching agent is 1: 5. and filtering to obtain lead-tin slag with main components of lead and tin after leaching. And (3) evaporating, crystallizing, washing and roasting the solution to finally obtain a ZnO powder product, wherein the recovered and evaporated gas can be returned to the leaching process to be used as a leaching solution for recycling.
(5) And (3) carrying out wet grinding on the dezincification ash by using a ball mill, wherein the mass ratio of water to the dezincification ash to the grinding balls is 0.8:1:4 during the wet grinding. After wet grinding, a magnetic separation tube is adopted for iron separation, and the magnetic separation field intensity is 2000 Oe. And drying to obtain iron concentrate powder.
And respectively detecting the purities of the recovered potassium chloride, crude salt, lead-tin slag and zinc oxide powder, and calculating according to the mass percentage: the purity of the potassium chloride is 96.2 percent; the purity of sodium chloride in the crude salt is 69.7 percent, and the main impurities in the crude salt are KCl (24.1 percent) and Na2SO4(4.0%); the content of PbO in the lead-tin slag is 58.3 percent, and SnO2The content is 38.7%; the purity of the zinc oxide powder was 98%.
Example 2
The embodiment 2 is different from the embodiment 1 in the following points:
leaching for 45min at 90 ℃ in the dealkalization leaching process in the step (1);
the temperature schedule during the reduction roasting in the step (3) is as follows: the furnace temperature is increased from room temperature to 1150 ℃ at the speed of 10 ℃/min, the temperature is kept for 60min, the temperature is increased to 950 ℃ at the speed of 5 ℃/min, and then the furnace is cooled.
The concentration of ammonia water in the ammonia water-ammonium carbonate aqueous solution in the step (4) is 2mol/L, and NH is contained in the system4+：NH3The molar ratio of the zinc oxide in the secondary ZnO powder to the ammonia water in the leaching agent is 1: 4.
otherwise, the other steps are the same as those of example 1.
And respectively detecting the purities of the recovered potassium chloride, crude salt, lead-tin slag and zinc oxide powder, and calculating according to the mass percentage: the purity of the potassium chloride is 96.7 percent; the purity of sodium chloride in the crude salt is 71.3 percent, and the main impurities in the crude salt are KCl (25.1 percent) and Na2SO4(3.6%); the content of PbO in the lead-tin slag is 58.3 percent, and SnO2The content is 38.7%; the purity of the zinc oxide powder was 98.9%.
Example 3
Example 3 differs from example 1 in the following points:
leaching at 60 ℃ for 60min in the dealkalization leaching process in the step (1);
the temperature schedule during the reduction roasting in the step (3) is as follows: the furnace temperature is increased from room temperature to 1200 ℃ at the speed of 10 ℃/min, the temperature is kept for 30min, the temperature is increased to 950 ℃ at the speed of 5 ℃/min, and then the furnace is cooled.
The ammonia water concentration of the ammonia water-ammonium carbonate aqueous solution in the step (4) is 6mol/L, and NH is contained in the system4+：NH3The molar ratio of the zinc oxide in the secondary ZnO powder to the ammonia water in the leaching agent is 1: 6.
otherwise, the other steps are the same as those of example 1.
And respectively detecting the purities of the recovered potassium chloride, crude salt, lead-tin slag and zinc oxide powder, and calculating according to the mass percentage: the purity of the potassium chloride is 97.9 percent; the purity of sodium chloride in the crude salt is 72.6 percent, and the main impurities in the crude salt are KCl (22.0 percent) and Na2SO4(5.4 percent); the content of PbO in the lead-tin slag is 58.3 percent, and SnO2The content is 38.7%; the purity of the zinc oxide powder was 99.1%.
Example 4
The iron ore concentrate powder separated in the example 1 is used for preparing the foamed ceramics, and the preparation method of the foamed ceramics in the example comprises the following steps:
taking the iron tailings obtained in the example 1 as raw materials and silica sand and clay as auxiliary materials, and preparing MgO, borax and Na2CO3And CaCO3 as additives; mixing the raw materials, the auxiliary materials and the additives to prepare a mixed material; the mixed materials comprise the following components in percentage by mass: 50% of iron tailings, 30% of clay, 10% of silica sand, 2.5% of MgO, and CaCO3 1％，Na2CO33.5 percent and 3 percent of borax.
Wet grinding all the raw materials by a ball mill, uniformly mixing the materials and preparing into slurry, wherein the mass ratio of water, all the raw materials and grinding balls in the wet grinding is 0.7:1: 3.
Drying the slurry, stirring in the drying process to ensure the uniformity of the slurry, and preparing the dried raw material into fine powder particles; putting the powder particle raw material into a mould made of refractory bricks, and putting the mould into a medium-temperature box type resistance furnace for roasting and foaming. The temperature system in the foaming and pore-forming process is as follows: the furnace temperature is increased from room temperature to 800 ℃ at the speed of 10 ℃/min and is kept for 10min, then the temperature is increased to 1100 ℃ at the speed of 5 ℃/min and is kept for 55min, and the roasting, foaming and pore-forming are completed.
The performance of the prepared foamed ceramic is detected, and the detection result is as follows:
the bulk density is 0.458g/cm3The heat conductivity coefficient is 0.117W/(m.K), the compressive strength is 4.022MPa, and the volume water absorption is 4.75%.
Example 5
The iron ore concentrate powder separated in example 2 is used for preparing the foamed ceramics, and the preparation method of the foamed ceramics in the present example is different from that in example 4 in that: the foamed ceramic material comprises 49% of iron ore tailings, 32% of clay, 10% of silica sand, 2% of MgO and CaCO by mass percent31.5% of Na2CO32 percent of borax 3.5 percent
The remaining steps were the same as in example 4.
The performance of the prepared foamed ceramic is detected, and the detection result is as follows:
the bulk density is 0.467g/cm3Coefficient of thermal conductivity of 0.11The 8W/(m.K) compressive strength is 4.028MPa, and the volume water absorption is 4.73%.
Example 6
The iron ore concentrate powder separated in example 3 is used for preparing the foamed ceramics, and the preparation method of the foamed ceramics in the present example is different from that in example 4 in that: the foamed ceramic material comprises 53 percent of iron ore tailings, 29 percent of clay, 9 percent of silica sand, 3.0 percent of MgO and CaCO by mass percent31.5% of Na2CO32.0 percent of borax and 2.5 percent of
The remaining steps were the same as in example 4.
The performance of the prepared foamed ceramic is detected, and the detection result is as follows:
the bulk density was 0.451g/cm3The heat conductivity coefficient is 0.119W/(mK), the compressive strength is 4.031MPa, and the volume water absorption is 4.75%.
Finally, it should be noted that: the above embodiments are only used to illustrate the technical solution of the present invention, and not to limit the same; while the invention has been described in detail and with reference to the foregoing embodiments, it will be understood by those skilled in the art that: the technical solutions described in the foregoing embodiments may still be modified, or some or all of the technical features may be equivalently replaced; and the modifications or the substitutions do not make the essence of the corresponding technical solutions depart from the scope of the technical solutions of the embodiments of the present invention.
Claims (10)
1. A method for recovering high-alkali high-zinc blast furnace gas ash is characterized by comprising the following steps:
(1) dealkalizing blast furnace gas ash: leaching the blast furnace gas ash by using water, and continuously carrying out solid-liquid separation to obtain an alkali leaching solution and dealkalized ash;
(2) separation of potassium and sodium elements: evaporating and recrystallizing the alkali leaching solution to obtain industrial-grade potassium chloride and crude salt with the main component of sodium chloride;
(3) dezincification treatment of blast furnace gas ash: mixing the dealkalized ash with a carbonaceous reducing agent to carry out carbothermic reduction reaction, and separating a zinc-containing substance from the dealkalized ash to obtain secondary ZnO powder and dezincized ash;
(4) leaching and purifying the secondary ZnO powder in the step (3) by adopting an ammonia-containing leaching agent, dissolving a zinc element in the secondary ZnO powder in a zinc ammonia complex form into a solution, filtering and separating to obtain a zinc ammonia complex solution and lead-tin slag, performing ammonia evaporation crystallization on the zinc ammonia complex solution to obtain basic zinc carbonate crystals, washing and roasting the basic zinc carbonate crystals to obtain zinc oxide powder;
(5) and (4) ball-milling and magnetic separation enriching the dezincification ash obtained in the step (3) to obtain iron concentrate powder and iron tailings.
2. The method for recovering blast furnace gas ash containing high alkali and high zinc according to claim 1, comprising: the method is characterized in that in the step (1): the temperature of the leaching treatment is 20-90 ℃, and preferably 30-70 ℃;
the leaching time is 30-90min, preferably 45-60 min.
3. The method for recovering blast furnace gas ash containing high soda and high zinc as claimed in claim 1, wherein: in the step (2), the alkali leaching solution is subjected to evaporation recrystallization, and the method specifically comprises the following steps:
heating the alkali leaching solution, performing color development adsorption on the alkali leaching solution by using an adsorbent, filtering the adsorbent, and adjusting the pH value of the alkali leaching solution by using dilute hydrochloric acid;
measuring the content of sodium ions in the alkali leaching solution, calculating the total content of sodium chloride in the alkali leaching solution according to the content of the sodium ions, calculating the saturated solvent amount of the sodium chloride in the solution by combining the sodium chloride solubility, and taking the saturated solvent amount of the sodium chloride as an evaporation critical point;
evaporating the alkali leaching solution after the pH is adjusted, stopping evaporation until the evaporation amount of the solvent reaches an evaporation critical point, cooling to separate out potassium chloride crystals, and filtering potassium chloride; and continuously evaporating and crystallizing the residual leaching solution to obtain crude salt with the main component of NaCl.
4. The method for recovering blast furnace gas ash containing high soda and high zinc as claimed in claim 3, wherein: adjusting pH of the alkaline leaching solution to 6.0-8.5, preferably 6.5-7.5;
the concentration of the dilute hydrochloric acid is 1-3.8mol/L, and preferably 1-2 mol/L.
5. The method for recovering blast furnace gas ash containing high soda and high zinc as claimed in claim 1, wherein: in the step (3): the mass ratio of the carbonaceous reducing agent to the dealkalized ash is 1-10: 1, preferably 2-5: 1;
the carbonaceous reducing agent is coke powder or coal powder.
6. The method for recovering blast furnace gas ash containing high soda and high zinc as claimed in claim 1, wherein: in the step (3):
the temperature of the reduction reaction is 900-1300 ℃, and preferably 1000-1200 ℃;
the time of the reduction reaction is 10-120 min, preferably 30-90 min.
7. The method for recovering blast furnace gas ash containing high soda and high zinc as claimed in claim 1, wherein: in the step (4):
the time for leaching and purifying the secondary ZnO powder is 0.5-2h, preferably 1-1.5 h; the leaching temperature is 25-65 ℃, preferably 25-40 ℃.
8. The method for recovering blast furnace gas ash of high soda and high zinc as set forth in any one of claims 1 to 7, wherein: in the step (4):
the leaching agent is a mixed solution of ammonia water and ammonium carbonate or a mixed solution of ammonia water and ammonium bicarbonate;
the concentration of ammonia water in the mixed solution is 4-8mol/L, preferably 6-8 mol/L;
NH in the leaching agent4+：NH3The molar ratio of (A) to (B) is 2-6: 1;
the molar ratio of zinc oxide in the secondary ZnO powder to ammonia water in the leaching agent is 1: 4-6.
9. The method for recovering blast furnace gas ash containing high soda and high zinc as claimed in claim 8, wherein: and (4) recovering ammonia and carbon dioxide obtained by evaporating gas during ammonia evaporation crystallization in the step (4), and returning the recovered ammonia and carbon dioxide to supplement a leaching agent so as to leach and purify the secondary ZnO powder.
10. The application of the high-alkali high-zinc blast furnace gas ash recyclate is characterized in that: the use of iron tailings recovered from the recovery process of high alkali high zinc blast furnace gas ash according to any one of claims 1 to 9 in the preparation of foamed ceramics comprising the steps of:
takes clay and silica sand as auxiliary materials, and takes MgO, borax and Na2CO3、CaCO3Mixing an additive with the iron tailings to prepare a mixed material;
wet grinding the mixed material, and drying the mixed material after the wet grinding is finished;
and filling the dried mixture into a mold, placing the mold in a heating furnace for roasting, foaming and pore-forming, and preserving the heat of the roasted, foamed and pore-formed material along with the furnace to obtain the foamed ceramic material.
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| CN115927873A (en) * | 2022-11-24 | 2023-04-07 | 东北大学 | Method for preparing zinc oxide by wet processing of secondary zinc oxide smoke dust |
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