CN117187596A - Comprehensive recovery method for ferrovanadium in vanadium-containing steel slag - Google Patents
Comprehensive recovery method for ferrovanadium in vanadium-containing steel slag Download PDFInfo
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- 229910052720 vanadium Inorganic materials 0.000 title claims abstract description 191
- LEONUFNNVUYDNQ-UHFFFAOYSA-N vanadium atom Chemical compound [V] LEONUFNNVUYDNQ-UHFFFAOYSA-N 0.000 title claims abstract description 181
- 239000002893 slag Substances 0.000 title claims abstract description 112
- 229910000831 Steel Inorganic materials 0.000 title claims abstract description 105
- 239000010959 steel Substances 0.000 title claims abstract description 105
- 238000000034 method Methods 0.000 title claims abstract description 90
- PNXOJQQRXBVKEX-UHFFFAOYSA-N iron vanadium Chemical compound [V].[Fe] PNXOJQQRXBVKEX-UHFFFAOYSA-N 0.000 title claims abstract description 60
- 238000011084 recovery Methods 0.000 title claims abstract description 26
- 229910000628 Ferrovanadium Inorganic materials 0.000 title claims abstract description 21
- 238000002386 leaching Methods 0.000 claims abstract description 142
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 claims abstract description 45
- 229910052921 ammonium sulfate Inorganic materials 0.000 claims abstract description 45
- 235000011130 ammonium sulphate Nutrition 0.000 claims abstract description 45
- UNTBPXHCXVWYOI-UHFFFAOYSA-O azanium;oxido(dioxo)vanadium Chemical compound [NH4+].[O-][V](=O)=O UNTBPXHCXVWYOI-UHFFFAOYSA-O 0.000 claims abstract description 37
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 claims abstract description 35
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims abstract description 25
- 235000011114 ammonium hydroxide Nutrition 0.000 claims abstract description 25
- 239000012452 mother liquor Substances 0.000 claims abstract description 23
- 239000000203 mixture Substances 0.000 claims abstract description 22
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 claims abstract description 21
- 238000001816 cooling Methods 0.000 claims abstract description 21
- 238000001914 filtration Methods 0.000 claims abstract description 21
- CURLTUGMZLYLDI-UHFFFAOYSA-N Carbon dioxide Chemical compound O=C=O CURLTUGMZLYLDI-UHFFFAOYSA-N 0.000 claims abstract description 18
- 239000001099 ammonium carbonate Substances 0.000 claims abstract description 14
- 229910052799 carbon Inorganic materials 0.000 claims abstract description 14
- 238000001556 precipitation Methods 0.000 claims abstract description 14
- ATRRKUHOCOJYRX-UHFFFAOYSA-N Ammonium bicarbonate Chemical compound [NH4+].OC([O-])=O ATRRKUHOCOJYRX-UHFFFAOYSA-N 0.000 claims abstract description 13
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 claims abstract description 13
- 150000003863 ammonium salts Chemical class 0.000 claims abstract description 12
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 12
- 238000000975 co-precipitation Methods 0.000 claims abstract description 11
- 229910021529 ammonia Inorganic materials 0.000 claims abstract description 10
- 229910000013 Ammonium bicarbonate Inorganic materials 0.000 claims abstract description 9
- 235000012538 ammonium bicarbonate Nutrition 0.000 claims abstract description 9
- 229910002092 carbon dioxide Inorganic materials 0.000 claims abstract description 9
- 239000001569 carbon dioxide Substances 0.000 claims abstract description 8
- 238000001704 evaporation Methods 0.000 claims abstract description 8
- 239000000243 solution Substances 0.000 claims description 66
- GNTDGMZSJNCJKK-UHFFFAOYSA-N divanadium pentaoxide Chemical compound O=[V](=O)O[V](=O)=O GNTDGMZSJNCJKK-UHFFFAOYSA-N 0.000 claims description 24
- 239000007787 solid Substances 0.000 claims description 24
- 238000003756 stirring Methods 0.000 claims description 18
- 230000001276 controlling effect Effects 0.000 claims description 10
- 239000007789 gas Substances 0.000 claims description 9
- 239000007788 liquid Substances 0.000 claims description 7
- 239000010413 mother solution Substances 0.000 claims description 7
- MNNHAPBLZZVQHP-UHFFFAOYSA-N diammonium hydrogen phosphate Chemical compound [NH4+].[NH4+].OP([O-])([O-])=O MNNHAPBLZZVQHP-UHFFFAOYSA-N 0.000 claims description 6
- 235000012501 ammonium carbonate Nutrition 0.000 claims description 4
- VBIXEXWLHSRNKB-UHFFFAOYSA-N ammonium oxalate Chemical compound [NH4+].[NH4+].[O-]C(=O)C([O-])=O VBIXEXWLHSRNKB-UHFFFAOYSA-N 0.000 claims description 4
- 239000004254 Ammonium phosphate Substances 0.000 claims description 3
- 229910000148 ammonium phosphate Inorganic materials 0.000 claims description 3
- 235000019289 ammonium phosphates Nutrition 0.000 claims description 3
- CJOBVZJTOIVNNF-UHFFFAOYSA-N cadmium sulfide Chemical compound [Cd]=S CJOBVZJTOIVNNF-UHFFFAOYSA-N 0.000 claims description 3
- 230000001105 regulatory effect Effects 0.000 claims description 3
- 239000011575 calcium Substances 0.000 abstract description 25
- OYPRJOBELJOOCE-UHFFFAOYSA-N Calcium Chemical compound [Ca] OYPRJOBELJOOCE-UHFFFAOYSA-N 0.000 abstract description 20
- 229910052791 calcium Inorganic materials 0.000 abstract description 20
- 238000000605 extraction Methods 0.000 abstract description 20
- 239000002253 acid Substances 0.000 abstract description 10
- 239000002910 solid waste Substances 0.000 abstract description 3
- ICAWQJNNPMTGPS-UHFFFAOYSA-N [Fe].[V].[Ca] Chemical compound [Fe].[V].[Ca] ICAWQJNNPMTGPS-UHFFFAOYSA-N 0.000 abstract 1
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 51
- 238000006243 chemical reaction Methods 0.000 description 34
- 229910052742 iron Inorganic materials 0.000 description 27
- 230000000052 comparative effect Effects 0.000 description 18
- 239000013078 crystal Substances 0.000 description 18
- 230000001089 mineralizing effect Effects 0.000 description 12
- 239000000126 substance Substances 0.000 description 11
- 239000000292 calcium oxide Substances 0.000 description 9
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 description 9
- 229910052751 metal Inorganic materials 0.000 description 9
- 239000002184 metal Substances 0.000 description 9
- 238000002441 X-ray diffraction Methods 0.000 description 8
- 230000033558 biomineral tissue development Effects 0.000 description 8
- 238000010438 heat treatment Methods 0.000 description 8
- 229910004298 SiO 2 Inorganic materials 0.000 description 7
- 239000000654 additive Substances 0.000 description 7
- 230000000996 additive effect Effects 0.000 description 7
- 238000001354 calcination Methods 0.000 description 7
- 239000003795 chemical substances by application Substances 0.000 description 7
- 238000002425 crystallisation Methods 0.000 description 7
- 230000008025 crystallization Effects 0.000 description 7
- 239000002244 precipitate Substances 0.000 description 7
- 238000004064 recycling Methods 0.000 description 7
- 239000002699 waste material Substances 0.000 description 7
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 6
- 229910010413 TiO 2 Inorganic materials 0.000 description 6
- 238000010586 diagram Methods 0.000 description 5
- 239000002994 raw material Substances 0.000 description 5
- -1 sulfur ions Chemical class 0.000 description 5
- 229910018072 Al 2 O 3 Inorganic materials 0.000 description 4
- NLXLAEXVIDQMFP-UHFFFAOYSA-N Ammonia chloride Chemical compound [NH4+].[Cl-] NLXLAEXVIDQMFP-UHFFFAOYSA-N 0.000 description 4
- VTYYLEPIZMXCLO-UHFFFAOYSA-L Calcium carbonate Chemical compound [Ca+2].[O-]C([O-])=O VTYYLEPIZMXCLO-UHFFFAOYSA-L 0.000 description 4
- 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 4
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 4
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical compound [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 description 4
- 238000005265 energy consumption Methods 0.000 description 4
- 238000005516 engineering process Methods 0.000 description 4
- 239000012535 impurity Substances 0.000 description 4
- 239000000047 product Substances 0.000 description 4
- 230000035484 reaction time Effects 0.000 description 4
- 229910052708 sodium Inorganic materials 0.000 description 4
- 239000011734 sodium Substances 0.000 description 4
- 159000000000 sodium salts Chemical class 0.000 description 4
- 238000009628 steelmaking Methods 0.000 description 4
- OAICVXFJPJFONN-UHFFFAOYSA-N Phosphorus Chemical compound [P] OAICVXFJPJFONN-UHFFFAOYSA-N 0.000 description 3
- 238000004458 analytical method Methods 0.000 description 3
- LCPUDZUWZDSKMX-UHFFFAOYSA-K azane;hydrogen sulfate;iron(3+);sulfate;dodecahydrate Chemical compound [NH4+].O.O.O.O.O.O.O.O.O.O.O.O.[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O LCPUDZUWZDSKMX-UHFFFAOYSA-K 0.000 description 3
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 description 3
- 238000004146 energy storage Methods 0.000 description 3
- 238000010304 firing Methods 0.000 description 3
- 238000002156 mixing Methods 0.000 description 3
- 229910052698 phosphorus Inorganic materials 0.000 description 3
- 239000011574 phosphorus Substances 0.000 description 3
- 230000001376 precipitating effect Effects 0.000 description 3
- 238000002360 preparation method Methods 0.000 description 3
- 238000009853 pyrometallurgy Methods 0.000 description 3
- BHPQYMZQTOCNFJ-UHFFFAOYSA-N Calcium cation Chemical compound [Ca+2] BHPQYMZQTOCNFJ-UHFFFAOYSA-N 0.000 description 2
- 238000003723 Smelting Methods 0.000 description 2
- 235000019270 ammonium chloride Nutrition 0.000 description 2
- 230000002308 calcification Effects 0.000 description 2
- 229910000019 calcium carbonate Inorganic materials 0.000 description 2
- 239000000378 calcium silicate Substances 0.000 description 2
- 229910052918 calcium silicate Inorganic materials 0.000 description 2
- OYACROKNLOSFPA-UHFFFAOYSA-N calcium;dioxido(oxo)silane Chemical compound [Ca+2].[O-][Si]([O-])=O OYACROKNLOSFPA-UHFFFAOYSA-N 0.000 description 2
- 239000003245 coal Substances 0.000 description 2
- 238000004090 dissolution Methods 0.000 description 2
- 238000009854 hydrometallurgy Methods 0.000 description 2
- 239000000463 material Substances 0.000 description 2
- 239000011259 mixed solution Substances 0.000 description 2
- 238000012216 screening Methods 0.000 description 2
- 238000000926 separation method Methods 0.000 description 2
- 239000000377 silicon dioxide Substances 0.000 description 2
- 239000004575 stone Substances 0.000 description 2
- BVKZGUZCCUSVTD-UHFFFAOYSA-M Bicarbonate Chemical compound OC([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-M 0.000 description 1
- BVKZGUZCCUSVTD-UHFFFAOYSA-L Carbonate Chemical compound [O-]C([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-L 0.000 description 1
- 229910000640 Fe alloy Inorganic materials 0.000 description 1
- AFCARXCZXQIEQB-UHFFFAOYSA-N N-[3-oxo-3-(2,4,6,7-tetrahydrotriazolo[4,5-c]pyridin-5-yl)propyl]-2-[[3-(trifluoromethoxy)phenyl]methylamino]pyrimidine-5-carboxamide Chemical compound O=C(CCNC(=O)C=1C=NC(=NC=1)NCC1=CC(=CC=C1)OC(F)(F)F)N1CC2=C(CC1)NN=N2 AFCARXCZXQIEQB-UHFFFAOYSA-N 0.000 description 1
- 229910001069 Ti alloy Inorganic materials 0.000 description 1
- 235000011124 aluminium ammonium sulphate Nutrition 0.000 description 1
- LCQXXBOSCBRNNT-UHFFFAOYSA-K ammonium aluminium sulfate Chemical compound [NH4+].[Al+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O LCQXXBOSCBRNNT-UHFFFAOYSA-K 0.000 description 1
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 229910052797 bismuth Inorganic materials 0.000 description 1
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 description 1
- 238000007664 blowing Methods 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 229910001424 calcium ion Inorganic materials 0.000 description 1
- DNWNZRZGKVWORZ-UHFFFAOYSA-N calcium oxido(dioxo)vanadium Chemical compound [Ca+2].[O-][V](=O)=O.[O-][V](=O)=O DNWNZRZGKVWORZ-UHFFFAOYSA-N 0.000 description 1
- WETINTNJFLGREW-UHFFFAOYSA-N calcium;iron;tetrahydrate Chemical compound O.O.O.O.[Ca].[Fe].[Fe] WETINTNJFLGREW-UHFFFAOYSA-N 0.000 description 1
- 238000004364 calculation method Methods 0.000 description 1
- 239000004568 cement Substances 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 150000001875 compounds Chemical class 0.000 description 1
- 238000000354 decomposition reaction Methods 0.000 description 1
- 230000007423 decrease Effects 0.000 description 1
- 230000018109 developmental process Effects 0.000 description 1
- 239000012065 filter cake Substances 0.000 description 1
- 239000000706 filtrate Substances 0.000 description 1
- 238000011010 flushing procedure Methods 0.000 description 1
- 230000004907 flux Effects 0.000 description 1
- 238000000227 grinding Methods 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N iron oxide Inorganic materials [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 description 1
- SZVJSHCCFOBDDC-UHFFFAOYSA-N iron(II,III) oxide Inorganic materials O=[Fe]O[Fe]O[Fe]=O SZVJSHCCFOBDDC-UHFFFAOYSA-N 0.000 description 1
- 230000001788 irregular Effects 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 238000002844 melting Methods 0.000 description 1
- 230000008018 melting Effects 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 230000006911 nucleation Effects 0.000 description 1
- 238000010899 nucleation Methods 0.000 description 1
- 229910052760 oxygen Inorganic materials 0.000 description 1
- 239000001301 oxygen Substances 0.000 description 1
- 229920000642 polymer Polymers 0.000 description 1
- 230000001737 promoting effect Effects 0.000 description 1
- 235000015598 salt intake Nutrition 0.000 description 1
- 238000007086 side reaction Methods 0.000 description 1
- 235000012239 silicon dioxide Nutrition 0.000 description 1
- 238000005245 sintering Methods 0.000 description 1
- 239000002002 slurry Substances 0.000 description 1
- CDBYLPFSWZWCQE-UHFFFAOYSA-L sodium carbonate Substances [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 description 1
- 229910000029 sodium carbonate Inorganic materials 0.000 description 1
- CMZUMMUJMWNLFH-UHFFFAOYSA-N sodium metavanadate Chemical compound [Na+].[O-][V](=O)=O CMZUMMUJMWNLFH-UHFFFAOYSA-N 0.000 description 1
- KKCBUQHMOMHUOY-UHFFFAOYSA-N sodium oxide Chemical compound [O-2].[Na+].[Na+] KKCBUQHMOMHUOY-UHFFFAOYSA-N 0.000 description 1
- 229910001948 sodium oxide Inorganic materials 0.000 description 1
- NTHWMYGWWRZVTN-UHFFFAOYSA-N sodium silicate Chemical compound [Na+].[Na+].[O-][Si]([O-])=O NTHWMYGWWRZVTN-UHFFFAOYSA-N 0.000 description 1
- 239000006104 solid solution Substances 0.000 description 1
- 239000011029 spinel Substances 0.000 description 1
- 229910052596 spinel Inorganic materials 0.000 description 1
- 239000011232 storage material Substances 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- ZXQVPEBHZMCRMC-UHFFFAOYSA-R tetraazanium;iron(2+);hexacyanide Chemical compound [NH4+].[NH4+].[NH4+].[NH4+].[Fe+2].N#[C-].N#[C-].N#[C-].N#[C-].N#[C-].N#[C-] ZXQVPEBHZMCRMC-UHFFFAOYSA-R 0.000 description 1
- LSGOVYNHVSXFFJ-UHFFFAOYSA-N vanadate(3-) Chemical compound [O-][V]([O-])([O-])=O LSGOVYNHVSXFFJ-UHFFFAOYSA-N 0.000 description 1
- 229910000166 zirconium phosphate Inorganic materials 0.000 description 1
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a comprehensive recovery method of vanadium iron from vanadium-containing steel slag, belongs to the technical field of clean vanadium extraction from vanadium-containing secondary resources, and solves the problems that in the prior art, the recovery rate of vanadium extraction from vanadium-containing steel slag is low, calcium cannot be effectively utilized, or the acid consumption is large, and the environment is not friendly. Comprising the following steps: roasting the vanadium-containing steel slag and ammonium sulfate at a low temperature; absorbing the ammonia-containing tail gas by water to obtain an ammonia water solution; cooling the roasted clinker, and obtaining a first leaching solution and a first leaching residue through first leaching and filtering; adjusting the pH value of the first leaching solution by using an ammonia water solution to carry out ferrovanadium coprecipitation; performing second leaching on the mixture of the ammonium salt (or ammonia water) of the ammonium falcate and the ammonium polyvanadate; cooling and crystallizing the second leaching solution to obtain ammonium metavanadate and vanadium precipitation mother liquor; adding ammonia water and carbon dioxide or ammonium bicarbonate into the first leaching residue to mineralize and fix carbon; and (3) evaporating and crystallizing the mineralized mother liquor in the step (7) and the first mother liquor in the step (4) to obtain ammonium sulfate. The method can realize the comprehensive utilization of the vanadium-iron-calcium in the vanadium-containing solid waste.
Description
Technical Field
The invention relates to the technical field of vanadium-containing secondary resource clean vanadium extraction, in particular to a comprehensive recovery method of vanadium iron from vanadium-containing steel slag.
Background
Vanadium is a strategic key metal and is listed in the list of key minerals in developed countries such as europe and america. China is the country with the most abundant vanadium resources in the world, and is also the country with the highest vanadium yield. The vanadium slag is the main raw material for extracting vanadium in China, and accounts for more than 88% of the vanadium productivity in China. Vanadium is mainly applied to the fields of steel, titanium alloy, chemical industry and energy storage, the vanadium in China is applied to more than 91% in the field of steel, and the vanadium in China is applied to less than 5% in the field of energy storage. Along with the increase of vanadium demand in non-steel fields such as vanadium-based energy storage materials, bismuth vanadate chemical materials and the like and the limitation of steel productivity on vanadium slag productivity, the clean and efficient utilization of secondary and low-grade vanadium resources such as vanadium-containing steel slag, stone coal and the like is increasingly emphasized.
The vanadium-containing steel slag is generated in the steelmaking process of the vanadium molten iron produced by the vanadium titano-magnetite, and although the vanadium content (1% -5%) is lower than that of the vanadium slag (10% -25%), the vanadium content is far higher than that of stone coal (0.13% -1.2%), so that the vanadium-containing steel slag is a valuable vanadium-containing secondary resource. The vanadium-containing steel slag is a byproduct of blast furnace-converter smelting by taking vanadium titanomagnetite as a raw material, and the sources of the vanadium-containing steel slag are divided into two types: one is that the residual vanadium in the semisteel enters the slag through converter steelmaking, and the other is that vanadium-containing molten iron is obtained through direct steelmaking without converting vanadium slag. The annual discharge of iron and steel enterprises in China reaches millions of tons, and most vanadium-containing steel slag is not effectively utilized except a small part of the vanadium-containing steel slag which is returned to sintering utilization. Therefore, the extraction of valuable metal vanadium in the vanadium-containing steel slag is realized, and the method has important significance for promoting the development of vanadium industry and the comprehensive utilization of secondary resources in China.
Vanadium extraction from vanadium-containing steel slag at home and abroad is mainly classified into pyrometallurgy and hydrometallurgy. The pyrometallurgy vanadium extraction mainly comprises ore blending smelting of high-grade vanadium slag, and the wet vanadium extraction mainly comprises a sodium roasting method, a direct leaching method and the like. The steel slag returning method is a main method for pyrometallurgy of vanadium-containing steel slag, and the method takes the vanadium-containing steel slag as a flux to enter a blast furnace along with sinter. The vanadium-containing phase is reduced, smelted and converted into vanadium simple substance, and then enters into blast furnace molten iron to produce high-vanadium molten iron containing 3% -10% of vanadium, and then high-grade vanadium slag with vanadium pentoxide grade reaching 46% is obtained by blowing, so that the vanadium pentoxide or vanadium-iron alloy is prepared. The research shows that the TFe of the sinter is reduced by about 0.116 percent, the phosphorus content is increased by about 0.005 percent, and the phosphorus content in the molten iron is increased by about 0.006 percent every 1 percent of the mixture ratio of the sinter steel slag. Therefore, although vanadium can be recovered by the steel slag returning method, impurity phosphorus is easy to be circularly enriched in molten iron, and dephosphorization burden of a steelmaking process is increased. The content of effective active calcium oxide in the vanadium-containing steel slag is low, the grade of the sinter is reduced, the slag quantity is increased, the energy consumption in the iron-making process is increased, and a large amount of the active calcium oxide is not suitable for being added.
Sodium oxide calcification roasting vanadium extraction technology is mature, but various problems exist. However, the content of calcium in the vanadium-containing steel slag is up to 40%, the higher sodium salt addition is required to be maintained in order to improve the vanadium extraction rate, the calcium silicate wrapping phase is decomposed in the roasting process, and a sodium silicate glass phase with a low melting point is easily generated, and the new wrapping phase is melted to prevent the generation of sodium vanadate, so that the conversion rate of vanadium is reduced. In order to eliminate the influence of calcium vanadate insoluble matters in the sodium roasting and leaching process, researchers adopt a sodium roasting-sodium carbonate leaching method to improve the leaching rate of vanadium. However, the technology has the problems of high sodium salt consumption, low vanadium extraction rate and unavailable sodium salt content in tailings. In addition, aiming at the field of hydrometallurgy of vanadium-containing steel slag, researchers also propose technologies such as calcification roasting-ammonium carbonate leaching, calcium reduction roasting-water leaching, and the like, but the technologies take calcium as a harmful resource, and comprehensive extraction of vanadium and calcium is not considered.
CN114150165a proposes a method for enriching vanadium from steel slag containing vanadium and simultaneously preparing nano calcium carbonate, the technique realizes the enrichment of vanadium in steel slag containing vanadium through acid dissolution reaction, and impurities enter acid solution; then removing impurities from the acid solution and performing a calcium precipitation reaction to obtain nano calcium carbonate; and (5) evaporating and concentrating the calcium-precipitated solution to obtain ammonium chloride, and returning the ammonium chloride to the acid dissolution process for recycling. The disadvantage is that the technology requires a large amount of acid to participate in the reaction because of the large content of calcium in the steel slag containing vanadium, and the cost is high.
In summary, no matter the fire method/wet method process, calcium is taken as a harmful element, in order to eliminate the influence of calcium in the raw materials on the extraction of vanadium, sodium salt needs to be introduced from the source or strong acid is used for leaching, so that the tailings contain sodium/sulfur ions, the tailings are difficult to be absorbed and can only be subjected to landfill treatment, the environment is polluted, and the recovery rate of vanadium in the existing process is lower. Therefore, the method has very important significance on how to realize clean extraction of valuable metals from the vanadium-containing steel slag.
Disclosure of Invention
In view of the above, the invention aims to provide a comprehensive recovery method of vanadium iron from vanadium-containing steel slag, which is used for solving the problems that the recovery rate of vanadium extraction from the existing vanadium-containing steel slag is low, calcium cannot be effectively utilized, the acid consumption is large, and the environment is not friendly.
The aim of the invention is mainly realized by the following technical scheme:
the invention provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, which comprises the following steps:
step 1, roasting vanadium-containing steel slag and ammonium sulfate at a low temperature to obtain roasting clinker and ammonia-containing tail gas;
step 2, ammonia-containing tail gas is absorbed by water to obtain ammonia water solution;
step 3, cooling the roasted clinker, and obtaining a first leaching solution and a first leaching residue through first leaching and filtering;
step 4, regulating the pH value of the first leaching solution by using the ammonia water solution obtained in the step 2 to carry out ferrovanadium coprecipitation, filtering to obtain a first mother solution and filter residues, wherein the filter residues are a mixture of ammonium ferroyellow and ammonium polyvanadate;
step 5, performing second leaching and filtering on the mixture of the ammonium polyvanadate and the ammonium ferrosulfate to obtain second leaching liquid and unreacted ammonium ferrosulfate;
step 6, cooling and crystallizing the second leaching solution to obtain ammonium metavanadate and vanadium precipitation mother liquor;
step 7, adding ammonia water and carbon dioxide or ammonium bicarbonate into the first leaching residue to mineralize and fix carbon, and filtering to obtain mineralized mother liquor and mineralized residue;
and 8, evaporating and crystallizing the mineralized mother liquor in the step 7 and the first mother liquor in the step 4 to obtain ammonium sulfate.
Further, in the step 1, the granularity of the vanadium-containing slag is controlled to be 44-150 mu m.
Further, in the step 1, the mass ratio of the steel slag containing vanadium to the ammonium sulfate is controlled to be 1:1-1:10.
Further, in step 3, the first leaching process includes: stirring at 20-90 deg.c for 30-120 min.
Further, in the step 3, the liquid-solid ratio of the first leaching is controlled to be 2:1-10:1, and the unit of the liquid-solid ratio is l/g.
Further, in the step 4, the temperature of the first leaching solution in the coprecipitation of the ferrovanadium is controlled to be 30-90 ℃.
In step 4, the pH is controlled to be 1 to 7.
Further, in step 5, ammonium salt or ammonia water is used for the second leaching, and the ammonium salt is ammonium carbonate, ammonium bicarbonate, ammonium oxalate, ammonium phosphate or ammonium hydrogen phosphate.
Further, in step 5, the process of controlling the second leaching includes: leaching for 10-120 min at 20-90 ℃.
Further, in the step 6, the ammonium metavanadate can be calcined to obtain vanadium pentoxide.
Compared with the prior art, the invention has the following beneficial effects:
a) According to the comprehensive recovery method for vanadium iron in the vanadium-containing steel slag, after roasting under the mild roasting condition, the vanadium iron is co-precipitated, and then vanadium leaching and separation of ammonium ferrovanadium are carried out, so that the high-efficiency extraction of the vanadium iron component in the vanadium-containing steel slag can be realized, and meanwhile, the calcium-containing tailings can be used as a carbon-fixing raw material. The method has the comprehensive utilization of vanadium, iron and calcium in vanadium-containing solid waste of vanadium-containing steel slag, and has multiple benefits of valuable metal recovery, calcium-containing tailing utilization and high purity of vanadium products. The extraction rate of vanadium in the roasting slag is more than 97%, the extraction rate of iron is more than 97%, and the conversion rate of calcium is more than 95%.
b) The comprehensive recycling method of vanadium iron in vanadium-containing steel slag utilizes calcium element in the vanadium-containing steel slag to mineralize and seal carbon dioxide, and each ton of vanadium-containing steel slag can seal more than or equal to 270kg of CO 2 。
c) The method has simple technical flow, and can recover valuable metal vanadium and iron and simultaneously realize the utilization of calcium and the fixation of carbon dioxide. In addition, the technical route of the invention does not need acid, and is environment-friendly; the method achieves the purpose of treating waste by waste while recovering valuable metal ferrovanadium, and has the prospect of large-scale popularization and application.
Additional features and advantages of the invention will be set forth in the description which follows, and in part will be obvious from the description, or may be learned by practice of the invention. The objectives and other advantages of the invention will be realized and attained by the structure particularly pointed out in the written description and claims thereof as well as the appended drawings.
Drawings
The drawings are only for purposes of illustrating particular embodiments and are not to be construed as limiting the invention, like reference numerals being used to refer to like parts throughout the several views.
FIG. 1 is a process flow diagram of the comprehensive recovery method of vanadium iron in vanadium-containing steel slag;
FIG. 2 is an XRD pattern of the first leaching residue of example 1;
FIG. 3 is an XRD pattern of ammonium metavanadate crystals of example 1;
FIG. 4 is an SEM-EDS diagram of ammonium metavanadate crystals of example 1;
FIG. 5 is an XRD pattern for vanadium pentoxide of example 1;
FIG. 6 is an SEM-EDS diagram of vanadium pentoxide of example 1;
FIG. 7 is an XRD pattern for a mixture of ammonium polyvanadate and ammonium polyvanadate as described in example 1;
FIG. 8 is an SEM-EDS of a mixture of ammonium polyvanadate and ammonium polyvanadate of example 1.
Detailed Description
Preferred embodiments of the present invention are described in detail below with reference to the attached drawing figures, which form a part of the present invention and are used in conjunction with embodiments of the present invention to illustrate the principles of the present invention.
The invention provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, which comprises the following steps:
step 1, roasting vanadium-containing steel slag and ammonium sulfate at a low temperature to obtain roasting clinker and ammonia-containing tail gas;
step 2, ammonia-containing tail gas is absorbed by water to obtain ammonia water solution;
step 3, cooling the roasted clinker, and obtaining a first leaching solution and a first leaching residue through first leaching and filtering;
step 4, regulating the pH value of the first leaching solution by using the ammonia water solution obtained in the step 2 to carry out ferrovanadium coprecipitation, filtering to obtain a first mother solution and filter residues, wherein the filter residues are a mixture of ammonium ferroyellow and ammonium polyvanadate;
step 5, performing second leaching and filtering on the mixture of the ammonium salt or ammonia water to obtain a second leaching solution and unreacted ammonium salt;
step 6, cooling and crystallizing the second leaching solution to obtain ammonium metavanadate and vanadium precipitation mother liquor;
step 7, adding ammonia water and carbon dioxide or ammonium bicarbonate into the first leaching residue to indirectly mineralize and fix carbon, and filtering to obtain mineralized mother liquor and mineralized residue rich in ammonium sulfate;
and 8, evaporating and crystallizing the mineralized mother liquor in the step 7 and the first mother liquor in the step 4 to obtain ammonium sulfate.
Specifically, in the step 1, the phases of the vanadium-containing steel slag mainly include calcium silicate, perovskite oxide, oxide phase RO (a solid solution composed of FeO, mnO, mgO), calcium ferrite, free calcium oxide, spinel and the like.
Specifically, in the step 1, the chemical composition (weight percentage) of the steel slag containing vanadium mainly includes: 32 to 41 percent of CaO, 20 to 23 percent of Fe 2 O 3 、3％～15％MgO、11％～15％SiO 2 、2％～6.5％Al 2 O 3 、0.2％～3.5％P 2 O 5 、1％～4％V 2 O 5 、1％～5％TiO 2 0.8% -5% MnO and other oxygen-containing compounds.
Specifically, in the step 1, considering that the roasting reaction is incomplete due to the excessively large granularity of the steel slag containing vanadium, the excessively small granularity increases the ore grinding cost. Therefore, the granularity of the steel slag containing vanadium is controlled to be 44-150 mu m.
Specifically, in the step 1, considering that the excessive mass ratio of the vanadium-containing steel slag to the ammonium sulfate causes insufficient ammonium sulfate addition, and the excessive small mass ratio causes excessive ammonium sulfate addition, so that the loss of the ammonium sulfate is increased. Therefore, the mass ratio of the vanadium-containing steel slag to the ammonium sulfate is controlled to be 1:1-1:10.
Specifically, in the step 1, the occurrence of the ammonium sulfate decomposition side reaction is considered to be caused by the too high roasting temperature, and the incomplete roasting reaction is caused by the too low roasting temperature; too long a roasting time increases roasting energy consumption, and too short a roasting reaction is incomplete. Thus, the firing temperature is controlled to be 250 to 450 ℃, preferably 250 to 350 ℃, and the firing time is controlled to be 30 to 300 minutes, for example, 90 to 300 minutes.
Specifically, in the above step 1, the calcination may be performed in a tube furnace. Considering that too fast a temperature rise speed leads to too short roasting time, the roasting is not thorough, and too small will increase the roasting time. Therefore, the temperature is controlled to rise to the firing temperature at a rate of 5 to 10 ℃/min.
Specifically, in the step 3, the too high temperature of the first leaching is considered to increase the leaching energy consumption, and the too low temperature can reduce the molecular movement rate and the reaction efficiency; too long a leaching time reduces the reaction efficiency, too short results in incomplete leaching. Thus, the process of controlling the first leaching comprises: stirring at 20-90 deg.c for 30-120 min.
Specifically, in the step 3, considering that the liquid-solid ratio of the first leaching is too high, the leaching agent is excessively excessive, so that the leaching agent is wasted; if the dosage of the leaching agent is too low, the leaching is incomplete. Thus, the liquid-to-solid ratio (l/g) of the first leaching is controlled to be 2:1 to 10:1.
Specifically, in the step 3, the first leaching solution is rich in vanadium and iron ions.
Specifically, in the step 3, the first leaching residue is rich in calcium sulfate and silicon dioxide.
Specifically, in the step 3, the calculation formulas of the leaching rates of vanadium and iron are as follows:
wherein: c x 、V x The concentration and the volume of the leaching solution are g/L and L respectively; m is m 0 The unit is g, which is the mass of the steel slag containing vanadium; w (w) x Is the quality of vanadium/iron in the steel slag containing vanadiumThe amount fraction, in%.
Specifically, in the step 4, considering that the temperature of the first leaching solution is too high and the solution is seriously volatilized when the ferrovanadium is co-precipitated; too low a reaction rate is slow. Therefore, the temperature of the first leaching solution in the coprecipitation of the ferrovanadium is controlled to be 30-90 ℃. Preferably, the temperature of the first leaching solution in the coprecipitation of the ferrovanadium is controlled to be 50-80 ℃.
Specifically, in the step 4, considering that the pH value is too low and the vanadium iron precipitation is incomplete when the vanadium iron coprecipitation is performed; excessive levels cause precipitation of other impurity elements, affecting the purity of the ferrovanadium co-precipitate. Therefore, the pH is controlled to 1 to 7. Preferably, the pH is controlled to be 2.5-5.
Specifically, in the step 5, the ammonium salt may be one or more of ammonium carbonate, ammonium bicarbonate, ammonia water, ammonium oxalate, ammonium phosphate or ammonium hydrogen phosphate.
Specifically, in the step 5, considering that the temperature of the second leaching is too high, the ammonium salt is easily decomposed; too low increases the reaction time and decreases the reaction efficiency. Thus, the process of controlling the second leaching comprises: leaching for 10-120 min at 20-90 ℃. Preferably, leaching is carried out at 40-70 ℃.
Specifically, in the step 5, considering that the liquid-solid ratio of the second leaching is too high, the leaching agent is excessively large, so that waste is caused; too low an amount of leaching agent is insufficient, resulting in incomplete reaction. Thus, the liquid-to-solid ratio (l/g) of the second leaching is controlled to be 2:1 to 20:1, preferably, the liquid-to-solid ratio (l/g) of the second leaching is controlled to be 5:1 to 10:1.
Specifically, in the above step 5, NH leached by ammonium salt 4 + The molar ratio to V is 1:1 to 10:1, preferably 2:1 to 5:1.
Specifically, in the step 5, the second leaching solution is rich in ammonium metavanadate.
Specifically, in the step 6, the second leaching solution is cooled and crystallized at 20-50 ℃, the concentration of ammonium metavanadate before crystallization is 10-30 g/L, the stirring speed is 100-500 r/min, the adding amount of the seed crystal is 0.1-3% of the mass of the ammonium metavanadate in the solution, for example, the adding amount of the seed crystal is 0.5-1.0% of the mass of the ammonium metavanadate in the solution. The seed crystal can be one or a combination of more of ammonium metavanadate, ammonium polyvanadate and vanadium pentoxide, and is preferably ammonium metavanadate.
Specifically, in the step 6, considering that the cooling crystallization temperature is too high, the solubility of ammonium metavanadate is too high and the supersaturation degree is too low, so that the crystallization efficiency is low; too low a cooling temperature requires artificial manufacturing of low temperatures and increases energy consumption. Therefore, the temperature of the crystallized vanadium deposit is controlled to be 20-50 ℃.
Specifically, in the above step 6, the stirring speed is selected to be 100 to 500r/min in view of controlling the crystal nucleus growth and the nucleation number during the cooling crystallization.
Specifically, in the above step 6, the formation of crystalline nuclei is affected by the waste of the seed crystals due to the excessive amount of the cooling crystalline seed. Therefore, the adding amount of the seed crystal is selected to be 0.1-3% of the mass of the ammonium metavanadate in the solution.
Specifically, in the step 6, the vanadium precipitation mother liquor can circularly leach the filter residue in the step 4, and the ammonium metavanadate can be calcined at 350-500 ℃ to obtain vanadium pentoxide.
Specifically, in the step 7, the mineralized slag can be used as a cement additive.
Specifically, in the step 7, the temperature of indirectly mineralizing and fixing carbon is too high, and the liquid volatilizes seriously; too low, the reaction rate is too slow, reducing the reaction efficiency. Therefore, the temperature of the indirect mineralization carbon fixation is controlled to be 30-80 ℃ and the time is controlled to be 10-300 min. Preferably, the temperature of the indirectly mineralized carbon fixation is 30-75 ℃.
Specifically, in the step 7, if the liquid-solid ratio of the indirectly mineralized carbon-fixing agent is too high, the carbon-fixing agent is not completely fixed because the excessive solution containing the carbon source; too low, the slurry is too thick to be easily filtered and separated and the leachate is recycled. Therefore, the liquid-solid ratio (l/g) is controlled to be 1:1-20:1; CO 3 2- /HCO 3 - With SO 4 2- If the molar ratio is too high, reagent waste is caused, and if the molar ratio is too low, the carbonate/bicarbonate amount is insufficient, so that the carbon fixing efficiency is affected. Thus, control of CO 3 2- /HCO 3 - With SO 4 2- The molar ratio of (2) is 1:1-5:1. Preferably, the method comprises the steps of,CO 3 2- /HCO 3 - with SO 4 2- The molar ratio of (2) is 1:1-3:1.
Specifically, in the step 7, in the mineralizing step, the conversion rate of Ca is calculated as follows:
wherein: η (eta) Ca Conversion rate of calcium is expressed in units of; c 1 、c 2 The mass concentration of Ca ions in the solution before and after mineralization reaction is expressed in g/l; v (V) 1 、V 2 The volume of the solution before and after the reaction is given in ml.
In the mineralization step, CO 2 The mineralization rate is calculated as follows:
wherein:the mineralization rate is kg/t; m is m 3 、m 4 Respectively representing the mass before mineralizing and solidifying carbon and the mass after mineralizing and solidifying carbon of a certain substance, wherein the unit is g and c 3 、c 4 The mass fraction of C element before mineralizing and fixing carbon and the mass fraction of C element after mineralizing and fixing carbon are respectively represented. M is M CO2 、M C Respectively represent CO 2 Molar mass with C, in g/mol.
Specifically, in the step 8, the obtained ammonium sulfate can be used as a vanadium-containing steel slag roasting additive to be recycled to the step 1, so that the recycling of the ammonium sulfate is realized.
Compared with the prior art, the comprehensive recovery method of vanadium iron in the vanadium-containing steel slag is carried out after roasting under mild roasting conditions, vanadium iron coprecipitation is carried out first, and then leaching of vanadium and separation of ammonium alum are carried out, so that the high-efficiency extraction of vanadium iron components in the vanadium-containing steel slag can be realized, and meanwhile, tailings can be used as a carbon-fixing raw material. The method has the comprehensive utilization of vanadium, iron and calcium in vanadium-containing solid waste of vanadium-containing steel slag, and has multiple benefits of valuable metal recovery, calcium-containing tailing utilization and high purity of vanadium products. The extraction rate of vanadium is more than 97%, the extraction rate of iron is more than 97%, and the conversion rate of calcium is more than 95%.
The comprehensive recycling method of vanadium iron in vanadium-containing steel slag utilizes calcium element in the vanadium-containing steel slag to mineralize and seal carbon dioxide, and each ton of vanadium-containing steel slag can seal more than or equal to 270kg of CO 2 。
The preparation method is simple and low in cost, and the utilization of calcium and the fixation of carbon dioxide are both realized while valuable metal vanadium and iron are recovered. In addition, the preparation method does not need acid, and is environment-friendly; the method achieves the purpose of treating waste by waste while recovering valuable metal ferrovanadium, and has the prospect of large-scale popularization and application.
The following specific examples and comparative examples illustrate the advantages of the method for the integrated recovery of vanadium iron from vanadium-containing steel slag according to the present invention. The embodiment of the invention provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag.
Example 1:
the embodiment provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, and the chemical composition (weight percent) of the vanadium-containing steel slag in the embodiment comprises 32.09 percent of CaO and 20.28 percent of Fe 2 O 3 、14.79％MgO、11.1％SiO 2 、6.09％Al 2 O 3 、0.21％P 2 O 5 、1.05％V 2 O 5 、1.04％TiO 2 0.85% mno and other oxygenates; the process of the comprehensive recovery method of vanadium iron in the vanadium-containing steel slag in the embodiment is shown in fig. 1, and specifically comprises the following steps:
(1) The vanadium-containing steel slag is sieved to-125 mu m and is evenly mixed with ammonium sulfate, and the mass ratio of the vanadium-containing steel slag to the ammonium sulfate is controlled to be 1:10.
(2) And (3) placing the mixture obtained in the step (1) into a tube furnace, programming the temperature to 350 ℃ at a speed of 10 ℃/min, roasting for 120min, cooling the roasted clinker obtained after the reaction to room temperature, and absorbing ammonia generated in the roasting process by water.
(3) Leaching the baked clinker obtained in the step (2) by using water, stirring for 90min at the temperature of 70 ℃, controlling the liquid-solid ratio (l/g) to be 3:1, and filtering and separatingObtaining a first leaching solution rich in vanadium and iron ions and a first leaching solution containing CaSO as a main component 4 And SiO 2 Is a first leaching residue.
(4) Stirring and heating the first leaching solution obtained in the step (3) to 80 ℃, adding ammonia water to a pH value of 2.5, and filtering after the reaction is finished to obtain a first mother solution and filter residues, wherein the filter residues are ammonium ferrosulfate and ammonium polyvanadate precipitate.
(5) Adding the filter residue obtained in the step (4) into ammonium bicarbonate solution, stirring and heating to 70 ℃ for leaching reaction, wherein the leaching time is 30min, the liquid-solid ratio (l/g) is 5:1, and NH is contained in the solution 4 + The molar ratio of the ammonium metavanadate to V is 3:1, and a second leaching solution and unreacted Huang An iron alum are obtained after leaching, wherein the concentration of the ammonium metavanadate in the leaching solution is 25g/L.
(6) And (3) cooling and crystallizing the second leaching solution obtained in the step (5), wherein the cooling temperature is 30 ℃, the stirring speed is 200r/min, and the adding amount of the seed crystal (the seed crystal is ammonium metavanadate) is 1.0% of the mass of the ammonium metavanadate in the solution. Calcining ammonium metavanadate solid obtained by crystallization at 400 ℃ to obtain powdery V 2 O 5 。
(7) Adding ammonia water solution into the first leaching residue obtained in the step (2), and introducing CO 2 Indirectly mineralizing the gas, wherein the mineralizing temperature is 75 ℃, the reaction time is 30min, the liquid-solid ratio is 3:1, and the CO is 3 2- /HCO 3 - With SO 4 2- The molar ratio of (2) is 3:1.
(8) And (3) evaporating and crystallizing the mineralized mother liquor obtained in the step (7) and the first mother liquor to obtain ammonium sulfate, and using the ammonium sulfate as a roasting additive in the step (1) to realize recycling of the ammonium sulfate.
According to analysis, the leaching rate of V in the low-temperature roasting-leaching process of ammonium sulfate under the process condition is 95%, the leaching rate of Fe is 95%, and V 2 O 5 Purity 99.4%, conversion of mineralized step Ca 96%, total CO 2 The mineralization rate is 287.8kg/t of steel slag containing vanadium.
Example 2:
the embodiment provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, and the chemical composition (mass) of the vanadium-containing steel slag in the embodimentPercentage by weight) comprises 40.86% CaO, 22.87% Fe 2 O 3 、3.26％MgO、14.44SiO 2 、3.85Al 2 O 3 、3.25％P 2 O 5 、2.56％V 2 O 5 、2.83％TiO 2 4.54% mno and other oxygenates; the process of the comprehensive recovery method of vanadium iron in the vanadium-containing steel slag in the embodiment is shown in fig. 1, and specifically comprises the following steps:
(1) And (3) screening the vanadium-containing steel slag to be 75 mu m, uniformly mixing the vanadium-containing steel slag with ammonium sulfate, and controlling the mass ratio of the vanadium-containing steel slag to the ammonium sulfate to be 1:6.
(2) And (3) placing the mixture obtained in the step (1) into a tube furnace, programming the temperature to 300 ℃ at a speed of 10 ℃/min, roasting for 90min, cooling the roasted clinker obtained after the reaction to room temperature, and absorbing ammonia generated in the roasting process by water.
(3) Leaching the baked clinker obtained in the step (2) by using water, stirring for 100min at 90 ℃, controlling the liquid-solid ratio (l/g) to be 10:1, and filtering and separating to obtain a first leaching solution rich in vanadium and iron ions and a main component CaSO 4 And SiO 2 Is a first leaching residue.
(4) Stirring and heating the first leaching solution obtained in the step (3) to 90 ℃, adding ammonia water to a pH value of 3.0, and filtering after the reaction is finished to obtain a first mother solution and filter residues, wherein the filter residues are ammonium ferrosulfate and ammonium polyvanadate precipitate.
(5) Adding the filter residue obtained in the step (4) into a mixed solution of ammonium oxalate and ammonia water, stirring and heating to 50 ℃ for leaching reaction, wherein the leaching condition is that the time is 120min, the liquid-solid ratio (l/g) is 10:1, and the NH in the solution is the same as that of the solution 4 + The molar ratio of the ammonium metavanadate to V is 5:1, and a second leaching solution and unreacted Huang An iron alum are obtained after leaching, wherein the concentration of the ammonium metavanadate in the leaching solution is 25g/L.
(6) And (3) cooling and crystallizing the second leaching solution obtained in the step (5), wherein the cooling temperature is 25 ℃, the stirring speed is 400r/min, and the adding amount of the seed crystal (the seed crystal is ammonium metavanadate) is 0.5% of the mass of the ammonium metavanadate in the solution. Calcining ammonium metavanadate solid obtained by crystallization at 380 ℃ to obtain powdery V 2 O 5 。
(7) The first leaching obtained in the step (2)Adding ammonia water solution into the slag, and introducing CO 2 Indirectly mineralizing the gas, wherein the mineralizing temperature is 55 ℃, the reaction time is 250min, the liquid-solid ratio is 10:1, and the CO is 3 2- /HCO 3 - With SO 4 2- The molar ratio of (2) to (1).
(8) And (3) evaporating and crystallizing the mineralized mother liquor obtained in the step (7) and the first mother liquor to obtain ammonium sulfate, and using the ammonium sulfate as a roasting additive in the step (1) to realize recycling of the ammonium sulfate.
According to analysis, the leaching rate of V in the low-temperature roasting-leaching process of ammonium sulfate under the process condition is 98%, the leaching rate of Fe is 99%, and V 2 O 5 Purity 98.9%, conversion of mineralized step Ca 99%, total CO 2 The mineralization rate is 294.9kg/t of steel slag containing vanadium.
Example 3:
the embodiment provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, and the chemical composition (weight percent) of the vanadium-containing steel slag comprises 40.51 percent of CaO and 21.83 percent of Fe 2 O 3 、11.80％MgO、11.23％SiO 2 、2.79％Al 2 O 3 、1.35％P 2 O 5 、3.63％V 2 O 5 、4.82％TiO 2 1.70% mno and other oxygenates; the process of the comprehensive recovery method of vanadium iron in the vanadium-containing steel slag in the embodiment is shown in fig. 1, and specifically comprises the following steps:
(1) And (3) screening the vanadium-containing steel slag to-45 mu m, uniformly mixing the vanadium-containing steel slag with ammonium sulfate, and controlling the mass ratio of the vanadium-containing steel slag to the ammonium sulfate to be 1:3.
(2) And (3) placing the mixture obtained in the step (1) into a tube furnace, programming the temperature to 250 ℃ at the speed of 10 ℃/min, roasting for 300min, cooling the roasted clinker obtained after the reaction to room temperature, and absorbing ammonia generated in the roasting process by water.
(3) Leaching the baked clinker obtained in the step (2) by using water, stirring for 50min at the temperature of 30 ℃, controlling the liquid-solid ratio (l/g) to be 5:1, and filtering and separating to obtain a first leaching solution rich in vanadium and iron ions and a main component of CaSO 4 And SiO 2 Is a first leaching residue.
(4) Stirring and heating the first leaching solution obtained in the step (3) to 60 ℃, adding ammonia water to a pH value of 5.0, and filtering after the reaction is finished to obtain a first mother solution and filter residues, wherein the filter residues are ammonium ferrosulfate and ammonium polyvanadate precipitate.
(5) Adding the filter residue obtained in the step (4) into a mixed solution of ammonium carbonate and ammonium bicarbonate, stirring and heating to 40 ℃ for leaching reaction, wherein the leaching time is 100min, the liquid-solid ratio (l/g) is 10:1, and the NH in the solution is the same as that of the solution 4 + The molar ratio of the ammonium metavanadate to V is 5:1, and a second leaching solution and unreacted Huang An iron alum are obtained after leaching, wherein the concentration of the ammonium metavanadate in the leaching solution is 15g/L.
(6) And (3) cooling and crystallizing the second leaching solution obtained in the step (5), wherein the cooling temperature is 20 ℃, the stirring speed is 400r/min, and the adding amount of the seed crystal (the seed crystal is ammonium metavanadate) is 0.5% of the mass of the ammonium metavanadate in the solution. Calcining ammonium metavanadate solid obtained by crystallization at 500 ℃ to obtain powdery V 2 O 5 。
(7) Adding ammonia water solution into the first leaching residue obtained in the step (2), and introducing CO 2 Indirectly mineralizing the gas, wherein the mineralizing temperature is 30 ℃, the reaction time is 300min, the liquid-solid ratio is 10:1, and the CO is 3 2- /HCO 3 - With SO 4 2- The molar ratio of (2) was 5:1.
(8) And (3) evaporating and crystallizing the mineralized mother liquor obtained in the step (7) and the first mother liquor to obtain ammonium sulfate, and using the ammonium sulfate as a roasting additive in the step (1) to realize recycling of the ammonium sulfate.
According to analysis, the leaching rate of V in the low-temperature roasting-leaching process of ammonium sulfate under the process condition is 98%, the leaching rate of Fe is 98%, and V 2 O 5 Purity 98.0%, conversion of mineralized step Ca 98%, total CO 2 The mineralization rate is 313.6kg/t of steel slag containing vanadium.
FIG. 2 is an XRD pattern of the first leaching residue of example 1; FIG. 3 is an XRD pattern of ammonium metavanadate crystals of example 1; FIG. 4 is an SEM-EDS diagram of ammonium metavanadate crystals of example 1; FIG. 5 is an XRD pattern for vanadium pentoxide of example 1; FIG. 6 is an SEM-EDS diagram of vanadium pentoxide of example 1; FIG. 7 is an XRD pattern for a mixture of ammonium polyvanadate and ammonium polyvanadate as described in example 1; FIG. 8 is an SEM-EDS of a mixture of ammonium polyvanadate and ammonium polyvanadate of example 1. As can be seen from fig. 2, the first leaching residue is mainly composed of calcium sulfate and silica; it can be seen from fig. 3 that the cooled crystalline product is ammonium metavanadate; as can be seen from fig. 4, the ammonium metavanadate crystal product is cylindrical, and the crystal form is complete; it can be seen from fig. 5 that the material after calcination of ammonium metavanadate is vanadium pentoxide; as can be seen from fig. 6, the vanadium pentoxide is in a fine particulate form and is in an aggregated state; from fig. 7 it can be seen that the ferrovanadium co-precipitate is ammonium ferrosulfate and ammonium polyvanadate; from fig. 8 it can be seen that the ferrovanadium co-precipitate is an irregular large particulate polymer.
The inventors have conducted a great deal of research during the course of the study, and some of these methods, which are not effective, are now exemplified as comparative examples as follows:
comparative example 1:
the comparative example provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, and the chemical composition (weight percent) of the vanadium-containing steel slag in the comparative example comprises 32.09 percent of CaO and 20.28 percent of Fe 2 O 3 、14.79％MgO、11.1％SiO 2 、6.09％Al 2 O 3 、0.21％P 2 O 5 、1.05％V 2 O 5 、1.04％TiO 2 0.85% mno and other oxygenates; the method for comprehensively recovering vanadium iron from vanadium-containing steel slag in the comparative example is approximately the same as that in the example 1, except that: comparative example 1 vanadium iron was recovered stepwise by first precipitating iron and then precipitating vanadium; in example 1, vanadium iron coprecipitate is obtained by coprecipitation of vanadium iron, and ammonium salt is used for leaching vanadium and separating ammonium ferrocyanide.
Step 1, first, the first leaching solution is subjected to iron precipitation in comparative example 1.
Heating the first leaching solution, adding ammonia water to adjust the pH value of the filtrate while heating, continuously stirring in the reaction process, and filtering the feed liquid after the reaction is finished to obtain a first mother solution and filter residues, wherein the filter residues are the ammonium ferrovitriol.
And 2, carrying out vanadium precipitation on the first mother liquor obtained in the step 1.
The method is characterized in that the vanadium precipitation process is carried out in a water bath, and the vanadium precipitation additive adopts ammonium sulfate; ammonium in ammonium salt added with vanadium precipitation additive and leachingThe molar ratio of vanadium in the liquid is 1.5-2.5; the pH value of the system is maintained to be 2.5 by adding sulfuric acid solution with the volume fraction of 15 percent in the vanadium precipitation process; precipitating vanadium at 80-95 deg.c for 60min, filtering to separate solid and liquid, flushing the filter cake with dilute sulfuric acid of pH 2.5 to obtain polyvanadate precipitate and the second mother liquid, calcining ammonium polyvanadate at 450-600 deg.c to obtain V 2 O 5 Ammonia gas generated in the calcination process is absorbed by dilute sulfuric acid to prepare ammonium sulfate.
Comparative example 1 preparation V 2 O 5 The purity was only 85.6%, while V was obtained in example 1 2 O 5 The purity was 99.4%.
Comparative example 2:
the comparative example provides a comprehensive recovery method of vanadium iron in vanadium-containing steel slag, and the chemical composition (weight percent) of the vanadium-containing steel slag in the comparative example comprises 40.86 percent of CaO and 22.87 percent of Fe 2 O 3 、3.26％MgO、14.44SiO 2 、3.85Al 2 O 3 、3.25％P 2 O 5 、2.56％V 2 O 5 、2.83％TiO 2 4.54% mno and other oxygenates; the method for comprehensively recovering vanadium iron from vanadium-containing steel slag in the comparative example is approximately the same as that in the example 1, except that: the temperature of the ammonium sulfate roasting process is increased to 500 ℃ and roasting is carried out for 90min.
The leaching rate of the ammonium sulfate roasting-leaching process V of comparative example 2 was 83% and the leaching rate of Fe was 90%, whereas the leaching rate of the ammonium sulfate roasting-leaching process V of example 2 was 98% and the leaching rate of Fe was 99%.
Comparative example 3:
the chemical composition of the vanadium-containing steel slag of the comparative example is the same as that of the example 3, and the description is omitted here; the method for comprehensively recovering vanadium iron from vanadium-containing steel slag in this comparative example is substantially the same as that in example 3, except that: the mass ratio of the vanadium-containing steel slag to the ammonium sulfate is 2:1.
The leaching rate of the ammonium sulfate roasting-leaching process V of this comparative example was 52% and the leaching rate of Fe was 60%, whereas the leaching rate of the ammonium sulfate roasting-leaching process V of example 3 was 98% and the leaching rate of Fe was 98%.
The present invention is not limited to the above-mentioned embodiments, and any changes or substitutions that can be easily understood by those skilled in the art within the technical scope of the present invention are intended to be included in the scope of the present invention.
Claims (10)
1. The comprehensive recovery method of vanadium iron in vanadium-containing steel slag is characterized by comprising the following steps:
step 1, roasting vanadium-containing steel slag and ammonium sulfate at a low temperature to obtain roasting clinker and ammonia-containing tail gas;
step 2, ammonia-containing tail gas is absorbed by water to obtain ammonia water solution;
step 3, cooling the roasted clinker, and obtaining a first leaching solution and a first leaching residue through first leaching and filtering;
step 4, regulating the pH value of the first leaching solution by using the ammonia water solution obtained in the step 2 to carry out ferrovanadium coprecipitation, filtering to obtain a first mother solution and filter residues, wherein the filter residues are a mixture of ammonium ferroyellow and ammonium polyvanadate;
step 5, performing second leaching and filtering on the mixture of the ammonium polyvanadate and the ammonium ferrosulfate to obtain second leaching liquid and unreacted ammonium ferrosulfate;
step 6, cooling and crystallizing the second leaching solution to obtain ammonium metavanadate and vanadium precipitation mother liquor;
step 7, adding ammonia water and carbon dioxide or ammonium bicarbonate into the first leaching residue to mineralize and fix carbon, and filtering to obtain mineralized mother liquor and mineralized residue;
and 8, evaporating and crystallizing the mineralized mother liquor in the step 7 and the first mother liquor in the step 4 to obtain ammonium sulfate.
2. The method for comprehensively recovering ferrovanadium from vanadium-containing steel slag according to claim 1, wherein in the step 1, the granularity of the vanadium-containing steel slag is controlled to be 44-150 μm.
3. The method for comprehensively recovering vanadium iron from vanadium-containing steel slag according to claim 1, wherein in the step 1, the mass ratio of the vanadium-containing steel slag to the ammonium sulfate is controlled to be 1:1-1:10.
4. The method for comprehensively recovering ferrovanadium from vanadium-containing steel slag according to claim 1, wherein in the step 3, the first leaching process comprises: stirring at 20-90 deg.c for 30-120 min.
5. The method for comprehensively recovering vanadium iron from vanadium-containing steel slag according to claim 1, wherein in the step 3, the liquid-solid ratio of the first leaching is controlled to be 2:1-10:1, and the unit of the liquid-solid ratio is l/g.
6. The method for comprehensively recovering vanadium iron from vanadium-containing steel slag according to claim 1, wherein in the step 4, the temperature of the first leaching solution in the coprecipitation of the vanadium iron is controlled to be 30-90 ℃.
7. The method for comprehensively recovering ferrovanadium from vanadium-containing steel slag according to claim 1, wherein in the step 4, the pH value is controlled to be 1-7.
8. The method for comprehensively recovering ferrovanadium from vanadium-containing steel slag according to claim 1, wherein in the step 5, ammonium salt or ammonia water is adopted for the second leaching, and the ammonium salt adopts ammonium carbonate, ammonium bicarbonate, ammonium oxalate, ammonium phosphate or ammonium hydrogen phosphate.
9. The method for comprehensively recovering ferrovanadium from vanadium-containing steel slag according to claim 1, wherein in the step 5, the process for controlling the second leaching comprises: leaching for 10-120 min at 20-90 ℃.
10. The method for comprehensively recovering vanadium iron from vanadium-containing steel slag according to any one of claims 1 to 9, wherein in the step 6, the ammonium metavanadate is calcined to obtain vanadium pentoxide.
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