CN116332127A - Method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals - Google Patents
Method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals Download PDFInfo
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- CN116332127A CN116332127A CN202310269322.5A CN202310269322A CN116332127A CN 116332127 A CN116332127 A CN 116332127A CN 202310269322 A CN202310269322 A CN 202310269322A CN 116332127 A CN116332127 A CN 116332127A
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- lithium
- aluminum
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- purity
- silicon
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- 229910052744 lithium Inorganic materials 0.000 title claims abstract description 136
- WHXSMMKQMYFTQS-UHFFFAOYSA-N Lithium Chemical compound [Li] WHXSMMKQMYFTQS-UHFFFAOYSA-N 0.000 title claims abstract description 131
- 239000001257 hydrogen Substances 0.000 title claims abstract description 61
- 229910052739 hydrogen Inorganic materials 0.000 title claims abstract description 61
- 229910052710 silicon Inorganic materials 0.000 title claims abstract description 55
- 239000010703 silicon Substances 0.000 title claims abstract description 55
- 229910052500 inorganic mineral Inorganic materials 0.000 title claims abstract description 35
- 239000011707 mineral Substances 0.000 title claims abstract description 35
- 238000004519 manufacturing process Methods 0.000 title abstract description 37
- 125000004435 hydrogen atom Chemical class [H]* 0.000 title 1
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims abstract description 114
- 239000007788 liquid Substances 0.000 claims abstract description 84
- 229910052782 aluminium Inorganic materials 0.000 claims abstract description 63
- CSDREXVUYHZDNP-UHFFFAOYSA-N alumanylidynesilicon Chemical compound [Al].[Si] CSDREXVUYHZDNP-UHFFFAOYSA-N 0.000 claims abstract description 53
- 229910000676 Si alloy Inorganic materials 0.000 claims abstract description 52
- 230000009467 reduction Effects 0.000 claims abstract description 52
- 239000000292 calcium oxide Substances 0.000 claims abstract description 46
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 claims abstract description 46
- 238000002156 mixing Methods 0.000 claims abstract description 44
- UFHFLCQGNIYNRP-UHFFFAOYSA-N Hydrogen Chemical compound [H][H] UFHFLCQGNIYNRP-UHFFFAOYSA-N 0.000 claims abstract description 42
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 claims abstract description 38
- 239000000463 material Substances 0.000 claims abstract description 37
- 238000000926 separation method Methods 0.000 claims abstract description 36
- PMZURENOXWZQFD-UHFFFAOYSA-L Sodium Sulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=O PMZURENOXWZQFD-UHFFFAOYSA-L 0.000 claims abstract description 27
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims abstract description 27
- 239000000428 dust Substances 0.000 claims abstract description 27
- 229910052938 sodium sulfate Inorganic materials 0.000 claims abstract description 27
- 235000011152 sodium sulphate Nutrition 0.000 claims abstract description 27
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Chemical compound O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 27
- PQVSTLUFSYVLTO-UHFFFAOYSA-N ethyl n-ethoxycarbonylcarbamate Chemical compound CCOC(=O)NC(=O)OCC PQVSTLUFSYVLTO-UHFFFAOYSA-N 0.000 claims abstract description 26
- GLXDVVHUTZTUQK-UHFFFAOYSA-M lithium hydroxide monohydrate Substances [Li+].O.[OH-] GLXDVVHUTZTUQK-UHFFFAOYSA-M 0.000 claims abstract description 26
- 229940040692 lithium hydroxide monohydrate Drugs 0.000 claims abstract description 26
- USOPFYZPGZGBEB-UHFFFAOYSA-N calcium lithium Chemical compound [Li].[Ca] USOPFYZPGZGBEB-UHFFFAOYSA-N 0.000 claims abstract description 23
- 150000001875 compounds Chemical class 0.000 claims abstract description 23
- 239000000706 filtrate Substances 0.000 claims abstract description 23
- 150000002431 hydrogen Chemical class 0.000 claims abstract description 20
- 238000002386 leaching Methods 0.000 claims abstract description 17
- TWNQGVIAIRXVLR-UHFFFAOYSA-N oxo(oxoalumanyloxy)alumane Chemical compound O=[Al]O[Al]=O TWNQGVIAIRXVLR-UHFFFAOYSA-N 0.000 claims abstract description 17
- 239000002351 wastewater Substances 0.000 claims abstract description 17
- 229910001868 water Inorganic materials 0.000 claims abstract description 15
- 238000001354 calcination Methods 0.000 claims abstract description 10
- 229910052751 metal Inorganic materials 0.000 claims abstract description 8
- 239000002184 metal Substances 0.000 claims abstract description 8
- 239000003575 carbonaceous material Substances 0.000 claims abstract description 7
- 238000002425 crystallisation Methods 0.000 claims abstract description 5
- 230000008025 crystallization Effects 0.000 claims abstract description 5
- 238000004821 distillation Methods 0.000 claims abstract description 5
- 238000003825 pressing Methods 0.000 claims abstract description 5
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims description 31
- 229910052802 copper Inorganic materials 0.000 claims description 31
- 239000010949 copper Substances 0.000 claims description 31
- 235000010755 mineral Nutrition 0.000 claims description 31
- 238000000034 method Methods 0.000 claims description 30
- 239000003638 chemical reducing agent Substances 0.000 claims description 20
- 239000000843 powder Substances 0.000 claims description 16
- 230000008569 process Effects 0.000 claims description 10
- 238000006460 hydrolysis reaction Methods 0.000 claims description 8
- 230000007062 hydrolysis Effects 0.000 claims description 7
- 238000005470 impregnation Methods 0.000 claims description 4
- -1 sequentially mixing Substances 0.000 abstract description 7
- 230000008901 benefit Effects 0.000 abstract description 5
- IJGRMHOSHXDMSA-UHFFFAOYSA-N Atomic nitrogen Chemical compound N#N IJGRMHOSHXDMSA-UHFFFAOYSA-N 0.000 description 60
- 238000006722 reduction reaction Methods 0.000 description 50
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 description 49
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 35
- 239000002893 slag Substances 0.000 description 33
- 229910052757 nitrogen Inorganic materials 0.000 description 30
- 238000006243 chemical reaction Methods 0.000 description 27
- 239000002994 raw material Substances 0.000 description 25
- 238000000889 atomisation Methods 0.000 description 24
- PNEYBMLMFCGWSK-UHFFFAOYSA-N aluminium oxide Inorganic materials [O-2].[O-2].[O-2].[Al+3].[Al+3] PNEYBMLMFCGWSK-UHFFFAOYSA-N 0.000 description 21
- 239000002245 particle Substances 0.000 description 21
- 238000001035 drying Methods 0.000 description 20
- 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 19
- FUJCRWPEOMXPAD-UHFFFAOYSA-N lithium oxide Chemical compound [Li+].[Li+].[O-2] FUJCRWPEOMXPAD-UHFFFAOYSA-N 0.000 description 19
- 229910001947 lithium oxide Inorganic materials 0.000 description 19
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N silicon dioxide Inorganic materials O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 18
- 229910052642 spodumene Inorganic materials 0.000 description 18
- 229910052742 iron Inorganic materials 0.000 description 16
- 238000012360 testing method Methods 0.000 description 13
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 description 12
- 239000010439 graphite Substances 0.000 description 12
- 229910002804 graphite Inorganic materials 0.000 description 12
- 229910052629 lepidolite Inorganic materials 0.000 description 12
- 239000000047 product Substances 0.000 description 12
- 239000007787 solid Substances 0.000 description 12
- 239000004927 clay Substances 0.000 description 11
- 238000002360 preparation method Methods 0.000 description 10
- 239000000779 smoke Substances 0.000 description 10
- 239000008188 pellet Substances 0.000 description 9
- 239000000376 reactant Substances 0.000 description 9
- 239000003832 thermite Substances 0.000 description 9
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 8
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 8
- 229910000831 Steel Inorganic materials 0.000 description 8
- 238000001816 cooling Methods 0.000 description 8
- PUZPDOWCWNUUKD-UHFFFAOYSA-M sodium fluoride Chemical compound [F-].[Na+] PUZPDOWCWNUUKD-UHFFFAOYSA-M 0.000 description 8
- 239000010959 steel Substances 0.000 description 8
- 229910002091 carbon monoxide Inorganic materials 0.000 description 7
- 239000000571 coke Substances 0.000 description 7
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N iron oxide Inorganic materials [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 description 7
- 235000012239 silicon dioxide Nutrition 0.000 description 7
- 229910018072 Al 2 O 3 Inorganic materials 0.000 description 6
- 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 6
- 238000005868 electrolysis reaction Methods 0.000 description 6
- 239000003792 electrolyte Substances 0.000 description 6
- 239000010453 quartz Substances 0.000 description 6
- 230000035484 reaction time Effects 0.000 description 6
- 239000011734 sodium Substances 0.000 description 6
- 229910052708 sodium Inorganic materials 0.000 description 6
- 230000009471 action Effects 0.000 description 5
- XFWJKVMFIVXPKK-UHFFFAOYSA-N calcium;oxido(oxo)alumane Chemical compound [Ca+2].[O-][Al]=O.[O-][Al]=O XFWJKVMFIVXPKK-UHFFFAOYSA-N 0.000 description 5
- 239000003795 chemical substances by application Substances 0.000 description 5
- 230000000694 effects Effects 0.000 description 5
- 238000002844 melting Methods 0.000 description 5
- 230000008018 melting Effects 0.000 description 5
- 238000007670 refining Methods 0.000 description 5
- 239000000377 silicon dioxide Substances 0.000 description 5
- 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 4
- KLZUFWVZNOTSEM-UHFFFAOYSA-K Aluminum fluoride Inorganic materials F[Al](F)F KLZUFWVZNOTSEM-UHFFFAOYSA-K 0.000 description 4
- BHPQYMZQTOCNFJ-UHFFFAOYSA-N Calcium cation Chemical compound [Ca+2] BHPQYMZQTOCNFJ-UHFFFAOYSA-N 0.000 description 4
- 238000003723 Smelting Methods 0.000 description 4
- DNEHKUCSURWDGO-UHFFFAOYSA-N aluminum sodium Chemical compound [Na].[Al] DNEHKUCSURWDGO-UHFFFAOYSA-N 0.000 description 4
- OYLGJCQECKOTOL-UHFFFAOYSA-L barium fluoride Chemical compound [F-].[F-].[Ba+2] OYLGJCQECKOTOL-UHFFFAOYSA-L 0.000 description 4
- 229910001632 barium fluoride Inorganic materials 0.000 description 4
- 229910001424 calcium ion Inorganic materials 0.000 description 4
- 239000012295 chemical reaction liquid Substances 0.000 description 4
- 229910001648 diaspore Inorganic materials 0.000 description 4
- 150000004683 dihydrates Chemical class 0.000 description 4
- YDEXUEFDPVHGHE-GGMCWBHBSA-L disodium;(2r)-3-(2-hydroxy-3-methoxyphenyl)-2-[2-methoxy-4-(3-sulfonatopropyl)phenoxy]propane-1-sulfonate Chemical compound [Na+].[Na+].COC1=CC=CC(C[C@H](CS([O-])(=O)=O)OC=2C(=CC(CCCS([O-])(=O)=O)=CC=2)OC)=C1O YDEXUEFDPVHGHE-GGMCWBHBSA-L 0.000 description 4
- 238000001914 filtration Methods 0.000 description 4
- 238000010438 heat treatment Methods 0.000 description 4
- INHCSSUBVCNVSK-UHFFFAOYSA-L lithium sulfate Inorganic materials [Li+].[Li+].[O-]S([O-])(=O)=O INHCSSUBVCNVSK-UHFFFAOYSA-L 0.000 description 4
- 238000007885 magnetic separation Methods 0.000 description 4
- 239000011780 sodium chloride Substances 0.000 description 4
- 239000011775 sodium fluoride Substances 0.000 description 4
- 235000013024 sodium fluoride Nutrition 0.000 description 4
- 239000000243 solution Substances 0.000 description 4
- 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 description 4
- 238000002441 X-ray diffraction Methods 0.000 description 3
- KMWBBMXGHHLDKL-UHFFFAOYSA-N [AlH3].[Si] Chemical compound [AlH3].[Si] KMWBBMXGHHLDKL-UHFFFAOYSA-N 0.000 description 3
- 239000011575 calcium Substances 0.000 description 3
- 238000002474 experimental method Methods 0.000 description 3
- 238000000605 extraction Methods 0.000 description 3
- 238000012921 fluorescence analysis Methods 0.000 description 3
- 239000007789 gas Substances 0.000 description 3
- 238000009533 lab test Methods 0.000 description 3
- 239000012452 mother liquor Substances 0.000 description 3
- 238000005453 pelletization Methods 0.000 description 3
- 230000001376 precipitating effect Effects 0.000 description 3
- 238000011084 recovery Methods 0.000 description 3
- CURLTUGMZLYLDI-UHFFFAOYSA-N Carbon dioxide Chemical compound O=C=O CURLTUGMZLYLDI-UHFFFAOYSA-N 0.000 description 2
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 2
- 229910018068 Li 2 O Inorganic materials 0.000 description 2
- 229910045601 alloy Inorganic materials 0.000 description 2
- 239000000956 alloy Substances 0.000 description 2
- 238000000227 grinding Methods 0.000 description 2
- 239000012535 impurity Substances 0.000 description 2
- XGZVUEUWXADBQD-UHFFFAOYSA-L lithium carbonate Chemical compound [Li+].[Li+].[O-]C([O-])=O XGZVUEUWXADBQD-UHFFFAOYSA-L 0.000 description 2
- 229910052808 lithium carbonate Inorganic materials 0.000 description 2
- KWGKDLIKAYFUFQ-UHFFFAOYSA-M lithium chloride Chemical compound [Li+].[Cl-] KWGKDLIKAYFUFQ-UHFFFAOYSA-M 0.000 description 2
- 229910001760 lithium mineral Inorganic materials 0.000 description 2
- VNWKTOKETHGBQD-UHFFFAOYSA-N methane Chemical compound C VNWKTOKETHGBQD-UHFFFAOYSA-N 0.000 description 2
- 238000011946 reduction process Methods 0.000 description 2
- 229910052814 silicon oxide Inorganic materials 0.000 description 2
- 238000001179 sorption measurement Methods 0.000 description 2
- 241000510672 Cuminum Species 0.000 description 1
- 229910000502 Li-aluminosilicate Inorganic materials 0.000 description 1
- 229910010093 LiAlO Inorganic materials 0.000 description 1
- HBBGRARXTFLTSG-UHFFFAOYSA-N Lithium ion Chemical compound [Li+] HBBGRARXTFLTSG-UHFFFAOYSA-N 0.000 description 1
- 229910021536 Zeolite Inorganic materials 0.000 description 1
- 230000006978 adaptation Effects 0.000 description 1
- 238000003556 assay Methods 0.000 description 1
- 239000012267 brine Substances 0.000 description 1
- 229910002092 carbon dioxide Inorganic materials 0.000 description 1
- 239000001569 carbon dioxide Substances 0.000 description 1
- 125000005587 carbonate group Chemical group 0.000 description 1
- 229910052570 clay Inorganic materials 0.000 description 1
- 239000003245 coal Substances 0.000 description 1
- 239000013078 crystal Substances 0.000 description 1
- 238000011161 development Methods 0.000 description 1
- GUJOJGAPFQRJSV-UHFFFAOYSA-N dialuminum;dioxosilane;oxygen(2-);hydrate Chemical compound O.[O-2].[O-2].[O-2].[Al+3].[Al+3].O=[Si]=O.O=[Si]=O.O=[Si]=O.O=[Si]=O GUJOJGAPFQRJSV-UHFFFAOYSA-N 0.000 description 1
- HNPSIPDUKPIQMN-UHFFFAOYSA-N dioxosilane;oxo(oxoalumanyloxy)alumane Chemical compound O=[Si]=O.O=[Al]O[Al]=O HNPSIPDUKPIQMN-UHFFFAOYSA-N 0.000 description 1
- 238000007599 discharging Methods 0.000 description 1
- 238000005485 electric heating Methods 0.000 description 1
- 239000010433 feldspar Substances 0.000 description 1
- 239000003546 flue gas Substances 0.000 description 1
- 150000002500 ions Chemical class 0.000 description 1
- 238000011068 loading method Methods 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 229910052901 montmorillonite Inorganic materials 0.000 description 1
- 239000003345 natural gas Substances 0.000 description 1
- 239000008239 natural water Substances 0.000 description 1
- 239000007773 negative electrode material Substances 0.000 description 1
- NDLPOXTZKUMGOV-UHFFFAOYSA-N oxo(oxoferriooxy)iron hydrate Chemical compound O.O=[Fe]O[Fe]=O NDLPOXTZKUMGOV-UHFFFAOYSA-N 0.000 description 1
- 230000000737 periodic effect Effects 0.000 description 1
- 238000010587 phase diagram Methods 0.000 description 1
- 229910021420 polycrystalline silicon Inorganic materials 0.000 description 1
- 238000004064 recycling Methods 0.000 description 1
- 239000011435 rock Substances 0.000 description 1
- 239000013535 sea water Substances 0.000 description 1
- 239000002002 slurry Substances 0.000 description 1
- HPALAKNZSZLMCH-UHFFFAOYSA-M sodium;chloride;hydrate Chemical compound O.[Na+].[Cl-] HPALAKNZSZLMCH-UHFFFAOYSA-M 0.000 description 1
- 238000005507 spraying Methods 0.000 description 1
- 239000004575 stone Substances 0.000 description 1
- 230000001502 supplementing effect Effects 0.000 description 1
- 239000008399 tap water Substances 0.000 description 1
- 235000020679 tap water Nutrition 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
- 238000004065 wastewater treatment Methods 0.000 description 1
- 239000010457 zeolite Substances 0.000 description 1
- 229910052644 β-spodumene Inorganic materials 0.000 description 1
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Classifications
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01B—NON-METALLIC ELEMENTS; COMPOUNDS THEREOF; METALLOIDS OR COMPOUNDS THEREOF NOT COVERED BY SUBCLASS C01C
- C01B3/00—Hydrogen; Gaseous mixtures containing hydrogen; Separation of hydrogen from mixtures containing it; Purification of hydrogen
- C01B3/02—Production of hydrogen or of gaseous mixtures containing a substantial proportion of hydrogen
- C01B3/06—Production of hydrogen or of gaseous mixtures containing a substantial proportion of hydrogen by reaction of inorganic compounds containing electro-positively bound hydrogen, e.g. water, acids, bases, ammonia, with inorganic reducing agents
- C01B3/10—Production of hydrogen or of gaseous mixtures containing a substantial proportion of hydrogen by reaction of inorganic compounds containing electro-positively bound hydrogen, e.g. water, acids, bases, ammonia, with inorganic reducing agents by reaction of water vapour with metals
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01B—NON-METALLIC ELEMENTS; COMPOUNDS THEREOF; METALLOIDS OR COMPOUNDS THEREOF NOT COVERED BY SUBCLASS C01C
- C01B33/00—Silicon; Compounds thereof
- C01B33/02—Silicon
- C01B33/037—Purification
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B26/00—Obtaining alkali, alkaline earth metals or magnesium
- C22B26/10—Obtaining alkali metals
- C22B26/12—Obtaining lithium
-
- 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
- C22B5/00—General methods of reducing to metals
- C22B5/02—Dry methods smelting of sulfides or formation of mattes
- C22B5/04—Dry methods smelting of sulfides or formation of mattes by aluminium, other metals or silicon
-
- C—CHEMISTRY; METALLURGY
- C25—ELECTROLYTIC OR ELECTROPHORETIC PROCESSES; APPARATUS THEREFOR
- C25C—PROCESSES FOR THE ELECTROLYTIC PRODUCTION, RECOVERY OR REFINING OF METALS; APPARATUS THEREFOR
- C25C3/00—Electrolytic production, recovery or refining of metals by electrolysis of melts
- C25C3/06—Electrolytic production, recovery or refining of metals by electrolysis of melts of aluminium
- C25C3/24—Refining
-
- 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
- Y02E—REDUCTION OF GREENHOUSE GAS [GHG] EMISSIONS, RELATED TO ENERGY GENERATION, TRANSMISSION OR DISTRIBUTION
- Y02E60/00—Enabling technologies; Technologies with a potential or indirect contribution to GHG emissions mitigation
- Y02E60/30—Hydrogen technology
- Y02E60/36—Hydrogen production from non-carbon containing sources, e.g. by water electrolysis
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- Chemical & Material Sciences (AREA)
- Engineering & Computer Science (AREA)
- Organic Chemistry (AREA)
- Materials Engineering (AREA)
- Chemical Kinetics & Catalysis (AREA)
- Metallurgy (AREA)
- Mechanical Engineering (AREA)
- Manufacturing & Machinery (AREA)
- Inorganic Chemistry (AREA)
- Electrochemistry (AREA)
- Combustion & Propulsion (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- General Health & Medical Sciences (AREA)
- Health & Medical Sciences (AREA)
- Silicon Compounds (AREA)
Abstract
The invention provides a method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals, belonging to the field of resource utilization. Mixing lithium-containing minerals, carbon materials and papermaking wastewater, and then carrying out electrothermal reduction to obtain aluminum-silicon alloy liquid and lithium-rich material dust; carrying out electrochemical separation on the aluminum-silicon alloy liquid to obtain high-purity aluminum and high-purity silicon; atomizing high-purity aluminum to obtain aluminum powder; mixing aluminum powder and water vapor to hydrolyze to prepare hydrogen, and obtaining hydrogen; mixing lithium-rich material dust, sodium sulfate, calcium oxide and water, and performing pressure leaching and solid-liquid separation to obtain filtrate; mixing the filtrate with sodium hydroxide, performing distillation crystallization to obtain lithium hydroxide monohydrate, sequentially mixing, ball pressing and calcining the lithium hydroxide monohydrate, aluminum oxide and calcium oxide to obtain a calcium lithium aluminate compound, and mixing the calcium lithium aluminate compound with high-purity aluminum to perform vacuum aluminothermic reduction to obtain metal lithium. The invention comprehensively utilizes the lithium-containing minerals to obtain the metallic lithium, the high-purity silicon and the hydrogen, and has high economic benefit.
Description
Technical Field
The invention relates to the technical field of resource utilization, in particular to a method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals.
Background
Lithium metal, silicon and hydrogen are the three most important raw materials in the world that are closely related to new energy. Silicon is an important raw material for developing solar silicon cells, hydrogen is a basic stone for hydrogen energy development, and lithium metal is a negative electrode material of solid-state lithium cells. Many methods for producing these materials, such as mainstream processes, use quartz as a raw material to produce elemental silicon by carbothermal reduction; metallic lithium is produced by electrolysis of lithium chloride; there are also many methods of producing hydrogen, such as from coal, natural gas or water electrolysis.
Lithium ores refer to naturally occurring, economically mined lithium resources, and more than 150 lithium minerals and lithium-containing ores have been found in nature. The mineral raw material for preparing lithium is mainly spodumene (Li) 2 O5.8-8.1 wt%), lepidolite (Li-containing) 2 O3.2-6.45 wt%), clay-type lithium ore (containing Li) 2 2.2 to 5.45 weight percent of O). At present, the lithium element is mainly extracted from spodumene ore and lepidolite, and other valuable elements are not utilized, so that resource waste is caused.
Disclosure of Invention
In view of the above, the present invention aims to provide a method for producing hydrogen, metallic lithium and high purity silicon from lithium-containing minerals. The invention comprehensively utilizes the lithium-containing minerals to obtain the metallic lithium, the high-purity silicon and the hydrogen, and compared with the traditional simple lithium extraction, the invention has the advantages of improved economic benefit and broad prospect.
In order to achieve the above object, the present invention provides the following technical solutions:
the invention provides a method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals, which comprises the following steps:
mixing lithium-containing minerals, a reducing agent and papermaking wastewater, and then carrying out electrothermal reduction to obtain aluminum-silicon alloy liquid and lithium-rich material dust;
carrying out electrochemical separation on the aluminum-silicon alloy liquid to obtain high-purity aluminum and high-purity silicon;
atomizing the high-purity aluminum to prepare powder, so as to obtain aluminum powder;
mixing the aluminum powder and water vapor to hydrolyze to prepare hydrogen, so as to obtain hydrogen;
mixing the lithium-rich material dust, sodium sulfate, calcium oxide and water, and then sequentially carrying out pressure leaching and solid-liquid separation to obtain filtrate;
mixing the filtrate with sodium hydroxide, and then carrying out distillation crystallization to obtain lithium hydroxide monohydrate;
mixing, ball pressing and calcining the lithium hydroxide monohydrate, aluminum oxide and calcium oxide in sequence to obtain a calcium lithium aluminate compound;
and mixing the calcium lithium aluminate compound with aluminum powder for vacuum aluminothermic reduction to obtain metal lithium.
Preferably, the mass percentage of the lithium-containing mineral in the material obtained by mixing the lithium-containing mineral, the carbon material and the papermaking wastewater is 65-75%, the mass percentage of the reducing agent is 25-30%, and the mass percentage of the papermaking wastewater is 5-10%.
Preferably, the temperature of the electrothermal reduction is 1900-2200 ℃ and the time is 3-4 h.
Preferably, the electrochemical separation further comprises adding copper, wherein the addition amount of the copper accounts for 30-40% of the total mass of the electrochemical separation system.
Preferably, the electrochemical separation temperature is 700-750 ℃.
Preferably, the temperature of the water vapor is 150-180 ℃.
Preferably, the time for hydrogen production by hydrolysis is 45-60 minutes, and the pressure is 1.2-1.8 MPa.
Preferably, the mass ratio of the sodium sulfate to the lithium-rich material dust is 1-1.5: 2 to 2.5, the mass ratio of the calcium oxide to the lithium-rich material dust is 1 to 1.5:50 to 55.
Preferably, the temperature of the pressure impregnation is 200-250 ℃, the pressure is 3-5 MPa, and the time is 2-5 h.
Preferably, li in the calcium lithium aluminate compound 5 AlO 4 The mass ratio of CaO to aluminum powder is 2.6-3.0: 2.6 to 3.0:0.8 to 1.2.
The invention provides a method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals, which comprises the following steps: mixing lithium-containing minerals, a reducing agent and papermaking wastewater, and then carrying out electrothermal reduction to obtain aluminum-silicon alloy liquid and lithium-rich material dust; carrying out electrochemical separation on the aluminum-silicon alloy liquid to obtain high-purity aluminum and high-purity silicon; atomizing the high-purity aluminum to prepare powder, so as to obtain aluminum powder; mixing the aluminum powder with water vapor to hydrolyze to prepare hydrogen, so as to obtain hydrogen; mixing the lithium-rich material dust, sodium sulfate, calcium oxide and water, and then sequentially carrying out pressure leaching and solid-liquid separation to obtain filtrate; mixing the filtrate with sodium hydroxide, and then carrying out distillation crystallization to obtain lithium hydroxide monohydrate; mixing, ball pressing and calcining the lithium hydroxide monohydrate, aluminum oxide and calcium oxide in sequence to obtain a calcium lithium aluminate compound; and mixing the calcium lithium aluminate compound with aluminum powder for vacuum aluminothermic reduction to obtain metal lithium.
The invention comprehensively utilizes the lithium-containing minerals to obtain metallic lithium, high-purity silicon and hydrogen. Compared with the traditional simple lithium extraction, the method has the advantages of improved economic benefit and wide prospect, and is worthy of popularization. In the invention, the lithium-containing mineral is a lithium aluminosilicate mineral, aluminum-silicon alloy can be obtained through electrothermal reduction with a reducing agent, high-purity silicon and high-purity aluminum are obtained through electrochemical separation of the aluminum-silicon alloy, the high-purity silicon is used as a raw material for producing solar polycrystalline silicon, aluminum powder is obtained after the high-purity aluminum is atomized, and hydrogen is obtained through hydrolysis reaction of the aluminum powder and water vapor. Meanwhile, aluminum powder is used as a reducing agent, aluminum and silicon resources are comprehensively utilized, lithium oxide can enter fine dust particles for recycling, crystal forms are changed, and a foundation is laid for the next step of producing lithium hydroxide monohydrate; leaching the lithium ore with sodium sulfate and calcium oxide, calcining at high temperature, crystallizing to obtain lithium hydroxide monohydrate, and performing thermit reduction on the lithium hydroxide monohydrate, aluminum oxide and calcium oxide to obtain metallic lithium.
Drawings
Fig. 1 is a flow chart of a method for producing hydrogen, metallic lithium and high purity silicon from lithium-containing minerals in an embodiment of the invention.
Detailed Description
The invention provides a method for producing hydrogen, metallic lithium and high-purity silicon from lithium-containing minerals, which comprises the following steps:
mixing lithium-containing minerals, a reducing agent and papermaking wastewater, and then carrying out electrothermal reduction to obtain aluminum-silicon alloy liquid and lithium-rich material dust;
carrying out electrochemical separation on the aluminum-silicon alloy liquid to obtain high-purity aluminum and high-purity silicon;
atomizing the high-purity aluminum to prepare powder, so as to obtain aluminum powder;
mixing the aluminum powder and water vapor to hydrolyze to prepare hydrogen, so as to obtain hydrogen;
mixing the lithium-rich material dust, sodium sulfate, calcium oxide and water, and then sequentially carrying out pressure leaching and solid-liquid separation to obtain filtrate;
mixing the filtrate with sodium hydroxide, and then carrying out distillation crystallization to obtain lithium hydroxide monohydrate;
mixing, ball pressing and calcining the lithium hydroxide monohydrate, aluminum oxide and calcium oxide in sequence to obtain a calcium lithium aluminate compound;
and mixing the calcium lithium aluminate compound with aluminum powder for vacuum aluminothermic reduction to obtain metal lithium.
In the present invention, all materials used are commercial products in the art unless otherwise specified.
According to the invention, lithium-containing minerals, a reducing agent and papermaking wastewater are mixed and subjected to electrothermal reduction, so that aluminum-silicon alloy liquid and lithium-rich material dust are obtained.
In the present invention, the lithium-containing minerals preferably include one or more of diaspore, lepidolite, and clay lithium.
In the present invention, the spodumene ore preferably includes spodumene ore, the mass content of spodumene in the spodumene ore is preferably 95%, and the spodumene ore preferably includes quartz and feldspar.
In the present invention, the ore structure of the diaspore preferably includes a coarse grain structure, a fine grain structure, a texture structure, and a surrounding rock structure.
In the present invention, the assay results of the diaspore are preferably as shown in table 1.
TABLE 1 results of testing main component of Lithosite (mass%)
In the invention, the mass content of lithium oxide in the lepidolite ore is preferably 2.2-6.2%, the mass content of silicon oxide is preferably 50-65%, and the mass content of aluminum oxide is preferably 25-35%.
In the present invention, the mass content of lithium oxide in the clay lithium ore is preferably 2.2 to 6.2%, the mass content of silicon oxide is preferably 50 to 65%, and the mass content of aluminum oxide is preferably 25 to 35%.
In the present invention, the particle diameter of the lithium-containing mineral is preferably 0 to 0.5 mm and the particle diameter of the lithium-containing mineral is not 0.
In the present invention, the reducing agent is preferably a carbon material, and the carbon material preferably includes coke.
In the invention, the mass content of sodium lignin sulfonate in the papermaking wastewater is preferably 90-95%.
In the invention, the mass percentage of the lithium-containing mineral in the material obtained by mixing the lithium-containing mineral, the carbon material and the papermaking wastewater is preferably 65-75%, the mass percentage of the reducing agent is preferably 25-30%, and the mass percentage of the papermaking wastewater is preferably 5-10%.
In the present invention, the lithium-containing minerals and carbon materials are preferably first pulverized and then mixed with the papermaking wastewater.
In the invention, the particle size of the material obtained after the powder preparation is preferably 0-1 mm and the particle size of the material is not 0.
In the invention, the powder preparation preferably further comprises sequentially removing iron, wherein the iron removal is preferably a magnetic separation method, so that the mass content of the ferric oxide is preferably less than 0.5%.
In the present invention, the mixing is preferably further performed by sequentially granulating and drying, and the specific manner of granulating is not particularly limited in the present invention, and may be performed by a manner well known to those skilled in the art.
In the specific embodiment of the invention, the pressure for ball making is preferably 20-30 MPa.
In the present invention, the drying is preferably a drying at a temperature of preferably 150 to 200 ℃ for a time of preferably not more than 1wt% of the moisture content.
In the present invention, the temperature of the electrothermal reduction is preferably 1900 to 2200 ℃, and the time is preferably 3 to 4 hours.
In the present invention, the chemical reaction occurring during the electrothermal reduction is as follows:
Li 2 O·Al 2 O 3 ·4SiO 2 +7C＝2AlSi+Li 2 O+2Si+4CO 2 +3CO。
in the invention, the electrothermal reduction is preferably carried out in a direct-current submerged arc furnace with the electrode diameter phi of 100kVA, the diameter phi of a crucible in the furnace is preferably 150mm, the depth of the furnace is preferably 380mm, the primary voltage of a transformer is preferably 380V, the direct-current voltage of a three-phase bridge rectifier is preferably divided into four stages of 22V, 25V, 28V and 32V, the ratio of current to voltage is preferably 4, the maximum current is preferably 4000A at the voltage of 1000V, a graphite electrode is preferably used in the furnace, and the purity is preferably 99.9 weight percent.
In the invention, in the electrothermal reduction process, the periodic feeding-periodic discharging are preferable, and the interval time is preferably 2-4 hours.
In the invention, the aluminum-silicon alloy liquid preferably comprises the following components in percentage by mass: 35-50% of aluminum, 49-64% of silicon, and not more than 1% of iron and trace other impurity elements.
In the invention, the aluminum-silicon alloy liquid is preferably released into a containing container from an aluminum outlet, and because the aluminum-silicon alloy liquid contains certain nonmetallic impurities, it is necessary to add a refining agent into the containing container to remove slag, and then the slag is removed by filtration, so that the refined aluminum-silicon alloy liquid of pure liquid is obtained.
In the invention, the refining agent preferably comprises the following components in percentage by mass: 30-45% of NaCl and MgCl 2 15-20% and 35-40% of KCl.
In the present invention, the electrothermal reduction preferably also yields slag, preferably as a deoxidizer or dephosphorizer in iron and steel smelting, and flue gas comprising carbon dioxide, carbon monoxide and dust, preferably for the preparation of lithium carbonate.
After the aluminum-silicon alloy liquid is obtained, the aluminum-silicon alloy liquid is subjected to electrochemical separation, so that high-purity aluminum and high-purity silicon are obtained.
In the present invention, the electrochemical separation preferably further comprises adding copper, wherein the addition amount of copper is preferably 30-40% of the total mass of the electrochemical separation system, the purity of copper is preferably 99.5wt%, and the copper is preferably electrolytic copper.
In the invention, the electrolyte for electrochemical separation preferably comprises the following components in percentage by mass: 20% of sodium fluoride, 50% of barium fluoride, 30% of aluminum fluoride and 650 ℃ of melting point.
According to the ternary phase diagram of the silicon-aluminum-copper, the alloy density of the electrochemical separation system is preferably 3.1-3.6 g/cm 3 The density of the electrolyte is 0.5 to 1g/cm higher than that of the electrolyte 3 The melting point of the alloy is 520-580 ℃, copper aluminum silicon liquid formed by mixing the aluminum silicon alloy liquid and electrolytic copper is used as an anode, and is distributed at the bottom, and a high-purity graphite plate is arranged at the bottom and connected with a steel rod; the cathode is aluminum liquid, the aluminum liquid is connected with a high-purity graphite rod, the aluminum-silicon alloy liquid is added periodically, aluminum is continuously precipitated in the cathode region, when the aluminum is continuously reduced, silicon in the anode region is also continuously precipitated, and the high-purity aluminum and the high-purity silicon are taken out periodically; by supplementing raw materials and taking out products, the mass percentage ratio of liquid copper, silicon and aluminum in the anode region is preferably as follows: 30-40% of copper, 10-20% of silicon and 50-60% of aluminum.
In the invention, the precipitated product silicon obtained by electrochemical separation contains aluminum and copper, and is preferably washed by hydrochloric acid to obtain the high-purity silicon, wherein the purity of the high-purity silicon is preferably more than 99.9 weight percent, and the high-purity silicon is used as a raw material for preparing solar cell materials; the high purity aluminum is precipitated in the cathode zone and periodically withdrawn, preferably with a purity of greater than 99.9wt%.
In the present invention, the temperature of the electrochemical separation is preferably 700 to 750 ℃.
In the present invention, the electrochemical separation is preferably performed in an aluminum-silicon separation electrolysis apparatus.
After obtaining high-purity aluminum, the invention atomizes the high-purity aluminum to prepare powder, thereby obtaining aluminum powder.
In the present invention, the high purity aluminum exists in the form of an aluminum liquid.
In the invention, the aluminum liquid is preferably sent into a nitrogen atomization device through a liquid guide groove to be heated, atomized into small liquid drops under the action of venturi effect, and rapidly solidified into aluminum powder under the protection of ambient nitrogen and cooling.
In the present invention, the nitrogen atomizing apparatus preferably includes a nozzle, a high-pressure fan, and a housing.
In the invention, the process of changing the aluminum liquid into aluminum powder is as follows: continuously heating in a nitrogen atomization device, keeping the atomization temperature, atomizing aluminum liquid by nitrogen under the action of the liquid level pressure in the nitrogen atomization device and the venturi effect of nitrogen, spraying an atomization nozzle at the front end of equipment into an atomization chamber, atomizing into small liquid drops, rapidly solidifying the small liquid drops into aluminum powder under the protection and cooling of ambient nitrogen, and sucking the aluminum powder into a trough through a high-pressure fan.
In the invention, the temperature of the atomized powder preparation is preferably 720-750 ℃, the pressure is preferably 2.2-2.5 MPa, the nitrogen flow rate is preferably 300-350 m/s, and the nozzle gap is preferably 0.30-0.45 mm.
After the aluminum powder is obtained, the aluminum powder and the water vapor are mixed for hydrolysis to prepare hydrogen, so that the hydrogen is obtained.
In the present invention, the temperature of the water vapor is preferably 150 to 180 ℃.
In the invention, the hydrolysis hydrogen production preferably comprises adding water, and the mass ratio of the aluminum powder to the water is preferably 45-50: 55 to 50, more preferably 50:50.
in the present invention, the water preferably includes one or more of tap water, dirty water, seawater, alkaline water and brine.
In the present invention, the time for the hydrolysis to produce hydrogen is preferably 45 to 60 minutes, more preferably 40 minutes, the temperature is preferably 150 to 180 ℃, more preferably 160 ℃, and the pressure is preferably 1.2 to 1.8MPa, more preferably 1MPa.
In the invention, the reaction equation for the hydrolysis hydrogen production is as follows:
2Al+3H 2 O→Al 2 O 3 ↓+3H 2 ↑
in the present invention, the hydrogen gas is preferably fed into a hydrogen drying column to be dried, so that hydrogen gas having a purity of 99.9wt% is obtained.
After the hydrolysis hydrogen production is completed, the invention preferably cleans the hydrogen production equipment, pours out the reaction liquid, and alumina is deposited at the bottom of the hydrogen production equipment.
The alumina is preferably dried to obtain the dried alumina, the drying temperature is preferably 140-150 ℃, and the alumina is preferably reduced again in aluminum-silicon alloy reduction equipment and enters the next cycle to form a closed loop.
After the lithium-rich material dust is obtained, the invention mixes the lithium-rich material dust, sodium sulfate, calcium oxide and water and then sequentially carries out pressure leaching and solid-liquid separation to obtain filtrate.
In the invention, the mass ratio of the sodium sulfate to the lithium-rich material dust is preferably 1-1.5: 2 to 2.5, more preferably 1:2, the mass ratio of the calcium oxide to the lithium-rich material dust is preferably 1-1.5: 50 to 55, more preferably 1:50.
in the present invention, the concentration of the pressure impregnation system is preferably 10 to 15wt%.
In the present invention, the particle size of the lithium rich material dust, sodium sulfate and calcium oxide is preferably no greater than 40 microns.
In the present invention, the pressure impregnation is preferably performed at a temperature of 200 to 250 ℃, more preferably 220 to 230 ℃, for a time of 2 to 5 hours, more preferably 3 to 4 hours, and a pressure of 3 to 5MPa.
In the invention, the reaction slag obtained by pressure leaching preferably comprises analcite (sodium aluminum silicate dihydrate) and quartz, the reaction slag is sodium-rich slag, the microstructure of the sodium-rich slag is spherical, a micropore structure is arranged in the interior of the reaction slag, the reaction slag is rough in surface, the reaction slag is zeolite with high added value, the reaction slag can be modified by loading ammonium ions or other ions and is used for agriculture and wastewater treatment, and the application of the sodium-rich slag can not only solve the problem of lithium slag extraction, but also has wide market application prospect.
In the present invention, the filtrate preferably contains sodium sulfate, lithium sulfate and calcium ions.
After the filtrate is obtained, the filtrate is mixed with sodium hydroxide and then distilled and crystallized to obtain the lithium hydroxide monohydrate.
After obtaining the lithium hydroxide monohydrate, the invention sequentially mixes, presses and calcines the lithium hydroxide monohydrate, aluminum oxide and calcium oxide to obtain the calcium lithium aluminate compound.
In the invention, the mass ratio of the lithium hydroxide monohydrate, the aluminum oxide and the calcium oxide is preferably 3.8-4.3: 0.8 to 1.2:2.2 to 2.6.
In the present invention, the temperature of the calcination is preferably 600 to 820 ℃, and the time is preferably 2 to 5 hours, more preferably 3 to 4 hours.
In the present invention, the lithium calcium aluminate compound preferably includes Li 5 AlO 4 And CaO.
After the calcium lithium aluminate compound is obtained, the calcium lithium aluminate compound and aluminum powder are mixed for vacuum aluminothermic reduction to obtain metal lithium.
In the present invention, li in the calcium lithium aluminate compound 5 AlO 4 The mass ratio of CaO to aluminum powder is preferably 2.6-3.0: 2.6 to 3.0:0.8 to 1.2, the excessive aluminum powder can inhibit the progress of the reduction process of the metal lithium, and the reduction rate of the metal lithium can not be improved.
In the present invention, the temperature of the vacuum thermite reduction is preferably 1150-1250 ℃ and the time is preferably 110-130 minutes.
In the present invention, the pressure of the vacuum thermite reduction is preferably 10 to 20Pa.
In the present invention, the vacuum aluminothermic reduction preferably also yields calcium aluminate slag, which preferably includes 12 CaO.7Al 2 O 3 、CaO·2Al 2 O 3 、CaO·Al 2 O 3 、LiAlO 2 CaO and Al, the calcium aluminate slag is preferably used to provide alumina and calcium oxide.
In the present invention, the vacuum thermite reduction is preferably performed in a 120kW continuous lithium reduction test furnace, and the process conditions preferably include: the reaction was carried out in semi-continuous feeds of 30 kg each, operating the lithium test furnace in two to four hours.
In order to further illustrate the present invention, the method of producing hydrogen, metallic lithium and high purity silicon from lithium-containing minerals provided by the present invention is described in detail below with reference to examples, which are not to be construed as limiting the scope of the present invention.
Fig. 1 is a flow chart of a method for producing hydrogen, metallic lithium and high purity silicon from lithium-containing minerals in an embodiment of the invention.
Example 1
Test materials
Spodumene ore powder is obtained from Sichuan Cuminum and has a size ranging from 0 to 0.5 mm. The results of the tests for diaspore are shown in Table 1.
Aluminum silicon alloy production part
Laboratory experiments for producing aluminum-silicon alloys were carried out in a 100kVA direct current submerged arc furnace under the following process conditions: the diameter phi of the electrode is 150mm; the diameter phi of the crucible in the furnace is 300mm; the depth of the furnace is 380mm; the primary voltage of the transformer is 380V, and the direct current voltage of the three-phase bridge rectifier is divided into four stages of 22V, 25V, 28V and 32V; the ratio of current to voltage is 4; maximum current 4000A at a voltage of 1000V. The electrode used in the furnace is a graphite electrode.
The spodumene ore powder has an alumina content of 29.22wt% and a silica content of 62.31wt%. Pure coke (iron content less than 0.5 wt%) is used as reducing agent, and the mass percentage is as follows: 65% of spodumene ore raw material, 25% of pure coke and 10% of papermaking wastewater (containing 90wt% of sodium lignin sulfonate). Grinding spodumene ore raw materials and a reducing agent into powder with the particle size of 0-1 mm, and carrying out iron removal treatment. The iron removal method is a magnetic separation method, so that the iron oxide content is less than 0.5wt%; all materials are then mixed uniformly in a mixer to form the reactants for the next operation. Pelletizing the pretreated reactants: will beThe pretreated reactant is prepared into pellets in a ball press, and the granulating pressure is 20MPa; and (3) heating, drying and dehydrating the prepared pellets at 150 ℃ to ensure that the moisture content is not more than 1 weight percent. The temperature in the direct current arc furnace is adjusted to 1900 ℃, the pellets are added into the direct current arc furnace for reduction reaction for 4 hours, and the aluminum-silicon alloy liquid is obtained, and is fed periodically and discharged periodically (at intervals of 2 hours). Releasing the aluminum-silicon alloy liquid from the aluminum outlet into a container, and adding refining agent (NaCl 40wt% +MgCl) into the container 2 20wt% +40 wt% of KCl) and removing slag by filtration to obtain refined aluminum-silicon alloy liquid of pure liquid.
The slag was analyzed by elemental fluorescence analysis as follows (wt%). Al 18.13,Si 20.50,O 21.88,Ca 4.54,Fe 2.53,Mg 1.03,Cr 0.05,S 0.04,K 0.02,Co 0.02,Zr 0.01,Ni 0.01 and C31.24. The slag is used as deoxidizer or dephosphorizer in iron and steel smelting.
The smoke was analyzed in the experiment as follows. N (N) 2 78 vol.％，O 2 19 vol.％，CO 10vol.％，CO 2 11.5vol.％，SO 2 140 mg/Nm 3 Solid particles 0.6g/Nm 3 ；Li 2 O26 mass% of solid particles Al 2 O 3 10mass% of solid particles, siO 2 62mass% of solid particles. The fine powder collected from the exhaust gas contains up to 26wt% of lithium oxide, and the recovery rate of lithium oxide reaches 94%, and is used as a raw material for extracting metallic lithium.
Preparation of high purity silicon and metallic aluminum by aluminium-silicon separation
The aluminum-silicon alloy liquid from the direct current arc furnace flows into an aluminum-silicon separation electrolysis device after deslagging, and electrolytic copper is added. The purity of electrolytic copper is 99.5wt%, the addition amount is 40% of the total mass of the electrochemical separation system, and the electrolyte adopts 20wt% of sodium fluoride, 50wt% of barium fluoride, 30wt% of aluminum fluoride and the melting point is 650 ℃. The operating temperature was 700 ℃. Taking copper aluminum silicon alloy liquid formed by mixing aluminum silicon alloy liquid and electrolytic copper as an anode, distributing the anode at the bottom, and connecting the anode with a steel rod, wherein the bottom is a high-purity graphite plate; the cathode is aluminum liquid which is connected with the high-purity graphite rod. And (3) adding refined aluminum-silicon alloy liquid at regular intervals, and continuously precipitating aluminum in a cathode region. When aluminum is continuously reduced, silicon in the anode region also starts to be continuously separated out, and high-purity aluminum and high-purity silicon are periodically taken out. By replenishing the raw materials and withdrawing the product, the ratio of liquid copper, silicon and aluminum in the anode region is maintained at: the copper content is 30-40 wt%, the silicon content is 10-20 wt%, and the aluminum content is 50-60 wt%. The precipitated product silicon contains a small amount of aluminum and copper, is washed by hydrochloric acid, has the purity of 99.9 weight percent and is used for preparing the raw materials of the solar cell material; high purity aluminum was precipitated in the cathode zone, periodically removed, and had a purity of 99.9wt%.
Preparation of aluminum powder and production of hydrogen
The aluminum liquid is fed into a nitrogen atomization device through a liquid guide groove to be continuously heated, atomized into small liquid drops under the action of venturi effect, and rapidly solidified into aluminum powder under the protection and cooling of ambient nitrogen, and the nitrogen atomization device comprises a nozzle, a high-pressure fan and a shell. The atomizing nozzle at the front end of the equipment is sprayed into an atomizing chamber to be atomized into small liquid drops, the small liquid drops are rapidly solidified into aluminum powder under the protection and cooling of ambient nitrogen, and the aluminum powder is sucked into a trough through a high-pressure fan. In the nitrogen atomization device, the atomization temperature is 750 ℃, the atomization nitrogen pressure is 2.5MPa, the flow rate of the atomization nitrogen is 350m/s, and the nozzle gap is 0.45mm.
Aluminum powder and water vapor (180 ℃) were mixed at 50:50, the reaction time is 40 minutes, the reaction temperature is 180 ℃, the reaction pressure is 1.6MPa, and the prepared hydrogen enters a hydrogen drying tower for drying to obtain the hydrogen with the purity of 99.9 weight percent. After the hydrogen production is completed, the hydrogen production equipment is cleaned, the reaction liquid is poured out, and alumina is deposited at the bottom of the hydrogen production equipment. Taking out the alumina for drying to obtain dried alumina, wherein the drying temperature is 150 ℃, and then the alumina is reduced again in aluminum-silicon alloy reduction equipment to enter the next cycle to form a closed loop.
Production of lithium hydroxide monohydrate and lithium metal
Because spodumene in the alpha-type spodumene is changed into smoke dust through an electrothermal reduction device, the collected spodumene is converted into beta-type spodumene through X-ray diffraction analysis, the beta-type spodumene contains 26 weight percent of lithium oxide, the beta-type spodumene, sodium sulfate and calcium oxide are mixed and then added with water for slurry mixing, and the pressure leaching reaction is carried out under the following control conditions: the mass ratio of sodium sulfate to beta spodumene is 1:2, the mass ratio of the calcium oxide to the spodumene is 1:50, the reaction temperature is 230 ℃, the reaction time is 3 hours, the concentration is 15wt%, and the leaching rate of lithium can reach 95%. The main components of the combined leaching reaction slag of calcium oxide and sodium sulfate are analcite (sodium aluminum silicate dihydrate) and quartz, and are sodium-rich slag; the pressed and immersed filtrate consists of sodium sulfate, lithium sulfate and a small amount of calcium ions, sodium hydroxide is added into the filtrate, the filtrate is evaporated and frozen, most of sodium sulfate in the solution is separated, and then mother liquor is evaporated and crystallized to prepare a crude lithium hydroxide monohydrate product.
The lithium production equipment is a continuous lithium reduction test furnace with 120kW, and the process conditions are as follows: the reaction was carried out in semi-continuous feeds of 30 kg each, operating the lithium test furnace at a temperature of 1200 ℃ and under vacuum at 10Pa for two hours.
Lithium hydroxide monohydrate, calcium oxide and aluminum oxide are mixed according to the mass ratio of 4:1:2.4 mixing and calcining (at 800 ℃ C. For 3.5 h) to obtain the calcium-lithium aluminate compound (containing Li) 5 AlO 4 With calcium oxide) and using a calcium lithium aluminate compound as a raw material and aluminum powder as a reducing agent to prepare metallic lithium, li 5 AlO 4 : caO: al mass ratio = 2.8:2.8:1, the vacuum aluminothermic reduction temperature is 1200 ℃, the time is 120 minutes, the reduction rate of lithium is 97.76%, and the obtained calcium aluminate slag is reused for preparing aluminum oxide and calcium oxide.
Example 2
The same as in example 1, except that Li was reduced by vacuum thermite 5 AlO 4 : caO: the mass ratio of Al is 3.0:2.7:1, the reduction rate of lithium was 96.5%.
Example 3
The same as in example 1, except that the temperature at the time of vacuum thermite reduction was 1250℃and the time was 120 minutes, the reduction rate of lithium was 98%.
Example 4
Test materials
Lepidolite is obtained from Jiangxi Ganz and has a size ranging from 0 to 0.5 mm. The lepidolite contains 2.2wt% of lithium oxide, 65wt% of silica, 25wt% of alumina and trace amounts of other components.
Aluminum silicon alloy production part
Laboratory experiments for producing aluminum-silicon alloys were carried out in a 100kVA direct current submerged arc furnace under the following process conditions: the diameter phi of the electrode is 150mm; the diameter phi of the crucible in the furnace is 300mm; the depth of the furnace is 380mm; the primary voltage of the transformer is 380V, and the direct current voltage of the three-phase bridge rectifier is divided into four stages of 22V, 25V, 28V and 32V; the ratio of current to voltage is 4; maximum current 4000A at a voltage of 1000V. The electrode used in the furnace is a graphite electrode.
The lepidolite had an alumina content of 25wt% and a silica content of 65wt%. Pure coke (iron content less than 0.5 wt%) is used as reducing agent, and the mass percentage is as follows: 70% of lepidolite ore raw material, 25% of pure coke and 5% of papermaking wastewater (containing 95wt% of sodium lignin sulfonate). Grinding lepidolite raw material and reducing agent into powder with the particle size of 0-1 mm, and carrying out iron removal treatment. The iron removal method is a magnetic separation method, so that the iron oxide content is less than 0.5wt%; all materials are then mixed uniformly in a mixer to form the reactants for the next operation. Pelletizing the pretreated reactants: preparing the pretreated reactant into pellets in a ball press, wherein the granulating pressure is 20MPa; and (3) heating, drying and dehydrating the prepared pellets at 180 ℃ to ensure that the moisture content is not more than 1 weight percent. The temperature in the direct current arc furnace is adjusted to 1900 ℃, the pellets are added into the direct current arc furnace for reduction reaction for 4 hours, and the aluminum-silicon alloy liquid is obtained, and is fed periodically and discharged periodically (at intervals of 2 hours). Releasing the aluminum-silicon alloy liquid from the aluminum outlet into a container, and adding refining agent (NaCl 40wt% +MgCl) into the container 2 20wt% +40 wt% of KCl) and removing slag by filtration to obtain refined aluminum-silicon alloy liquid of pure liquid.
The slag was analyzed by elemental fluorescence analysis as follows (wt%). Al 17.10,Si 21.53,O 20.88,Ca 5.55,Fe 3.53,Mg 1.02,Cr 0.06,S 0.03,K 0.02,Co 0.02,Zr 0.01,Ni 0.01 and C30.24. The slag is used as deoxidizer or dephosphorizer in iron and steel smelting.
The smoke was analyzed in the experiment as follows. N (N) 2 76 vol.％，O 2 19 vol.％，CO 10vol.％，CO 2 13.5vol.％，SO 2 137 mg/Nm 3 0.4g/Nm of solid particles 3 ；Li 2 O27 mass% of solid particles Al 2 O 3 9mass% of solid particles, siO 2 62mass% of solid particles. The fine powder collected from the exhaust gas contains up to 27wt% of lithium oxide, and the recovery rate of lithium oxide reaches 94%, and is used as a raw material for extracting metallic lithium.
Preparation of high purity silicon and metallic aluminum by aluminium-silicon separation
The aluminum-silicon alloy liquid from the direct current arc furnace flows into an aluminum-silicon separation electrolysis device after deslagging, and electrolytic copper is added. The purity of electrolytic copper is 99.5wt%, the addition amount is 30% of the total mass of the electrochemical separation system, and the electrolyte adopts 20wt% of sodium fluoride, 50wt% of barium fluoride, 30wt% of aluminum fluoride and the melting point is 650 ℃. The operating temperature was 750 ℃. Taking copper aluminum silicon alloy liquid formed by mixing aluminum silicon alloy liquid and electrolytic copper as an anode, distributing the anode at the bottom, and connecting the anode with a steel rod, wherein the bottom is a high-purity graphite plate; the cathode is aluminum liquid which is connected with the high-purity graphite rod. And (3) adding refined aluminum-silicon alloy liquid at regular intervals, and continuously precipitating aluminum in a cathode region. When aluminum is continuously reduced, silicon in the anode region also starts to be continuously separated out, and high-purity aluminum and high-purity silicon are periodically taken out. By replenishing the raw materials and withdrawing the product, the ratio of liquid copper, silicon and aluminum in the anode region is maintained at: the copper content was 40wt%, the silicon content was 10wt%, and the aluminum content was 50wt%. The precipitated product silicon contains a small amount of aluminum and copper, is washed by hydrochloric acid, has the purity of 99.9 weight percent and is used for preparing the raw materials of the solar cell material; high purity aluminum was precipitated in the cathode zone, periodically removed, and had a purity of 99.9wt%.
Preparation of aluminum powder and production of hydrogen
The aluminum liquid is fed into a nitrogen atomization device through a liquid guide groove to be continuously heated, atomized into small liquid drops under the action of venturi effect, and rapidly solidified into aluminum powder under the protection and cooling of ambient nitrogen, and the nitrogen atomization device comprises a nozzle, a high-pressure fan and a shell. The atomizing nozzle at the front end of the equipment is sprayed into an atomizing chamber to be atomized into small liquid drops, the small liquid drops are rapidly solidified into aluminum powder under the protection and cooling of ambient nitrogen, and the aluminum powder is sucked into a trough through a high-pressure fan. In the nitrogen atomization device, the atomization temperature is 720 ℃, the atomization nitrogen pressure is 2.2MPa, the flow rate of the atomization nitrogen is 300m/s, and the nozzle gap is 0.3mm.
Aluminum powder and water vapor (150 ℃) were mixed according to a ratio of 48:50, the reaction time is 50 minutes, the reaction temperature is 150 ℃, the reaction pressure is 1.8MPa, and the prepared hydrogen enters a hydrogen drying tower for drying to obtain the hydrogen with the purity of 99.9 weight percent. After the hydrogen production is completed, the hydrogen production equipment is cleaned, the reaction liquid is poured out, and alumina is deposited at the bottom of the hydrogen production equipment. Taking out the alumina for drying to obtain dried alumina, wherein the drying temperature is 145 ℃, and then the alumina is reduced again in aluminum-silicon alloy reduction equipment to enter the next cycle to form a closed loop.
Production of lithium hydroxide monohydrate and lithium metal
Because the lithium mineral in the lepidolite ore is changed into smoke dust through the electric heating reduction equipment, the lepidolite in the smoke dust is defluorinated through X-ray diffraction analysis, the ore phase is transformed, the lepidolite ore contains 26 weight percent of lithium oxide, the smoke dust lithium ore, sodium sulfate and calcium oxide are mixed, then water is added for size mixing, and the pressure leaching reaction is carried out, wherein the control conditions are as follows: the mass ratio of the sodium sulfate to the smoke lithium ore is 1:2, the mass ratio of the calcium oxide to the smoke lithium ore is 1:50, the reaction temperature is 230 ℃, the reaction time is 3 hours, the concentration is 15wt%, and the leaching rate of lithium is 95%. The main components of the combined leaching reaction slag of calcium oxide and sodium sulfate are analcite (sodium aluminum silicate dihydrate) and quartz, and are sodium-rich slag; the pressed and immersed filtrate consists of sodium sulfate, lithium sulfate and a small amount of calcium ions, sodium hydroxide is added into the filtrate, the filtrate is evaporated and frozen, most of sodium sulfate in the solution is separated, and then mother liquor is evaporated and crystallized to prepare a crude lithium hydroxide monohydrate product.
The lithium production equipment is a continuous lithium reduction test furnace with 120kW, and the process conditions are as follows: the reaction was carried out in semi-continuous feeds of 30 kg each, operating the lithium test furnace at a temperature of 1200 ℃ and under vacuum at 10Pa for two hours.
Lithium hydroxide monohydrate, calcium oxide and aluminum oxide are mixed according to the mass ratio of 4.2:1:2.3 mixing and calcining (the temperature is 750 ℃ C., the time is 2 hours),obtain a calcium-lithium aluminate compound (containing Li) 5 AlO 4 With calcium oxide) and using a calcium lithium aluminate compound as a raw material and aluminum powder as a reducing agent to prepare metallic lithium, li 5 AlO 4 : caO: al mass ratio = 2.8:2.8:1, the vacuum aluminothermic reduction temperature is 1200 ℃, the time is 120 minutes, the reduction rate of lithium is 97%, and the obtained calcium aluminate slag is reused for preparing aluminum oxide and calcium oxide.
Example 5
The same as in example 4, except that Li was reduced by vacuum thermite 5 AlO 4 : caO: the mass ratio of Al is 2.9:2.8:1, the reduction rate of lithium was 96%.
Example 6
The same as in example 4, except that the temperature at the time of vacuum thermite reduction was 1150 deg.c, the time was 120 minutes, and the reduction rate of lithium was 96%.
Example 7
Test materials
The clay lithium ore is obtained from Yunnan Yuxi with the size ranging from 0 to 0.5 mm. The Yunnan clay type lithium ore is carbonate clay type lithium ore, wherein lithium carbonate exists between clay type minerals such as montmorillonite mainly in an adsorption mode, and belongs to adsorption lithium. The content of lithium oxide in the clay lithium ore is 3.2wt%, the content of silicon dioxide is 65wt%, the content of aluminum oxide is 30wt%, and other components are trace.
Aluminum silicon alloy production part
Laboratory experiments for producing aluminum-silicon alloys were carried out in a 100kVA direct current submerged arc furnace under the following process conditions: the diameter phi of the electrode is 150mm; the diameter phi of the crucible in the furnace is 300mm; the depth of the furnace is 380mm; the primary voltage of the transformer is 380V, and the direct current voltage of the three-phase bridge rectifier is divided into four stages of 22V, 25V, 28V and 32V; the ratio of current to voltage is 4; maximum current 4000A at a voltage of 1000V. The electrode used in the furnace is a graphite electrode.
Wherein the alumina content is 30wt% and the silica content is 65wt%. Pure coke (iron content less than 0.5 wt%) is used as reducing agent, and the mass percentage is as follows: 65% of clay lithium ore raw material, 30% of pure coke and 5% of papermaking wastewater (containing 95wt% of sodium lignin sulfonate). Clay lithium oreGrinding the raw materials and the reducing agent into powder with the particle size of 0-1 mm, and carrying out iron removal treatment. The iron removal method is a magnetic separation method, so that the iron oxide content is less than 0.5wt%; all materials are then mixed uniformly in a mixer to form the reactants for the next operation. Pelletizing the pretreated reactants: preparing the pretreated reactant into pellets in a ball press, wherein the granulating pressure is 25MPa; and (3) heating, drying and dehydrating the prepared pellets at 200 ℃ to ensure that the moisture content is not more than 1 weight percent. The temperature in the direct current arc furnace is adjusted to 2000 ℃, the pellets are added into the direct current arc furnace for reduction reaction for 4 hours, and the aluminum-silicon alloy liquid is obtained, and is fed regularly and discharged regularly (at intervals of 2 hours). Releasing the aluminum-silicon alloy liquid from the aluminum outlet into a container, and adding refining agent (NaCl 40wt% +MgCl) into the container 2 20wt% +40 wt% of KCl) and removing slag by filtration to obtain refined aluminum-silicon alloy liquid of pure liquid.
The slag was analyzed by elemental fluorescence analysis as follows (wt%). Al 16.10,Si 22.53,O 22.88,Ca 5.54,Fe 3.53,Mg 2.03,Cr 0.03,S 0.06,K 0.03,Co 0.01,Zr 0.01,Ni 0.01 and C29.24. The slag is used as deoxidizer or dephosphorizer in iron and steel smelting.
The smoke was analyzed in the experiment as follows. N (N) 2 76 vol.％，O 2 18 vol.％，CO 12vol.％，CO 2 12.5vol.％，SO 2 130 mg/Nm 3 Solid particles 0.6g/Nm 3 ；Li 2 O24 mass% of solid particles Al 2 O 3 12mass% of solid particles, siO 2 62mass% of solid particles. The fine powder collected from the exhaust gas contains up to 24wt% of lithium oxide, and the recovery rate of lithium oxide reaches 95%, and is used as a raw material for extracting metallic lithium.
Preparation of high purity silicon and metallic aluminum by aluminium-silicon separation
The aluminum-silicon alloy liquid from the direct current arc furnace flows into an aluminum-silicon separation electrolysis device after deslagging, and electrolytic copper is added. The purity of electrolytic copper is 99.5wt%, the addition amount is 30% of the total mass of the electrochemical separation system, and the electrolyte adopts 20wt% of sodium fluoride, 50wt% of barium fluoride, 30wt% of aluminum fluoride and the melting point is 650 ℃. The operating temperature was 750 ℃. Taking copper aluminum silicon alloy liquid formed by mixing aluminum silicon alloy liquid and electrolytic copper as an anode, distributing the anode at the bottom, and connecting the anode with a steel rod, wherein the bottom is a high-purity graphite plate; the cathode is aluminum liquid which is connected with the high-purity graphite rod. And (3) adding refined aluminum-silicon alloy liquid at regular intervals, and continuously precipitating aluminum in a cathode region. When aluminum is continuously reduced, silicon in the anode region also starts to be continuously separated out, and high-purity aluminum and high-purity silicon are periodically taken out. By replenishing the raw materials and withdrawing the product, the ratio of liquid copper, silicon and aluminum in the anode region is maintained at: the copper content was 38wt%, the silicon content was 10wt%, and the aluminum content was 52wt%. The precipitated product silicon contains a small amount of aluminum and copper, is washed by hydrochloric acid, has the purity of 99.9 weight percent and is used for preparing the raw materials of the solar cell material; high purity aluminum was precipitated in the cathode zone, periodically removed, and had a purity of 99.9wt%.
Preparation of aluminum powder and production of hydrogen
The aluminum liquid is fed into a nitrogen atomization device through a liquid guide groove to be continuously heated, atomized into small liquid drops under the action of venturi effect, and rapidly solidified into aluminum powder under the protection and cooling of ambient nitrogen, and the nitrogen atomization device comprises a nozzle, a high-pressure fan and a shell. The atomizing nozzle at the front end of the equipment is sprayed into an atomizing chamber to be atomized into small liquid drops, the small liquid drops are rapidly solidified into aluminum powder under the protection and cooling of ambient nitrogen, and the aluminum powder is sucked into a trough through a high-pressure fan. In the nitrogen atomization device, the atomization temperature is 720 ℃, the atomization nitrogen pressure is 2.2MPa, the flow rate of the atomization nitrogen is 300m/s, and the nozzle gap is 0.3mm.
Aluminum powder and water vapor (140 ℃) were mixed at 48:52, the reaction time is 50 minutes, the reaction temperature is 160 ℃, the reaction pressure is 1.8MPa, and the prepared hydrogen enters a hydrogen drying tower for drying to obtain the hydrogen with the purity of 99.9 weight percent. After the hydrogen production is completed, the hydrogen production equipment is cleaned, the reaction liquid is poured out, and alumina is deposited at the bottom of the hydrogen production equipment. Taking out the alumina for drying to obtain dried alumina, wherein the drying temperature is 150 ℃, and then the alumina is reduced again in aluminum-silicon alloy reduction equipment to enter the next cycle to form a closed loop.
Production of lithium hydroxide monohydrate and lithium metal
Because the lithium ore in the clay type lithium ore is changed into smoke dust through the electrothermal reduction equipment, the clay type lithium ore is decomposed and converted into lithium oxide through X-ray diffraction analysis, the lithium oxide contains 24 weight percent of lithium oxide, the lithium oxide ore, sodium sulfate and calcium oxide are mixed, then water is added for size mixing, and the pressure leaching reaction is carried out, wherein the control conditions are as follows: the mass ratio of the sodium sulfate to the lithium oxide ore is 1:2, the mass ratio of the calcium oxide to the lithium oxide ore is 1:49, the reaction temperature is 220 ℃, the reaction time is 4 hours, the concentration is 18wt%, and the leaching rate of lithium is 96%. The main components of the combined leaching reaction slag of calcium oxide and sodium sulfate are analcite (sodium aluminum silicate dihydrate) and quartz, and are sodium-rich slag; the pressed and immersed filtrate consists of sodium sulfate, lithium sulfate and a small amount of calcium ions, sodium hydroxide is added into the filtrate, the filtrate is evaporated and frozen, most of sodium sulfate in the solution is separated, and then mother liquor is evaporated and crystallized to prepare a crude lithium hydroxide monohydrate product.
The lithium production equipment is a continuous lithium reduction test furnace with 120kW, and the process conditions are as follows: the reaction was carried out in semi-continuous feeds of 30 kg each, operating the lithium test furnace at a temperature of 1200 ℃ and under vacuum at 10Pa for two hours.
Lithium hydroxide monohydrate, calcium oxide and aluminum oxide are mixed according to the mass ratio of 3.9:0.9:2.4 mixing and calcining (at 800 ℃ C. For 3 hours) to obtain a calcium-lithium aluminate compound (containing Li) 5 AlO 4 With calcium oxide) and using a calcium lithium aluminate compound as a raw material and aluminum powder as a reducing agent to prepare metallic lithium, li 5 AlO 4 : caO: al mass ratio = 2.9:2.7:1.1, the vacuum aluminothermic reduction temperature is 1200 ℃, the time is 120 minutes, the reduction rate of lithium is 96.5 percent, and the obtained calcium aluminate slag is reused for preparing aluminum oxide and calcium oxide.
Example 8
The same as in example 7, except that Li was reduced by vacuum thermite 5 AlO 4 : caO: the mass ratio of Al is 2.6:2.6:1, the reduction rate of lithium was 95%.
Example 9
The same as in example 7, except that the temperature at the time of vacuum thermite reduction was 1150 deg.c, the time was 125 minutes, and the reduction rate of lithium was 96.5%.
The foregoing is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any way. It should be noted that modifications and adaptations to the present invention may occur to one skilled in the art without departing from the principles of the present invention and are intended to be comprehended within the scope of the present invention.
Claims (10)
1. A process for producing hydrogen, metallic lithium and high purity silicon from a lithium-containing mineral comprising the steps of:
mixing lithium-containing minerals, a reducing agent and papermaking wastewater, and then carrying out electrothermal reduction to obtain aluminum-silicon alloy liquid and lithium-rich material dust;
carrying out electrochemical separation on the aluminum-silicon alloy liquid to obtain high-purity aluminum and high-purity silicon;
atomizing the high-purity aluminum to prepare powder, so as to obtain aluminum powder;
mixing the aluminum powder and water vapor to hydrolyze to prepare hydrogen, so as to obtain hydrogen;
mixing the lithium-rich material dust, sodium sulfate, calcium oxide and water, and then sequentially carrying out pressure leaching and solid-liquid separation to obtain filtrate;
mixing the filtrate with sodium hydroxide, and then carrying out distillation crystallization to obtain lithium hydroxide monohydrate;
mixing, ball pressing and calcining the lithium hydroxide monohydrate, aluminum oxide and calcium oxide in sequence to obtain a calcium lithium aluminate compound;
and mixing the calcium lithium aluminate compound with aluminum powder for vacuum aluminothermic reduction to obtain metal lithium.
2. The method according to claim 1, wherein the mass percentage of the lithium-containing mineral in the material obtained by mixing the lithium-containing mineral, the carbon material and the papermaking wastewater is 65-75%, the mass percentage of the reducing agent is 25-30%, and the mass percentage of the papermaking wastewater is 5-10%.
3. The method according to claim 1 or 2, wherein the electrothermic reduction is carried out at a temperature of 1900-2200 ℃ for a time of 3-4 hours.
4. The method according to claim 1, wherein the electrochemical separation further comprises adding copper, wherein the addition amount of copper is 30-40% of the total mass of the electrochemical separation system.
5. The method according to claim 1 or 4, wherein the electrochemical separation temperature is 700-750 ℃.
6. The method according to claim 1, wherein the temperature of the water vapor is 150-180 ℃.
7. The method of claim 1 or 6, wherein the hydrolysis is performed for 45 to 60 minutes at a pressure of 1.2 to 1.8MPa.
8. The method according to claim 1, wherein the mass ratio of sodium sulfate to lithium-rich material dust is 1-1.5: 2 to 2.5, the mass ratio of the calcium oxide to the lithium-rich material dust is 1 to 1.5:50 to 55.
9. The method according to claim 1 or 8, wherein the pressure impregnation is carried out at a temperature of 200-250 ℃, a pressure of 3-5 MPa, and a time of 2-5 h.
10. The method of claim 1, wherein Li in the lithium calcium aluminate compound 5 AlO 4 The mass ratio of CaO to aluminum powder is 2.6-3.0: 2.6 to 3.0:0.8 to 1.2.
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