CN118047401A - Method for reducing lithium content in lithium-rich electrolyte of aluminum electrolysis and recovering lithium - Google Patents
Method for reducing lithium content in lithium-rich electrolyte of aluminum electrolysis and recovering lithium Download PDFInfo
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- WHXSMMKQMYFTQS-UHFFFAOYSA-N Lithium Chemical compound [Li] WHXSMMKQMYFTQS-UHFFFAOYSA-N 0.000 title claims abstract description 335
- 229910052744 lithium Inorganic materials 0.000 title claims abstract description 335
- 239000003792 electrolyte Substances 0.000 title claims abstract description 143
- 229910052782 aluminium Inorganic materials 0.000 title claims abstract description 132
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 title claims abstract description 130
- 238000005868 electrolysis reaction Methods 0.000 title claims abstract description 85
- 238000000034 method Methods 0.000 title claims abstract description 71
- 238000002386 leaching Methods 0.000 claims abstract description 64
- 239000000654 additive Substances 0.000 claims abstract description 39
- 230000000996 additive effect Effects 0.000 claims abstract description 38
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 37
- 239000012535 impurity Substances 0.000 claims abstract description 19
- 229910003002 lithium salt Inorganic materials 0.000 claims abstract description 16
- 159000000002 lithium salts Chemical class 0.000 claims abstract description 16
- 238000001704 evaporation Methods 0.000 claims abstract description 10
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 9
- 230000001376 precipitating effect Effects 0.000 claims abstract description 9
- 150000003839 salts Chemical class 0.000 claims description 30
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 20
- 229940037003 alum Drugs 0.000 claims description 18
- 239000008367 deionised water Substances 0.000 claims description 15
- 229910021641 deionized water Inorganic materials 0.000 claims description 15
- 239000002253 acid Substances 0.000 claims description 12
- 235000011126 aluminium potassium sulphate Nutrition 0.000 claims description 10
- 238000002156 mixing Methods 0.000 claims description 10
- 229940050271 potassium alum Drugs 0.000 claims description 10
- GRLPQNLYRHEGIJ-UHFFFAOYSA-J potassium aluminium sulfate Chemical compound [Al+3].[K+].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O GRLPQNLYRHEGIJ-UHFFFAOYSA-J 0.000 claims description 10
- 238000004321 preservation Methods 0.000 claims description 9
- 235000011127 sodium aluminium sulphate Nutrition 0.000 claims description 7
- KRHYYFGTRYWZRS-UHFFFAOYSA-M Fluoride anion Chemical compound [F-] KRHYYFGTRYWZRS-UHFFFAOYSA-M 0.000 claims description 6
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 6
- WZUKKIPWIPZMAS-UHFFFAOYSA-K Ammonium alum Chemical compound [NH4+].O.O.O.O.O.O.O.O.O.O.O.O.[Al+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O WZUKKIPWIPZMAS-UHFFFAOYSA-K 0.000 claims description 5
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims description 5
- 235000011124 aluminium ammonium sulphate Nutrition 0.000 claims description 5
- 229910017604 nitric acid Inorganic materials 0.000 claims description 5
- BVKZGUZCCUSVTD-UHFFFAOYSA-L Carbonate Chemical compound [O-]C([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-L 0.000 claims description 4
- 229910019142 PO4 Inorganic materials 0.000 claims description 4
- NBIIXXVUZAFLBC-UHFFFAOYSA-K phosphate Chemical compound [O-]P([O-])([O-])=O NBIIXXVUZAFLBC-UHFFFAOYSA-K 0.000 claims description 4
- 239000010452 phosphate Substances 0.000 claims description 4
- 239000003513 alkali Substances 0.000 claims description 3
- 238000011084 recovery Methods 0.000 abstract description 21
- YCKRFDGAMUMZLT-UHFFFAOYSA-N Fluorine atom Chemical compound [F] YCKRFDGAMUMZLT-UHFFFAOYSA-N 0.000 abstract description 17
- 229910052731 fluorine Inorganic materials 0.000 abstract description 17
- 239000011737 fluorine Substances 0.000 abstract description 17
- PQXKHYXIUOZZFA-UHFFFAOYSA-M lithium fluoride Chemical compound [Li+].[F-] PQXKHYXIUOZZFA-UHFFFAOYSA-M 0.000 abstract description 16
- 238000004064 recycling Methods 0.000 abstract description 6
- INHCSSUBVCNVSK-UHFFFAOYSA-L lithium sulfate Inorganic materials [Li+].[Li+].[O-]S([O-])(=O)=O INHCSSUBVCNVSK-UHFFFAOYSA-L 0.000 abstract description 5
- RBTVSNLYYIMMKS-UHFFFAOYSA-N tert-butyl 3-aminoazetidine-1-carboxylate;hydrochloride Chemical compound Cl.CC(C)(C)OC(=O)N1CC(N)C1 RBTVSNLYYIMMKS-UHFFFAOYSA-N 0.000 abstract description 5
- 230000007613 environmental effect Effects 0.000 abstract description 2
- 239000000243 solution Substances 0.000 description 66
- 239000000047 product Substances 0.000 description 33
- 238000006243 chemical reaction Methods 0.000 description 23
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 18
- 230000000052 comparative effect Effects 0.000 description 15
- IRPGOXJVTQTAAN-UHFFFAOYSA-N 2,2,3,3,3-pentafluoropropanal Chemical compound FC(F)(F)C(F)(F)C=O IRPGOXJVTQTAAN-UHFFFAOYSA-N 0.000 description 12
- KLZUFWVZNOTSEM-UHFFFAOYSA-K Aluminum fluoride Inorganic materials F[Al](F)F KLZUFWVZNOTSEM-UHFFFAOYSA-K 0.000 description 12
- 238000001035 drying Methods 0.000 description 11
- 238000005406 washing Methods 0.000 description 11
- 239000002893 slag Substances 0.000 description 9
- KRHYYFGTRYWZRS-UHFFFAOYSA-N Fluorane Chemical compound F KRHYYFGTRYWZRS-UHFFFAOYSA-N 0.000 description 8
- 230000018044 dehydration Effects 0.000 description 8
- 238000006297 dehydration reaction Methods 0.000 description 8
- 239000002699 waste material Substances 0.000 description 7
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 description 6
- 229910052783 alkali metal Inorganic materials 0.000 description 6
- 150000001340 alkali metals Chemical class 0.000 description 6
- -1 ammonium ions Chemical class 0.000 description 6
- 229910000040 hydrogen fluoride Inorganic materials 0.000 description 6
- 150000002500 ions Chemical class 0.000 description 6
- 229910001386 lithium phosphate Inorganic materials 0.000 description 6
- TWNQGVIAIRXVLR-UHFFFAOYSA-N oxo(oxoalumanyloxy)alumane Chemical compound O=[Al]O[Al]=O TWNQGVIAIRXVLR-UHFFFAOYSA-N 0.000 description 6
- PUZPDOWCWNUUKD-UHFFFAOYSA-M sodium fluoride Chemical compound [F-].[Na+] PUZPDOWCWNUUKD-UHFFFAOYSA-M 0.000 description 6
- TWQULNDIKKJZPH-UHFFFAOYSA-K trilithium;phosphate Chemical compound [Li+].[Li+].[Li+].[O-]P([O-])([O-])=O TWQULNDIKKJZPH-UHFFFAOYSA-K 0.000 description 6
- KWYUFKZDYYNOTN-UHFFFAOYSA-M Potassium hydroxide Chemical compound [OH-].[K+] KWYUFKZDYYNOTN-UHFFFAOYSA-M 0.000 description 5
- 239000007789 gas Substances 0.000 description 5
- 239000000203 mixture Substances 0.000 description 5
- 239000004411 aluminium Substances 0.000 description 4
- 238000001556 precipitation Methods 0.000 description 4
- 238000000926 separation method Methods 0.000 description 4
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 description 3
- HBBGRARXTFLTSG-UHFFFAOYSA-N Lithium ion Chemical compound [Li+] HBBGRARXTFLTSG-UHFFFAOYSA-N 0.000 description 3
- AZDRQVAHHNSJOQ-UHFFFAOYSA-N alumane Chemical class [AlH3] AZDRQVAHHNSJOQ-UHFFFAOYSA-N 0.000 description 3
- PNEYBMLMFCGWSK-UHFFFAOYSA-N aluminium oxide Inorganic materials [O-2].[O-2].[O-2].[Al+3].[Al+3] PNEYBMLMFCGWSK-UHFFFAOYSA-N 0.000 description 3
- DIZPMCHEQGEION-UHFFFAOYSA-H aluminium sulfate (anhydrous) Chemical compound [Al+3].[Al+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O DIZPMCHEQGEION-UHFFFAOYSA-H 0.000 description 3
- 230000009286 beneficial effect Effects 0.000 description 3
- 239000013078 crystal Substances 0.000 description 3
- 230000008020 evaporation Effects 0.000 description 3
- 238000000605 extraction Methods 0.000 description 3
- XGZVUEUWXADBQD-UHFFFAOYSA-L lithium carbonate Chemical compound [Li+].[Li+].[O-]C([O-])=O XGZVUEUWXADBQD-UHFFFAOYSA-L 0.000 description 3
- 229910052808 lithium carbonate Inorganic materials 0.000 description 3
- 229910001416 lithium ion Inorganic materials 0.000 description 3
- 238000004519 manufacturing process Methods 0.000 description 3
- 238000010979 pH adjustment Methods 0.000 description 3
- 239000002994 raw material Substances 0.000 description 3
- 238000007873 sieving Methods 0.000 description 3
- 229910000029 sodium carbonate Inorganic materials 0.000 description 3
- 239000011775 sodium fluoride Substances 0.000 description 3
- 235000013024 sodium fluoride Nutrition 0.000 description 3
- 238000003746 solid phase reaction Methods 0.000 description 3
- 238000012546 transfer Methods 0.000 description 3
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 2
- 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 description 2
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 2
- WNROFYMDJYEPJX-UHFFFAOYSA-K aluminium hydroxide Chemical compound [OH-].[OH-].[OH-].[Al+3] WNROFYMDJYEPJX-UHFFFAOYSA-K 0.000 description 2
- 229910052934 alunite Inorganic materials 0.000 description 2
- 239000010424 alunite Substances 0.000 description 2
- 229910001570 bauxite Inorganic materials 0.000 description 2
- 238000005260 corrosion Methods 0.000 description 2
- 230000007797 corrosion Effects 0.000 description 2
- 238000011161 development Methods 0.000 description 2
- 230000000694 effects Effects 0.000 description 2
- 238000010304 firing Methods 0.000 description 2
- 239000003546 flue gas Substances 0.000 description 2
- 239000000463 material Substances 0.000 description 2
- OTYBMLCTZGSZBG-UHFFFAOYSA-L potassium sulfate Chemical compound [K+].[K+].[O-]S([O-])(=O)=O OTYBMLCTZGSZBG-UHFFFAOYSA-L 0.000 description 2
- 229910052939 potassium sulfate Inorganic materials 0.000 description 2
- 230000029219 regulation of pH Effects 0.000 description 2
- 230000001105 regulatory effect Effects 0.000 description 2
- 238000012216 screening Methods 0.000 description 2
- 239000011734 sodium Substances 0.000 description 2
- 229910052708 sodium Inorganic materials 0.000 description 2
- 235000015424 sodium Nutrition 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- 239000011593 sulfur Substances 0.000 description 2
- 229910052717 sulfur Inorganic materials 0.000 description 2
- 231100000331 toxic Toxicity 0.000 description 2
- 230000002588 toxic effect Effects 0.000 description 2
- VRSRNLHMYUACMN-UHFFFAOYSA-H trilithium;hexafluoroaluminum(3-) Chemical compound [Li+].[Li+].[Li+].[F-].[F-].[F-].[F-].[F-].[F-].[Al+3] VRSRNLHMYUACMN-UHFFFAOYSA-H 0.000 description 2
- XFXPMWWXUTWYJX-UHFFFAOYSA-N Cyanide Chemical compound N#[C-] XFXPMWWXUTWYJX-UHFFFAOYSA-N 0.000 description 1
- ZLMJMSJWJFRBEC-UHFFFAOYSA-N Potassium Chemical compound [K] ZLMJMSJWJFRBEC-UHFFFAOYSA-N 0.000 description 1
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 description 1
- 229910001413 alkali metal ion Inorganic materials 0.000 description 1
- 239000012670 alkaline solution Substances 0.000 description 1
- 150000004645 aluminates Chemical class 0.000 description 1
- CNLWCVNCHLKFHK-UHFFFAOYSA-N aluminum;lithium;dioxido(oxo)silane Chemical compound [Li+].[Al+3].[O-][Si]([O-])=O.[O-][Si]([O-])=O CNLWCVNCHLKFHK-UHFFFAOYSA-N 0.000 description 1
- 239000007864 aqueous solution Substances 0.000 description 1
- 239000012752 auxiliary agent Substances 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 description 1
- 239000000292 calcium oxide Substances 0.000 description 1
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 description 1
- 239000000084 colloidal system Substances 0.000 description 1
- 229910001610 cryolite Inorganic materials 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000004090 dissolution Methods 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 239000003623 enhancer Substances 0.000 description 1
- 238000000227 grinding Methods 0.000 description 1
- 238000000713 high-energy ball milling Methods 0.000 description 1
- 238000003837 high-temperature calcination Methods 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 description 1
- 229910052629 lepidolite Inorganic materials 0.000 description 1
- 239000007788 liquid Substances 0.000 description 1
- 238000012423 maintenance Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 229910052700 potassium Inorganic materials 0.000 description 1
- 239000011591 potassium Substances 0.000 description 1
- 239000002244 precipitate Substances 0.000 description 1
- 230000001737 promoting effect Effects 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 230000000630 rising effect Effects 0.000 description 1
- 239000001488 sodium phosphate Substances 0.000 description 1
- 239000007787 solid Substances 0.000 description 1
- 239000007790 solid phase Substances 0.000 description 1
- 239000002910 solid waste Substances 0.000 description 1
- 229910052642 spodumene Inorganic materials 0.000 description 1
- 238000005979 thermal decomposition reaction Methods 0.000 description 1
- KPZTWMNLAFDTGF-UHFFFAOYSA-D trialuminum;potassium;hexahydroxide;disulfate Chemical compound [OH-].[OH-].[OH-].[OH-].[OH-].[OH-].[Al+3].[Al+3].[Al+3].[K+].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O KPZTWMNLAFDTGF-UHFFFAOYSA-D 0.000 description 1
- RYFMWSXOAZQYPI-UHFFFAOYSA-K trisodium phosphate Chemical compound [Na+].[Na+].[Na+].[O-]P([O-])([O-])=O RYFMWSXOAZQYPI-UHFFFAOYSA-K 0.000 description 1
- 229910000406 trisodium phosphate Inorganic materials 0.000 description 1
- 235000019801 trisodium phosphate Nutrition 0.000 description 1
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to the technical field of recycling of aluminum electrolyte, and discloses a method for reducing lithium content in aluminum electrolysis lithium-rich electrolyte and recycling lithium, wherein the aluminum electrolysis lithium-rich electrolyte and an additive are uniformly mixed and then baked to obtain a baked product A; leaching the roasting product A to obtain a lithium-containing leaching solution B and filter residues C; the filter residue C is washed and dried and then returned to the aluminum electrolysis cell to be continuously used as electrolyte; purifying, removing impurities, evaporating and concentrating the lithium-containing leaching solution B to obtain a lithium-rich solution, and adding a lithium precipitating agent into the lithium-rich solution to obtain a lithium salt product D. The method provided by the invention enables the lithium fluoride which is difficult to dissolve in water in the lithium-rich electrolyte of the aluminum electrolysis to be converted into the lithium sulfate which is easy to dissolve in water, reduces the lithium content in the lithium-rich electrolyte of the aluminum electrolysis, and simultaneously realizes the comprehensive utilization of lithium and fluorine resources in the electrolyte. Meanwhile, the method has the characteristics of high lithium recovery rate, simplicity in operation, environmental friendliness and the like.
Description
Technical Field
The invention relates to the technical field of recycling of aluminum electrolyte, in particular to a method for reducing the lithium content in an aluminum electrolysis lithium-rich electrolyte and recycling lithium.
Background
China is a country with large electrolytic aluminum yield, and the electrolytic aluminum yield is about 60% of the world. Since bauxite in China has low grade and large reserves of aluminum and lithium symbiotic ores, a large amount of lithium-containing bauxite is used for preparing alumina. The lithium-containing alumina is continuously conveyed into an aluminum electrolysis cell for producing electrolytic aluminum, and lithium is continuously enriched in the aluminum electrolyte due to the characteristics of low potential, difficult precipitation and the like in the production process, so that the lithium-rich electrolyte is finally formed (the lithium content can reach more than 1.35 percent). In an aluminum electrolysis molten salt system, the proper addition of lithium fluoride is beneficial to lowering the primary crystal temperature of the electrolyte and improving the conductivity and the current efficiency. However, when the lithium content is high, the electrolysis temperature is lowered, the dissolution capacity of alumina is reduced, and the stability of the electrode is deteriorated. For the aluminium electrolysis cell to operate normally and stably, part of the lithium-rich electrolyte must be replaced with an aluminium electrolyte of lower lithium content after a period of operation of the aluminium electrolysis cell to control the lithium content in the aluminium electrolyte.
In 2022, the yield of electrolytic aluminum in China reaches 4021.4 ten thousand tons, and it is estimated that the amount of the lithium-rich electrolyte generated in aluminum electrolysis is about 100 tons per 1 ten thousand tons of electrolytic aluminum produced, and the amount of the lithium-rich electrolyte is huge. Since the lithium-rich electrolyte contains a large amount of toxic and harmful substances such as fluoride and cyanide, the lithium-rich electrolyte is classified as dangerous solid waste, and the lithium-rich electrolyte must be subjected to harmless treatment. Meanwhile, with the rapid development of new energy automobiles, the demand of lithium salt is rapidly increased, and the lithium salt is gradually becoming a new lithium-containing raw material as a lithium-rich electrolyte with the lithium content equivalent to spodumene and lepidolite, and is highly paid attention to by researchers.
The Chinese patent No. CN109179457B discloses that concentrated sulfuric acid is reacted with electrolytic aluminum waste slag to obtain lithium-containing solution, and the lithium-containing solution is neutralized with calcium oxide to eliminate excessive sulfuric acid, evaporated and concentrated to prepare lithium carbonate. The Chinese patent publication No. CN105293536A, a method for extracting lithium from electrolytic aluminum waste residue, also adopts the reaction of the electrolytic aluminum waste residue and concentrated sulfuric acid at 200-400 ℃, and transfers the lithium in the electrolytic aluminum waste residue into solution for recovery; the Chinese patent No. CN116334410A discloses a method for separating lithium from waste residue of lithium-containing electrolyte in aluminum electrolysis, which adopts concentrated sulfuric acid, a reaction auxiliary agent and deionized water to be ball-milled and mixed with the lithium-containing electrolyte in aluminum electrolysis, then roasting, and leaching with water to obtain a lithium-containing solution, thereby realizing recovery of lithium in the aluminum electrolyte. The above patents can realize the recovery of lithium in the lithium-rich electrolysis of aluminum electrolysis, but the lithium-rich electrolyte contains a large amount of fluorine, so that hydrofluoric acid with extremely strong corrosiveness is easily generated when the lithium-rich electrolyte is treated by a concentrated sulfuric acid method, and the corrosion resistance requirement on equipment is extremely high; in addition, the above-mentioned patent does not fully utilize fluorine resources in the lithium-rich electrolyte.
The Chinese patent No. CN114438329A discloses a comprehensive recovery method of waste lithium-containing aluminum electrolyte, which adopts leaching enhancer to react with lithium-rich electrolyte in acid solution so as to recover lithium transfer leaching liquid in the electrolyte, and the method also has the problem of hydrogen fluoride escaping under the acid condition. The Chinese invention patent No. CN114890447A, a method for directly preparing aluminum fluoride by taking aluminum electrolyte as a raw material without roasting, carries out high-energy ball milling on an aluminum salt additive and the aluminum electrolyte, and the obtained ball grinding material is immersed in water to obtain a lithium-containing solution and leached slag. The Chinese patent publication No. CN116768246A, a method for efficiently separating and enriching lithium from lithium and aluminum in waste residue of aluminum electrolyte, mixes aluminum electrolyte and aluminum sulfate, and then carries out first-stage low-temperature roasting and second-stage high-temperature thermal decomposition, and the obtained roasting material is subjected to lithium sulfate solution obtained by water leaching, so that the recovery of lithium in the residue is realized; however, the process has high firing temperatures, high energy consumption, and additional treatment is required to produce sulfur-containing flue gas.
In summary, the treatment of the lithium-rich electrolyte in aluminum electrolysis in the prior art is easy to generate hydrogen fluoride with strong corrosiveness, has high requirement on corrosion resistance of equipment, and cannot fully utilize valuable resources (such as fluorine resources) in the electrolyte, and fluoride is a main component in an aluminum electrolysis molten salt system. Therefore, the development of a novel technology for reducing the lithium content in the lithium-rich electrolyte of aluminum electrolysis and comprehensively recovering valuable resources (lithium, fluorine, aluminum and the like) in the electrolyte has important significance for realizing comprehensive utilization of aluminum electrolysis byproducts, promoting clean production in the aluminum electrolysis industry and reducing the external dependence of lithium resources in China.
Disclosure of Invention
In view of the above-mentioned shortcomings of the prior art, the present invention aims to provide a method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte. The alum double salt and the aluminum electrolysis lithium-rich electrolyte are uniformly mixed and then baked, so that insoluble lithium fluoride and lithium cryolite in the aluminum electrolysis lithium-rich electrolyte are converted into water-soluble lithium sulfate, and fluorine is converted into water-insoluble aluminum fluoride, thereby realizing effective separation of lithium, aluminum, fluorine and other elements in the aluminum electrolysis lithium-rich electrolyte. In addition, the main components of the filter residue obtained after the roasting product is immersed in water are aluminum fluoride, aluminum oxide and the like, and the filter residue can be returned to the aluminum electrolysis cell for continuous use after being fully washed and dried, so that the comprehensive utilization of aluminum and fluorine resources is realized. And the lithium-containing leaching solution is used for lithium salt products after purification, impurity removal, evaporation and concentration, so that the efficient recovery of lithium is realized.
In order to achieve the above purpose, the invention is implemented according to the following technical scheme:
A method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte comprises the following steps:
S1, uniformly mixing an aluminum electrolysis lithium-rich electrolyte with an additive, and roasting to obtain a roasting product A;
s2, leaching the roasting product A to obtain a lithium-containing leaching solution B and filter residues C; the filter residue C is washed and dried and then returned to the aluminum electrolysis cell to be continuously used as electrolyte;
And S3, purifying, removing impurities, evaporating and concentrating the lithium-containing leaching solution B to obtain a lithium-rich solution, and adding a lithium precipitating agent into the lithium-rich solution to obtain a lithium salt product D.
The lithium-rich electrolyte for aluminum electrolysis is derived from lithium-containing electrolytic slag generated in the aluminum electrolysis process, in particular to electrolytic slag generated by electrolyzing aluminum oxide containing lithium in the north of China; in order to accelerate the reaction rate and make the reaction more sufficient, the lithium-rich electrolyte needs to be crushed and sieved;
preferably, in the step S1, the granularity of the aluminum electrolysis lithium-rich electrolyte is 75-500 mesh.
Further preferably, in the step S1, the granularity of the aluminum electrolysis lithium-rich electrolyte is 100 to 500 mesh.
In order to convert lithium fluoride which is difficult to dissolve in water in the lithium-rich electrolyte of aluminum electrolysis into lithium salt which is easy to dissolve in water, the extraction of lithium in the lithium-rich electrolyte of aluminum electrolysis is further realized. According to the invention, alum double salt is used as an additive to perform a high-temperature solid phase reaction with the aluminum electrolysis lithium-rich electrolyte, so that lithium fluoride is converted into aluminum sulfate dissolved in water, and fluorine in the aluminum electrolysis lithium-rich electrolyte is combined with aluminum in the alum double salt to generate aluminum fluoride which is more stable in thermodynamics and is difficult to dissolve in water, thereby realizing separation of lithium, aluminum, fluorine and other elements. Meanwhile, sodium, potassium, ammonium and other ions in the alum double salt are easy to act in the same way as lithium ions in the lithium-rich electrolyte in the high-temperature solid-phase reaction process, for example, sodium and lithium are similar to generate sodium fluoride, and the reaction of the sodium fluoride and aluminum in the alum double salt has a larger equilibrium constant, namely, the reaction can be easier to occur, so that the efficient extraction of lithium is realized. In addition, alum double salt is mainly sulfuric acid double salt of aluminum and alkali metal or ammonium, after the aluminum double salt is used as an additive to be roasted and immersed in water with aluminum electrolysis lithium-rich electrolyte, alkali metal or ammonium ions enter aqueous solution, and only aluminum exists in slag in the form of fluoride, namely, leaching slag returned to an electrolytic tank cannot introduce other impurity ions.
Preferably, in the step S1, the additive is alum double salt.
Further preferably, in the step S1, the alum double salt is a sulfuric acid double salt of aluminum and alkali metal or ammonium.
Preferably, in the step S1, the alum double salt is at least one of potassium alum, ammonium alum and sodium alum.
In order to sufficiently convert the poorly soluble lithium salt in the aluminum electrolysis lithium-rich electrolyte into a readily water-soluble lithium salt, it is necessary to control the addition amount of the additive. The addition amount of the additive is too small, and insoluble components such as lithium fluoride in the lithium-rich electrolyte are not thoroughly converted; too much additive is added, so that the impurity content in the leaching solution obtained after roasting is higher, and the burden is added to the subsequent process for preparing lithium carbonate.
Preferably, in the step S1, the molar ratio of the lithium-rich electrolyte to the additive for aluminum electrolysis is 1-3:1, based on the lithium content in the lithium-rich electrolyte and the aluminum content in the additive.
The original aluminum fluoride, cryolite, aluminum fluoride generated by subsequent reaction and the like of the aluminum electrolysis lithium-rich electrolyte are easy to react with water vapor at high temperature to generate aluminum oxide and hydrogen fluoride gas, and alum double salt generally has crystal water. In order to ensure that fluorine in the lithium-rich electrolyte is recovered as aluminum fluoride rather than hydrogen fluoride gas being generated and present in the tail gas, a certain temperature raising program is required to be adopted in the roasting process to sufficiently remove crystal water in alum double salts. Therefore, the invention adopts two-stage heating program, namely, the temperature is kept for a period of time at low temperature to fully remove water, and then the reaction is carried out at high temperature to realize the conversion of indissolvable lithium salt into water-soluble lithium salt; meanwhile, the high-temperature reaction stage should avoid the adoption of excessive reaction temperature so as to prevent the generation of sulfur-containing flue gas.
Preferably, in the step S1, the roasting process is that after heat preservation is carried out for 1-2 hours at 200-300 ℃, the temperature is raised to 400-750 ℃ and then the heat preservation is carried out for 0.5-12 hours.
Specifically, the step S1 includes the following steps: uniformly mixing crushed and screened aluminum electrolysis lithium-rich electrolyte with an additive according to a certain proportion, and roasting the mixture in an air atmosphere according to a set temperature-raising program to obtain a roasting product A; the aluminum electrolysis lithium-rich electrolyte is screened by a screen mesh with 75-500 meshes; the additive is alum double salt, preferably at least one of potassium alum, ammonium alum, sodium alum and the like; the molar ratio of the lithium content in the lithium-rich electrolyte to the aluminum content in the additive is 1:1-3:1; the temperature-raising procedure of the roasting reaction is to firstly carry out low-temperature dehydration at 200-300 ℃ for 1-2 h, and then continuously raise the temperature to 400-750 ℃ for 0.5-12 h to carry out solid-phase reaction.
The main components of a roasting product A obtained after the high-temperature solid-phase roasting reaction of the aluminum electrolysis lithium-rich electrolyte and the additive are lithium sulfate, aluminum fluoride, aluminum oxide and the like, and the solubility of the products in water or dilute acid is different to realize the effective separation of lithium and other components; the filter residue obtained after leaching is basically free of lithium aluminum fluoride, aluminum oxide and the like, and is the main component of an aluminum electrolysis molten salt system, and the filter residue can be fully washed by deionized water and dried and then returned to an aluminum electrolysis cell to be used as electrolyte.
Preferably, in the step S2, deionized water or acid is used in the leaching process; the acid is one of dilute sulfuric acid, dilute hydrochloric acid and dilute nitric acid.
Specifically, the liquid-solid ratio in the leaching process is 2-5:1.
Preferably, in the step S2, the concentration of the acid is 0.1 to 1mol/L.
In order to allow sufficient transfer of lithium from the calcined product a to the solution and thus to facilitate subsequent recovery, it is necessary to control the reaction conditions during leaching.
Preferably, in the step S2, the leaching temperature is 20-95 ℃ and the leaching time is 0.5-6 h.
Specifically, the step S2 includes the following steps: leaching the roasting product A by deionized water or dilute acid to obtain a lithium-containing leaching solution B and filter residues C, fully washing the filter residues C by the deionized water, drying, and returning the filter residues C to the aluminum electrolysis cell to continue to serve as electrolyte; the dilute acid is one of dilute sulfuric acid, dilute hydrochloric acid and dilute nitric acid, and the concentration of the dilute acid is 0.1-1 mol/L; the leaching temperature in the leaching process is 20-95 ℃ and the leaching time is 0.5-6 h.
Since the lithium-containing leaching solution B contains a small amount of impurity ions such as aluminum, the impurity ions are easy to co-precipitate with lithium and enter a lithium salt product when lithium is recovered from the leaching solution. Therefore, the lithium-containing leaching solution B needs to be purified and decontaminated. In view of the fact that aluminum hydroxide is colloid, is difficult to filter, and is easy to adsorb lithium ions, so that lithium is lost, aluminum removal cannot be achieved through a fractional precipitation method. The invention utilizes the characteristic that aluminum hydroxide belongs to amphoteric hydroxide, namely aluminum exists in the form of aluminate ions under alkaline condition, and aluminate does not react with carbonate, phosphate and the like in the subsequent lithium precipitation process, so that the separation of impurity ions such as lithium and aluminum in the leaching solution is realized.
Preferably, in the step S3, the process of purifying and removing impurities is to add an alkaline solution to the leaching solution B containing lithium, and adjust the pH value to be greater than 10.
Preferably, the alkali solution is at least one of sodium hydroxide solution and potassium hydroxide solution.
Because lithium salts such as lithium carbonate and lithium phosphate have the property of being slightly soluble in water, the lithium concentration needs to be controlled to be higher in the lithium precipitation process of the lithium-containing leaching solution so as to improve the recovery rate of lithium. Therefore, it is necessary to perform appropriate evaporation concentration of the lithium-containing leachate to obtain a lithium-rich solution having a higher lithium concentration.
Preferably, in the step S3, the lithium concentration of the lithium-rich solution is more than 10g/L.
In order to recover lithium in solution as a lithium salt, a lithium precipitating agent is used to precipitate lithium from the lithium-rich solution. The lithium precipitating agent is lithium salt which is difficult to dissolve in water and is formed with lithium ions.
Preferably, in the step S3, the lithium precipitating agent is at least one of soluble fluoride, carbonate and phosphate.
Specifically, the step S3 includes the following steps: purifying, removing impurities, evaporating and concentrating the lithium-containing leaching solution B to obtain a lithium-rich solution, and then adding a lithium precipitating agent into the lithium-rich solution to obtain a lithium salt product D. The purifying and impurity removing mode is to add alkali solution such as sodium hydroxide, potassium hydroxide and the like into the lithium-containing leaching solution B to adjust the pH value to be more than 10; the lithium concentration of the lithium-rich solution obtained by evaporation concentration is more than 10g/L; the lithium precipitating agent can be at least one of soluble fluoride, carbonate and phosphate.
The main chemical reactions involved in the method for reducing the lithium content and recycling lithium in the lithium-rich electrolyte of aluminum electrolysis provided by the invention are as follows (potassium alum is taken as an example):
6LiF+2KAl(SO4)2·12H2O＝3Li2SO4+K2SO4+2AlF3+24H2O;
2Li3AlF6+2KAl(SO4)2·12H2O＝3Li2SO4+K2SO4+4AlF3+24H2O.
the beneficial effects are that:
1) The alum double salt is adopted to react with the aluminum electrolysis lithium-rich electrolyte, so that lithium fluoride, lithium cryolite and the like which are difficult to dissolve in water in the aluminum electrolysis lithium-rich electrolyte are converted into lithium sulfate which is easy to dissolve in water, fluorine is converted into aluminum fluoride which is one of important components in an electrolytic aluminum molten salt system, the lithium content in the aluminum electrolysis lithium-rich electrolyte is reduced, and the comprehensive utilization of lithium and fluorine resources in the aluminum electrolysis lithium-rich electrolyte is realized;
2) The alunite compound salt is mainly sulfate of alkali metal and aluminum, the existence of the alkali metal can act like the lithium generation substance, the conversion efficiency of the lithium is improved, the alkali metal ions of the obtained product are transferred into the solution after being leached by water, other impurity ions are not introduced in the process of recovering the lithium, the main components of the leaching slag are aluminum fluoride, aluminum oxide and the like, and the leaching slag can be returned to the aluminum electrolysis cell for continuous use;
3) By adopting a two-stage temperature rising procedure of low-temperature dehydration and high-temperature roasting, fluorine in the lithium-rich electrolyte is not generated into toxic hydrogen fluoride gas in the process of roasting reaction of the lithium-rich electrolyte in aluminum electrolysis, but aluminum fluoride is generated by the fluorine-rich electrolyte and aluminum in alum double salt, so that fluorine and aluminum resources are recovered to the maximum extent;
4) The method provided by the invention has the advantages of high recovery rate, simple operation, environmental friendliness and low maintenance cost, and has important practical significance for the treatment of the lithium-rich electrolyte of aluminum electrolysis and the recycling of valuable resources.
Drawings
FIG. 1 is a process flow diagram of the present invention;
Fig. 2 is a graph showing comparison of lithium recovery rates of inventive examples 1, 2, 3, 4,5 and comparative examples 1, 2.
Detailed Description
The present invention will be described in further detail with reference to specific examples. It should be understood that the specific embodiments described herein are for purposes of illustration only and are not intended to limit the scope of the invention.
All the raw materials of the present invention are not particularly limited in purity, and the present invention preferably adopts industrial purity or conventional purity used in the art.
The apparatus used in the present invention is not particularly limited, and any apparatus commonly used in the art is employed.
TABLE 1 main components of lithium-rich electrolyte for aluminum electrolysis
| Composition of the components | Li | Al | F | Na | Ca | K |
| Content of% | 2.17 | 18.96 | 50.03 | 13.26 | 2.31 | 1.98 |
Example 1
The embodiment provides a method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte, wherein a principle process flow chart is shown in fig. 1, specifically potassium alum is used as an additive and deionized water is used for leaching, and the method comprises the following steps:
(1) Uniformly mixing 100g of aluminum electrolysis lithium-rich electrolyte (the main components are shown in table 1) obtained by sieving a 100-mesh screen with 147.03g of additive (potassium alum) according to the molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive being 1:1, then carrying out heat preservation at 300 ℃ for 1h for dehydration, and then carrying out reaction at 750 ℃ for 0.5h to obtain 180.14g of baked product A;
(2) Leaching the roasting product A with 500mL of deionized water at 90 ℃ for 0.5h to obtain a lithium-containing leaching solution B with the lithium concentration of 4.28g/L and 500mL and 106.20g of filter residue C, fully washing and drying the filter residue C by 5L of deionized water, and returning the filter residue C to an aluminum electrolysis cell to continue to serve as molten salt electrolyte in the aluminum electrolysis process;
(3) And (3) dropwise adding sodium hydroxide solution into the lithium-containing leaching solution B to adjust the pH value to 11.32 so as to purify and remove impurities, evaporating and concentrating the solution after the pH adjustment to obtain 203mL of lithium-rich solution with the lithium concentration of 10.56g/L, and adding trisodium phosphate into the lithium-rich solution to obtain 11.71g of lithium phosphate product D.
And (3) analyzing the lithium content in the filter residue C obtained after washing and drying, wherein compared with the lithium content in the initial lithium-rich electrolyte, the lithium content in the filter residue C is reduced to 0.03%, the lithium content in the electrolyte is effectively reduced, and the recovery rate of lithium in the lithium-rich electrolyte is calculated to be as high as 98.53%.
Example 2
The embodiment provides a method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte, specifically adopting sodium alum as an additive and adopting dilute hydrochloric acid for leaching, comprising the following steps:
(1) Uniformly mixing 500g of aluminum electrolysis lithium-rich electrolyte (the main components of which are shown in table 1) obtained by sieving a 500-mesh screen with 236.63g of additive (sodium alum) according to the molar ratio of the lithium content in the lithium-rich electrolyte to the aluminum content in the additive being 3:1, then carrying out heat preservation at 250 ℃ for 2h for dehydration, and then carrying out reaction at 600 ℃ for 6h to obtain 875.09g of baked product A;
(2) Leaching the roasting product A with 2000mL of 1mol/L hydrochloric acid at 20 ℃ for 3 hours to obtain a lithium-containing leaching solution B with a lithium concentration of 5.23g/L and 529.87g of filter residue C, fully washing and drying the filter residue C by 10L of deionized water, and returning the filter residue C to an aluminum electrolysis cell to continue to serve as molten salt electrolyte in the aluminum electrolysis process;
(3) The pH of the lithium-containing leaching solution B is regulated to 12.49 by dropwise adding sodium hydroxide solution to purify and remove impurities, then the solution after the pH regulation is evaporated and concentrated to obtain 490mL of lithium-rich solution with the lithium concentration of 21.35g/L, and then sodium carbonate is added into the lithium-rich solution to obtain 51.09g of lithium phosphate product D.
And (3) analyzing the lithium content in the filter residue C obtained after washing and drying, wherein the lithium content in the filter residue C is reduced to 0.08% compared with the lithium content in the initial lithium-rich electrolyte by 2.17%, so that the lithium content in the electrolyte is effectively reduced, and the recovery rate of lithium in the lithium-rich electrolyte is calculated to be as high as 96.08%.
Example 3
The embodiment provides a method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte, which specifically adopts ammonium alum as an additive and adopts dilute sulfuric acid for leaching, and comprises the following steps:
(1) Uniformly mixing 200g of aluminum electrolysis lithium-rich electrolyte (the main components of which are shown in table 1) obtained by sieving with a 75-mesh sieve and 181.87g of additive (ammonium alum) according to the molar ratio of the lithium content in the lithium-rich electrolyte to the aluminum content in the additive being 2:1, then carrying out heat preservation at 200 ℃ for 1.5h for dehydration, and then carrying out reaction at 400 ℃ for 12h to obtain 292.59g of baked product A;
(2) Leaching the roasting product A with 1000mL of 0.1mol/L dilute sulfuric acid solution at 50 ℃ for 6 hours to obtain a lithium-containing leaching solution B with the lithium concentration of 4.24g/L and 212.38g of filter residue C, fully washing and drying the filter residue C by 8L of deionized water, and returning the filter residue C to an aluminum electrolysis cell to continue to serve as molten salt electrolyte in the aluminum electrolysis process;
(3) And (3) dropwise adding sodium hydroxide solution into the lithium-containing leaching solution B to adjust the pH to 10.49 so as to purify and remove impurities, evaporating and concentrating the solution after the pH adjustment to obtain 250mL of lithium-rich solution with the lithium concentration of 16.93g/L, and adding sodium fluoride into the lithium-rich solution to obtain 15.09g of lithium phosphate product D.
And (3) analyzing the lithium content in the filter residue C obtained after washing and drying, wherein compared with the lithium content in the initial lithium-rich electrolyte, the lithium content in the filter residue C is reduced to 0.05%, the lithium content in the electrolyte is effectively reduced, and the recovery rate of lithium in the lithium-rich electrolyte is calculated to be as high as 97.55%.
Example 4
The embodiment provides a method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte, which specifically adopts potassium alum as an additive and adopts dilute nitric acid leaching, and comprises the following steps:
(1) Uniformly mixing 300g of aluminum electrolysis lithium-rich electrolyte (the main components of which are shown in table 1) obtained by screening a 200-mesh screen with 147.09g of additive (potassium alum) according to the molar ratio of the lithium content in the lithium-rich electrolyte to the aluminum content in the additive being 3:1, then carrying out heat preservation at 200 ℃ for 1.5h for dehydration, and then carrying out reaction at 400 ℃ for 12h to obtain 380.14g of baked product A;
(2) Leaching the roasting product A with 750mL of 0.5mol/L dilute nitric acid at 30 ℃ for 2h to obtain a lithium-containing leaching solution B with the lithium concentration of 8.32g/L and 301.86g of filter residue C, fully washing and drying the filter residue C by 6L of deionized water, and returning the filter residue C to an aluminum electrolysis cell to continue to serve as molten salt electrolyte in the aluminum electrolysis process;
(3) The pH of the lithium-containing leaching solution B is regulated to 11.19 by dropwise adding potassium hydroxide solution to purify and remove impurities, then the solution after the pH regulation is evaporated and concentrated to obtain 280mL of lithium-rich solution with the lithium concentration of 22.29g/L, and then sodium carbonate is added into the lithium-rich solution to obtain 30.62g of lithium phosphate product D.
And (3) analyzing the lithium content in the filter residue C obtained after washing and drying, wherein compared with the lithium content in the initial lithium-rich electrolyte, the lithium content in the filter residue C is reduced to 0.09%, the lithium content in the electrolyte is effectively reduced, and the recovery rate of lithium in the lithium-rich electrolyte is calculated to be up to 95.60%.
Example 5
The embodiment provides a method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte, which specifically adopts sodium alum as an additive and adopts deionized water for leaching, and comprises the following steps:
(1) Uniformly mixing 1000g of aluminum electrolysis lithium-rich electrolyte (the main components of which are shown in table 1) obtained by screening a 200-mesh screen with 473.27g of additive (sodium alum) according to the molar ratio of the lithium content in the lithium-rich electrolyte to the aluminum content in the additive being 3:1, then carrying out heat preservation at 300 ℃ for 1h for dehydration, and then carrying out reaction at 550 ℃ for 8h to obtain 1250.07g of baked product A;
(2) Leaching the roasting product A with 2500mL of deionized water at 60 ℃ for 1h to obtain a lithium-containing leaching solution B with the lithium concentration of 8.28g/L and 1006.20g of filter residue C, fully washing and drying the filter residue C by 15L of deionized water, and returning the filter residue C to an aluminum electrolysis cell to continue to serve as molten salt electrolyte in the aluminum electrolysis process;
(3) And (3) dropwise adding sodium hydroxide solution into the lithium-containing leaching solution B to adjust the pH value to 12.09 so as to purify and remove impurities, evaporating and concentrating the solution after the pH adjustment to obtain 815mL of lithium-rich solution with the lithium concentration of 25.41g/L, and adding sodium carbonate into the lithium-rich solution to obtain 101.66g of lithium phosphate product D.
And (3) analyzing the lithium content in the filter residue C obtained after washing and drying, wherein compared with the lithium content in the initial lithium-rich electrolyte, the lithium content in the filter residue C is reduced to 0.10%, the lithium content in the electrolyte is effectively reduced, and the recovery rate of lithium in the lithium-rich electrolyte is calculated to be up to 95.36%.
Comparative example 1
The comparative example was different from example 1 in that 100g of the aluminum electrolysis lithium-rich electrolyte was directly baked without adding an alunite additive and mixing with the aluminum electrolysis lithium-rich electrolyte, and the baking conditions, leaching conditions, and the like were the same as in example 1. The lithium concentration in the leaching solution of the product obtained in the comparative example is only 0.09g/L, the lithium content in the filter residue C is 2.13%, and the recovery rate of lithium in the lithium-rich electrolyte is only 2.12%.
Comparative example 2
This comparative example differs from example 1 in that 100g of a mixture of a lithium-rich electrolyte and potassium alum was calcined at 340 c for 12 hours, all other conditions being the same as in example 1. The lithium concentration in the leaching solution of the product obtained in the comparative example is only 0.25g/L, the lithium content in the filter residue C is 2.05%, and the recovery rate of lithium in the lithium-rich electrolyte is only 5.76%.
Lithium recovery rates of example 1, example 2, example 3, example 4, example 5 and comparative examples 1 and 2 were plotted, and the results are shown in fig. 2. From the results in the figures, the excellent effects of the method for reducing the lithium content and recovering the lithium in the lithium-rich electrolyte of the aluminum electrolysis provided by the invention depend on the addition of additives and the reasonable control of the roasting temperature, and are the result of the comprehensive effect.
Comparative example 3
This comparative example differs from example 1 in that the 100g mixture of lithium-rich electrolyte and potassium alum was not subjected to the temperature-increasing procedure of low-temperature dehydration and high-temperature calcination, but the mixture was directly heated to 750 ℃ and calcined for 0.5h, with the other conditions being the same as in example 1. The product obtained in this comparative example was analyzed to give a lithium concentration of 4.26g/L in the leachate, a lithium content of 0.04% in the filter residue C, and a lithium recovery rate of only 98.34% in the lithium-rich electrolyte. However, the obtained calcined product was 159.08g only, and the obtained residue C was 85.14g only, which means that a part of fluorine in the lithium-rich electrolyte reacted with water vapor at high temperature to generate hydrogen fluoride and was lost in the off-gas, and could not be recovered as aluminum fluoride for aluminum electrolysis.
Comparative example 4
This comparative example was different from example 2 in that aluminum sulfate (34.43 g, molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive 3:1) was used as an additive, and firing was performed after mixing with the lithium-rich electrolyte, except that the conditions were the same as in example 2. The lithium concentration in the leachate of the product obtained in this comparative example was 3.41g/L, the lithium content in the filter residue C was 0.47%, which was significantly higher than the lithium content in the residue of example 2 by 0.08%, and the recovery rate of lithium in the lithium-rich electrolyte was also only 78.57%.
As can be clearly seen from comparative examples 2 and 4, when alum double salt is used as an additive to mix and bake with a lithium-rich electrolyte, alkali metal or ammonium ions in the alum double salt can act in the same manner as lithium in the lithium-rich electrolyte during baking, which is beneficial to conversion of lithium fluoride, so that the content of lithium in slag can be effectively reduced, and deep extraction of lithium in the lithium-rich electrolyte can be realized.
The technical scheme of the invention is not limited to the specific embodiment, and all technical modifications made according to the technical scheme of the invention fall within the protection scope of the invention.
Claims (10)
1. A method for reducing lithium content and recovering lithium in an aluminum electrolysis lithium-rich electrolyte is characterized in that: the method comprises the following steps:
S1, uniformly mixing an aluminum electrolysis lithium-rich electrolyte with an additive, and roasting to obtain a roasting product A;
s2, leaching the roasting product A to obtain a lithium-containing leaching solution B and filter residues C; the filter residue C is washed and dried and then returned to the aluminum electrolysis cell to be continuously used as electrolyte;
And S3, purifying, removing impurities, evaporating and concentrating the lithium-containing leaching solution B to obtain a lithium-rich solution, and adding a lithium precipitating agent into the lithium-rich solution to obtain a lithium salt product D.
2. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S1, the additive is alum double salt.
3. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 2, wherein the method comprises the steps of: in the step S1, the alum double salt is at least one of potassium alum, ammonium alum and sodium alum.
4. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S1, the molar ratio of the lithium-rich electrolyte to the additive in the aluminum electrolysis is 1-3:1 according to the content of lithium in the lithium-rich electrolyte and the content of aluminum in the additive.
5. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S1, the roasting process is that after heat preservation is carried out for 1-2 hours at 200-300 ℃, the temperature is raised to 400-750 ℃ and the heat preservation is carried out for 0.5-12 hours.
6. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S2, deionized water or acid is adopted in the leaching process; the acid is one of dilute sulfuric acid, dilute hydrochloric acid and dilute nitric acid.
7. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 6, wherein the method comprises the steps of: in the step S2, the concentration of the acid is 0.1-1 mol/L.
8. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S2, the leaching temperature is 20-95 ℃ and the leaching time is 0.5-6 h.
9. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S3, the process of purifying and removing impurities is to add an alkali solution into the lithium-containing leaching solution B, and adjust the pH value of the solution to be more than 10.
10. The method for reducing the lithium content and recovering lithium in an aluminum-electrolyzed lithium-rich electrolyte according to claim 1, wherein the method comprises the steps of: in the step S3, the lithium precipitating agent is at least one of soluble fluoride, carbonate and phosphate.
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