CN117758072A - Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process - Google Patents
Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process Download PDFInfo
- Publication number
- CN117758072A CN117758072A CN202311637773.6A CN202311637773A CN117758072A CN 117758072 A CN117758072 A CN 117758072A CN 202311637773 A CN202311637773 A CN 202311637773A CN 117758072 A CN117758072 A CN 117758072A
- Authority
- CN
- China
- Prior art keywords
- scandium
- leaching
- solution
- equal
- acid
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Pending
Links
- 229910052706 scandium Inorganic materials 0.000 title claims abstract description 198
- SIXSYDAISGFNSX-UHFFFAOYSA-N scandium atom Chemical compound [Sc] SIXSYDAISGFNSX-UHFFFAOYSA-N 0.000 title claims abstract description 177
- 238000000034 method Methods 0.000 title claims abstract description 86
- 238000005660 chlorination reaction Methods 0.000 title claims abstract description 22
- 239000002910 solid waste Substances 0.000 title claims abstract description 21
- RTAQQCXQSZGOHL-UHFFFAOYSA-N Titanium Chemical compound [Ti] RTAQQCXQSZGOHL-UHFFFAOYSA-N 0.000 title claims abstract description 20
- 239000010936 titanium Substances 0.000 title claims abstract description 20
- 229910052719 titanium Inorganic materials 0.000 title claims abstract description 20
- 238000004064 recycling Methods 0.000 title claims abstract description 15
- 238000002386 leaching Methods 0.000 claims abstract description 120
- 239000000463 material Substances 0.000 claims abstract description 59
- 238000000605 extraction Methods 0.000 claims abstract description 52
- MUBZPKHOEPUJKR-UHFFFAOYSA-N Oxalic acid Chemical compound OC(=O)C(O)=O MUBZPKHOEPUJKR-UHFFFAOYSA-N 0.000 claims abstract description 42
- 239000002253 acid Substances 0.000 claims abstract description 42
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims abstract description 30
- 229910052782 aluminium Inorganic materials 0.000 claims abstract description 25
- 238000001556 precipitation Methods 0.000 claims abstract description 24
- 239000007788 liquid Substances 0.000 claims abstract description 21
- 230000001376 precipitating effect Effects 0.000 claims abstract description 18
- 239000012535 impurity Substances 0.000 claims abstract description 16
- 235000006408 oxalic acid Nutrition 0.000 claims abstract description 14
- LQPWUWOODZHKKW-UHFFFAOYSA-K scandium(3+);trihydroxide Chemical compound [OH-].[OH-].[OH-].[Sc+3] LQPWUWOODZHKKW-UHFFFAOYSA-K 0.000 claims abstract description 13
- 238000011084 recovery Methods 0.000 claims abstract description 12
- 238000004090 dissolution Methods 0.000 claims abstract description 10
- 239000002244 precipitate Substances 0.000 claims abstract description 8
- HYXGAEYDKFCVMU-UHFFFAOYSA-N scandium oxide Chemical compound O=[Sc]O[Sc]=O HYXGAEYDKFCVMU-UHFFFAOYSA-N 0.000 claims abstract description 7
- 239000000243 solution Substances 0.000 claims description 122
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 40
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 36
- 238000005406 washing Methods 0.000 claims description 34
- KWYUFKZDYYNOTN-UHFFFAOYSA-M Potassium hydroxide Chemical compound [OH-].[K+] KWYUFKZDYYNOTN-UHFFFAOYSA-M 0.000 claims description 32
- 239000012074 organic phase Substances 0.000 claims description 25
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 24
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims description 24
- 238000001914 filtration Methods 0.000 claims description 24
- 239000003513 alkali Substances 0.000 claims description 22
- 239000003638 chemical reducing agent Substances 0.000 claims description 20
- 238000010828 elution Methods 0.000 claims description 19
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 18
- 239000007787 solid Substances 0.000 claims description 17
- -1 scandium ions Chemical class 0.000 claims description 16
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical group OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 claims description 12
- 239000011777 magnesium Substances 0.000 claims description 12
- 229910052749 magnesium Inorganic materials 0.000 claims description 12
- 239000000843 powder Substances 0.000 claims description 12
- 239000002131 composite material Substances 0.000 claims description 11
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 claims description 10
- 150000001768 cations Chemical class 0.000 claims description 10
- JEIPFZHSYJVQDO-UHFFFAOYSA-N iron(III) oxide Inorganic materials O=[Fe]O[Fe]=O JEIPFZHSYJVQDO-UHFFFAOYSA-N 0.000 claims description 10
- 229910052751 metal Inorganic materials 0.000 claims description 10
- 239000002184 metal Substances 0.000 claims description 10
- NBIIXXVUZAFLBC-UHFFFAOYSA-N Phosphoric acid Chemical compound OP(O)(O)=O NBIIXXVUZAFLBC-UHFFFAOYSA-N 0.000 claims description 8
- 238000004537 pulping Methods 0.000 claims description 8
- 238000000227 grinding Methods 0.000 claims description 7
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims description 6
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 6
- 239000011259 mixed solution Substances 0.000 claims description 6
- 239000012670 alkaline solution Substances 0.000 claims description 5
- AZDRQVAHHNSJOQ-UHFFFAOYSA-N alumane Chemical class [AlH3] AZDRQVAHHNSJOQ-UHFFFAOYSA-N 0.000 claims description 5
- 239000008346 aqueous phase Substances 0.000 claims description 5
- 238000001035 drying Methods 0.000 claims description 5
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims description 4
- 229910000147 aluminium phosphate Inorganic materials 0.000 claims description 4
- JOPOVCBBYLSVDA-UHFFFAOYSA-N chromium(6+) Chemical compound [Cr+6] JOPOVCBBYLSVDA-UHFFFAOYSA-N 0.000 claims description 4
- GPRLSGONYQIRFK-UHFFFAOYSA-N hydron Chemical compound [H+] GPRLSGONYQIRFK-UHFFFAOYSA-N 0.000 claims description 4
- 150000002500 ions Chemical class 0.000 claims description 4
- 229910052742 iron Inorganic materials 0.000 claims description 4
- 229910017604 nitric acid Inorganic materials 0.000 claims description 4
- 239000007800 oxidant agent Substances 0.000 claims description 4
- 239000012071 phase Substances 0.000 claims description 4
- HRZFUMHJMZEROT-UHFFFAOYSA-L sodium disulfite Chemical compound [Na+].[Na+].[O-]S(=O)S([O-])(=O)=O HRZFUMHJMZEROT-UHFFFAOYSA-L 0.000 claims description 4
- 229940001584 sodium metabisulfite Drugs 0.000 claims description 4
- 235000010262 sodium metabisulphite Nutrition 0.000 claims description 4
- 239000002904 solvent Substances 0.000 claims description 4
- 239000003795 chemical substances by application Substances 0.000 claims description 3
- 239000012716 precipitator Substances 0.000 claims description 3
- BVKZGUZCCUSVTD-UHFFFAOYSA-M Bicarbonate Chemical compound OC([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-M 0.000 claims description 2
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 claims description 2
- BVKZGUZCCUSVTD-UHFFFAOYSA-L Carbonate Chemical compound [O-]C([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-L 0.000 claims description 2
- FYYHWMGAXLPEAU-UHFFFAOYSA-N Magnesium Chemical compound [Mg] FYYHWMGAXLPEAU-UHFFFAOYSA-N 0.000 claims description 2
- COKWAZNXKVPRTK-UHFFFAOYSA-N S(O)(O)(=O)=O.Cl.P(O)(O)(O)=O Chemical compound S(O)(O)(=O)=O.Cl.P(O)(O)(O)=O COKWAZNXKVPRTK-UHFFFAOYSA-N 0.000 claims description 2
- PMZURENOXWZQFD-UHFFFAOYSA-L Sodium Sulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=O PMZURENOXWZQFD-UHFFFAOYSA-L 0.000 claims description 2
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims description 2
- 235000010210 aluminium Nutrition 0.000 claims description 2
- ANBBXQWFNXMHLD-UHFFFAOYSA-N aluminum;sodium;oxygen(2-) Chemical compound [O-2].[O-2].[Na+].[Al+3] ANBBXQWFNXMHLD-UHFFFAOYSA-N 0.000 claims description 2
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 claims description 2
- 239000000920 calcium hydroxide Substances 0.000 claims description 2
- 229910001861 calcium hydroxide Inorganic materials 0.000 claims description 2
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 claims description 2
- 239000000292 calcium oxide Substances 0.000 claims description 2
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 claims description 2
- 229910052799 carbon Inorganic materials 0.000 claims description 2
- 229910052739 hydrogen Inorganic materials 0.000 claims description 2
- 239000001257 hydrogen Substances 0.000 claims description 2
- 239000003350 kerosene Substances 0.000 claims description 2
- 235000001055 magnesium Nutrition 0.000 claims description 2
- VTHJTEIRLNZDEV-UHFFFAOYSA-L magnesium dihydroxide Chemical compound [OH-].[OH-].[Mg+2] VTHJTEIRLNZDEV-UHFFFAOYSA-L 0.000 claims description 2
- 239000000347 magnesium hydroxide Substances 0.000 claims description 2
- 229910001862 magnesium hydroxide Inorganic materials 0.000 claims description 2
- 239000000395 magnesium oxide Substances 0.000 claims description 2
- CPLXHLVBOLITMK-UHFFFAOYSA-N magnesium oxide Inorganic materials [Mg]=O CPLXHLVBOLITMK-UHFFFAOYSA-N 0.000 claims description 2
- AXZKOIWUVFPNLO-UHFFFAOYSA-N magnesium;oxygen(2-) Chemical compound [O-2].[Mg+2] AXZKOIWUVFPNLO-UHFFFAOYSA-N 0.000 claims description 2
- 239000002245 particle Substances 0.000 claims description 2
- YXJYBPXSEKMEEJ-UHFFFAOYSA-N phosphoric acid;sulfuric acid Chemical compound OP(O)(O)=O.OS(O)(=O)=O YXJYBPXSEKMEEJ-UHFFFAOYSA-N 0.000 claims description 2
- 229910001388 sodium aluminate Inorganic materials 0.000 claims description 2
- 229910052938 sodium sulfate Inorganic materials 0.000 claims description 2
- 235000011152 sodium sulphate Nutrition 0.000 claims description 2
- 229910052725 zinc Inorganic materials 0.000 claims description 2
- 239000011701 zinc Substances 0.000 claims description 2
- 235000016804 zinc Nutrition 0.000 claims description 2
- 230000001590 oxidative effect Effects 0.000 claims 1
- 230000008901 benefit Effects 0.000 abstract description 4
- 239000000047 product Substances 0.000 abstract description 4
- 238000004519 manufacturing process Methods 0.000 abstract description 3
- 230000007613 environmental effect Effects 0.000 abstract description 2
- 229910052720 vanadium Inorganic materials 0.000 description 11
- 238000005191 phase separation Methods 0.000 description 7
- 239000002699 waste material Substances 0.000 description 7
- 238000009835 boiling Methods 0.000 description 6
- 239000000428 dust Substances 0.000 description 5
- 239000002994 raw material Substances 0.000 description 5
- 150000003839 salts Chemical class 0.000 description 5
- 238000005516 engineering process Methods 0.000 description 4
- 238000005070 sampling Methods 0.000 description 4
- 239000002893 slag Substances 0.000 description 4
- 229910052500 inorganic mineral Inorganic materials 0.000 description 3
- 239000011707 mineral Substances 0.000 description 3
- 230000004048 modification Effects 0.000 description 3
- 238000012986 modification Methods 0.000 description 3
- 229910052761 rare earth metal Inorganic materials 0.000 description 3
- 239000000126 substance Substances 0.000 description 3
- 125000002091 cationic group Chemical group 0.000 description 2
- 238000006243 chemical reaction Methods 0.000 description 2
- 239000000706 filtrate Substances 0.000 description 2
- HKIFKSTZUWVYNV-UHFFFAOYSA-N hydrogen peroxide phosphoric acid sulfuric acid Chemical compound P(O)(O)(O)=O.S(O)(O)(=O)=O.OO HKIFKSTZUWVYNV-UHFFFAOYSA-N 0.000 description 2
- 239000002367 phosphate rock Substances 0.000 description 2
- OJMIONKXNSYLSR-UHFFFAOYSA-N phosphorous acid Chemical compound OP(O)O OJMIONKXNSYLSR-UHFFFAOYSA-N 0.000 description 2
- 150000002910 rare earth metals Chemical class 0.000 description 2
- 238000011946 reduction process Methods 0.000 description 2
- WFKWXMTUELFFGS-UHFFFAOYSA-N tungsten Chemical compound [W] WFKWXMTUELFFGS-UHFFFAOYSA-N 0.000 description 2
- 229910052721 tungsten Inorganic materials 0.000 description 2
- 239000010937 tungsten Substances 0.000 description 2
- 238000003723 Smelting Methods 0.000 description 1
- GWEVSGVZZGPLCZ-UHFFFAOYSA-N Titan oxide Chemical compound O=[Ti]=O GWEVSGVZZGPLCZ-UHFFFAOYSA-N 0.000 description 1
- 229910052770 Uranium Inorganic materials 0.000 description 1
- 229910001570 bauxite Inorganic materials 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 238000007796 conventional method Methods 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 229910000514 dolomite Inorganic materials 0.000 description 1
- 239000010459 dolomite Substances 0.000 description 1
- 230000007062 hydrolysis Effects 0.000 description 1
- 238000006460 hydrolysis reaction Methods 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
- 239000010808 liquid waste Substances 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 239000000779 smoke Substances 0.000 description 1
- 238000001179 sorption measurement Methods 0.000 description 1
- 235000010215 titanium dioxide Nutrition 0.000 description 1
- JFALSRSLKYAFGM-UHFFFAOYSA-N uranium(0) Chemical compound [U] JFALSRSLKYAFGM-UHFFFAOYSA-N 0.000 description 1
- LEONUFNNVUYDNQ-UHFFFAOYSA-N vanadium atom Chemical compound [V] LEONUFNNVUYDNQ-UHFFFAOYSA-N 0.000 description 1
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for recycling scandium from scandium-containing solid waste generated by extracting titanium by a chlorination process, which comprises the steps of leaching and separating leaching liquid from scandium-containing solid waste generated by extracting titanium by the chlorination process, reducing and precipitating the leaching liquid, and repeating acid leaching-reduction-scandium precipitation procedures according to scandium enrichment conditions of scandium-rich materials; and carrying out alkali-aluminum dissolution, high-temperature roasting treatment and acid leaching on the precipitate, carrying out multistage countercurrent extraction, eluting to remove impurities and back extraction on the leaching solution to obtain scandium hydroxide, dissolving scandium hydroxide by hydrochloric acid, precipitating by oxalic acid, and roasting at high temperature, and finally obtaining scandium oxide product with purity of more than or equal to 99.9 percent after the treatment of the above multiple process steps. The method has the advantages of convenient operation, simple process flow, low production cost, good environmental protection benefit, easy realization of industrialization and high-efficiency recovery of valuable element scandium in scandium-containing solid waste generated by extracting titanium by a chlorination process.
Description
Technical Field
The invention belongs to the technical field of valuable resource recovery, and particularly relates to a method for recovering scandium from scandium-containing solid waste generated by extracting titanium by a chlorination process.
Background
Scandium is a rare earth element, belongs to strategic resources, has abundant global scandium resources, reserves of about 200 ten thousand t, and reserves of scandium of about 65 ten thousand t in China, but more than 75% of scandium is associated with other minerals, and the factors of complex scandium element extraction process, low overall process recovery rate, high cost and the like are the main reasons for causing the high price of scandium products, and scandium is one of the most expensive metals in the world. The scandium ore resources in China are mainly distributed in bauxite, phosphorite ore (containing weathering leaching type phosphorite bed), vanadium titano-magnetite, tungsten ore, rare earth ore and other minerals, the current scandium extraction raw material is mainly derived from secondary resources such as waste liquid or solid waste generated in the comprehensive utilization process of the co-associated minerals, the secondary resources which can be used as scandium extraction raw material comprise uranium ore byproducts, tungsten smelting waste residues, hydrolysis acid waste liquid generated in the production of titanium white by a sulfuric acid method, chlorinated smoke dust generated in the extraction of titanium by boiling chlorination, titanium-containing blast furnace slag, red mud, ion adsorption type rare earth ore, dolomite oblate tailings and the like, and the physical and chemical differences, especially the chemical compositions and the solubility in different leaches of the scandium extraction raw material are large, so that the involved scandium extraction process is different; the traditional scandium element extraction process has the technical defects of complex process, low recovery rate, high cost and the like.
Disclosure of Invention
In order to solve the bottleneck problem of the traditional scandium extraction process, the invention aims to provide the method for recycling scandium from scandium-containing solid waste generated in the titanium extraction process by the chlorination process, which has the advantages of convenient operation, simple process flow, high efficiency, low production cost, good environmental protection benefit, easy realization of industrialization and capability of efficiently recycling valuable element scandium in the scandium-containing solid waste generated in the titanium extraction process by the chlorination process. The scandium-extracting raw material is mainly scandium-containing solid waste generated in the process of extracting titanium by a chlorination method, and the process of extracting titanium by the chlorination method mainly comprises a molten salt chlorination method, a boiling chlorination method and a low-temperature chlorination method at present.
In order to realize the invention, the invention provides a method for recycling scandium from scandium-containing solid waste generated by extracting titanium by a chlorination process, which comprises the following steps:
s1, pulping scandium-containing solid waste with water, dissolving, washing and filtering to obtain leaching liquid, residues and washing water, wherein the washing water is circularly used for pulping and leaching;
s2, keeping the pH value of the leaching solution to be less than or equal to 0.5, and fully reducing target cations in the leaching solution by using a reducing agent;
s3, precipitating scandium ions in the leaching solution subjected to the reduction treatment by using a precipitator to prepare a primary scandium-rich material;
s4, leaching scandium in the primary scandium-rich material by using an acid solution, wherein the leaching end point pH value is less than or equal to 0.5, reducing target ions in the leaching solution by using a reducing agent, precipitating scandium ions in the leaching solution subjected to reduction treatment by using a precipitating agent, filtering to obtain a scandium-rich material, and repeating the acid leaching-reduction-scandium precipitation process according to the scandium enrichment condition of the scandium-rich material, wherein the repetition number is more than or equal to 1, so as to obtain a final scandium-rich material;
s5, countercurrent leaching of aluminum compounds in the final scandium-rich material at room temperature by using a strong alkaline solution, filtering to obtain an aluminum-containing leaching solution and a scandium-rich material after aluminum removal, washing the scandium-rich material after aluminum removal by water until the salinity is less than or equal to 0.5%, filtering to obtain filter residues and washing water, recycling the washing water to the alkali dissolution process, and preparing an alkali solution required by dissolution;
s6, drying the final scandium-rich material subjected to the alkali treatment to a water content of less than 2%, and then roasting at a high temperature, wherein the roasting temperature is more than or equal to 500 ℃, and preferably 600-900 ℃;
finely grinding the scandium-rich material after S7 roasting, and leaching scandium by using an acid solution, wherein the leaching end point pH value is less than or equal to 0.5, so as to obtain scandium leaching solution;
s8, reducing target cations in the leaching solution by using a reducing agent, so as to ensure H in the reduced solution + The concentration is more than or equal to 0.5mol/L;
s9, performing multistage countercurrent extraction on scandium ions in the scandium-containing leaching solution subjected to reduction treatment by using a composite extractant, and separating phases to obtain a scandium-loaded organic phase and a raffinate;
s10, carrying out multistage countercurrent elution on residual metal impurities in the loaded organic phase by using a composite acid solution;
s11, carrying out multistage countercurrent back extraction on the eluted scandium loaded organic phase by using alkali liquor to obtain scandium hydroxide with the total content of impurity elements less than 0.05%, dissolving scandium hydroxide by hydrochloric acid, precipitating by oxalic acid, and roasting at high temperature, and finally obtaining scandium oxide with the purity more than or equal to 99.9%.
In the technical scheme, in S1, the solid ratio of the leaching solution is more than or equal to 0.5:1ml/g, preferably 1-10:1 ml/g; scandium content in the scandium-containing solid waste is more than or equal to 10g/t, and the pH value of the leaching solution is kept to be less than or equal to 2 in the leaching process; during washing, the pH value of the washing water is kept to be less than or equal to 2, and the liquid-solid ratio during washing is the volume of the washing water: residue mass (50% water) =0.5 to 5:1, washing times are more than or equal to 2 times, washing water is returned to the leaching process for use, and residues are used for comprehensive recycling of titanium, carbon and iron.
Further, in S2, the target cation is ferric iron or hexavalent chromium, the reducing agent is metal powder or metal scraps of iron, magnesium, zinc or aluminum, sodium metabisulfite, sodium sulfate and oxalic acid, and the dosage of the reducing agent is 0.1-6 g/L.leaching solution; during reduction, the system temperature of the leaching solution is more than or equal to 40 ℃, preferably 60-90 ℃; in S3, the precipitant is one or more of sodium hydroxide, potassium hydroxide, carbonate, bicarbonate, calcium oxide, calcium hydroxide, magnesium oxide, magnesium hydroxide and ammonia water, and the pH of the precipitation end point is controlled to be 3-6.
Further, in S4, the acid solution for leaching scandium in the primary scandium-rich material is sulfuric acid, hydrochloric acid or nitric acid, the hydrogen ion concentration of the acid solution is 1-8 mol/L, and the solid ratio of the leaching solution is that the volume of the acid solution: primary scandium-rich material mass=0.5-10: 1, the leaching temperature is more than or equal to 20 ℃, preferably 50-95 ℃, the consumption of the reducing agent is 0.01-2 g/L.leaching liquid in the process of repeated acid leaching-reduction-scandium precipitation, the end pH value of the scandium precipitation process is 3-6, and the repeated acid leaching-reduction-scandium precipitation process is stopped when the scandium content of the scandium-rich material is more than or equal to 5%.
Further, in S5, the strong alkaline solution is a solution containing sodium hydroxide or potassium hydroxide, the alkali concentration is 0.1-5 mol/L, and the liquid-solid ratio during leaching is the volume of the alkali solution: final scandium-rich material mass (water content 50%) =1 to 20:1, when the concentration of sodium aluminate in the aluminum-containing leaching solution obtained by filtering after leaching reaches the requirement of industrial aluminum recovery concentration, the method can be used for aluminum recovery.
Further, in S7, the roasted scandium-rich material is required to be ground to be less than or equal to 30 microns in particle size during fine grinding, the acid solution is sulfuric acid, hydrochloric acid or nitric acid, the hydrogen ion concentration of the acid solution is 1-8 mol/L, and the solid ratio of the leaching solution is the volume of the acid solution: primary scandium-rich material mass (water content 50%) =0.5-10: 1, the leaching temperature is more than or equal to 20 ℃, preferably 50-95 ℃.
Further, in S8, the target cation to be reduced is ferric iron, the reducing agent is iron metal powder or metal scraps, the using amount of the reducing agent is 0.01-1 g/L.leaching solution, and the system temperature of the leaching solution during reduction is more than or equal to 40 ℃, preferably 60-90 ℃.
Further, in S9, the composite extractant comprises, by volume, 5-30% of P204, 5-20% of TBP, 0-10% of Cyanex572, and 50-90% of 260# solvent oil or sulfonated kerosene; during extraction, the volume ratio of the composite extractant of the organic phase to the leaching solution after the aqueous phase is reduced is 0.1-20: 1, the extraction stage number is more than or equal to 1, the extraction temperature is more than or equal to 5 ℃, preferably 20-50 ℃, and the extraction equipment is an extraction tank, a cyclone extractor and a microchannel extraction equipment.
Further, in S10, the acid solution includes an acid solution and a strong oxidizer, the acid solution is a sulfuric acid-phosphoric acid mixed solution or a hydrochloric acid-sulfuric acid-phosphoric acid mixed solution, the concentration of hydrogen ions in the acid solution is greater than or equal to 1mol/L, preferably 3-8 mol/L, the concentration of phosphoric acid is greater than or equal to 0.01mol/L, the strong oxidizer is hydrogen peroxide, the concentration of hydrogen peroxide is 0.5-2 mol/L, and the volume ratio of the organic phase to the aqueous phase during elution is 0.5-5: 1, the elution level is more than or equal to 1, the elution temperature is more than or equal to 5 ℃, preferably 20-50 ℃, and the elution equipment is an extraction tank, a cyclone extractor and a microchannel extraction equipment.
Further, in S11, the alkali liquor is NaOH, KOH or ammonia water, the concentration of the NaOH and the KOH is 1-5 mol/L, the concentration of the ammonia water is 1-10 mol/L, and the volume ratio of the loaded organic phase to the alkali liquor during back extraction is 1: 0.5-5, the back extraction temperature is more than or equal to 5 ℃, preferably 20-50 ℃, the back extraction stage number is more than or equal to 1mol/L, preferably 2-6 mol/L, and the liquid-solid ratio of the hydrochloric acid solution is that the volume of the hydrochloric acid solution is: strip mass=1 to 10:1, a step of; the concentration of oxalic acid is 0.1-1 mol/L, and the volume of scandium-containing solution is that during scandium precipitation: oxalic acid solution volume=0.5-1: 0.5-10, the roasting temperature of the precipitate is 700-1000 ℃, and the roasting time is more than or equal to 3 hours, preferably 4-6 hours.
Compared with the prior art, the invention has the beneficial effects that:
according to the method, the scandium-containing solid waste water generated by extracting titanium by the chlorination process is leached to separate leaching liquid, the leaching liquid is reduced and precipitated, the precipitate is subjected to alkali-dissolution aluminum, high-temperature roasting treatment and acid leaching, the leaching liquid is subjected to multistage countercurrent extraction, eluting to remove impurities and back extraction to obtain scandium hydroxide, and scandium oxide products with purity of more than or equal to 99.9% are finally obtained after treatment.
Detailed Description
The invention is further illustrated below in connection with specific examples, but is not limited in any way. For the sake of brevity, the raw materials in the following examples are all commercially available products unless otherwise specified, and the methods used are all conventional methods unless otherwise specified. Unless specifically indicated otherwise, the processing conditions, the operation steps, the equipment used, and the like, which are specifically referred to in the present invention, are all known to those skilled in the art on the basis of the prior art without any inventive work.
The method for recycling scandium from scandium-containing solid waste generated by extracting titanium by a chlorination method comprises the following steps:
s1, pulping scandium-containing solid waste with water, dissolving, washing and filtering to obtain leaching liquid, residues and washing water, wherein the washing water is circularly used for pulping and leaching;
s2, keeping the pH value of the leaching solution to be less than or equal to 0.5, and fully reducing target cations in the leaching solution by using a reducing agent;
s3, precipitating scandium ions in the leaching solution subjected to the reduction treatment by using a precipitator to prepare a primary scandium-rich material;
s4, leaching scandium in the primary scandium-rich material by using an acid solution, wherein the leaching end point pH value is less than or equal to 0.5, reducing target ions in the leaching solution by using a reducing agent, precipitating scandium ions in the leaching solution subjected to reduction treatment by using a precipitating agent, filtering to obtain a scandium-rich material, and repeating the acid leaching-reduction-scandium precipitation process according to the scandium enrichment condition of the scandium-rich material, wherein the repetition number is more than or equal to 1, so as to obtain a final scandium-rich material;
s5, countercurrent leaching of aluminum compounds in the final scandium-rich material at room temperature by using a strong alkaline solution, filtering to obtain an aluminum-containing leaching solution and a scandium-rich material after aluminum removal, washing the scandium-rich material after aluminum removal by water until the salinity is less than or equal to 0.5%, filtering to obtain filter residues and washing water, recycling the washing water to the alkali dissolution process, and preparing an alkali solution required by dissolution;
s6, drying the final scandium-rich material subjected to the alkali treatment to a water content of less than 2%, and then roasting at a high temperature, wherein the roasting temperature is more than or equal to 500 ℃, and preferably 600-900 ℃;
finely grinding the scandium-rich material after S7 roasting, and leaching scandium by using an acid solution, wherein the leaching end point pH value is less than or equal to 0.5, so as to obtain scandium leaching solution;
s8, reducing target cations in the leaching solution by using a reducing agent, so as to ensure H in the reduced solution + The concentration is more than or equal to 0.5mol/L;
s9, performing multistage countercurrent extraction on scandium ions in the scandium-containing leaching solution subjected to reduction treatment by using a composite extractant, and separating phases to obtain a scandium-loaded organic phase and a raffinate;
s10, carrying out multistage countercurrent elution on residual metal impurities in the loaded organic phase by using a composite acid solution;
s11, carrying out multistage countercurrent back extraction on the eluted scandium loaded organic phase by using alkali liquor to obtain scandium hydroxide with the total content of impurity elements less than 0.05%, dissolving scandium hydroxide by hydrochloric acid, precipitating by oxalic acid, and roasting at high temperature, and finally obtaining scandium oxide with the purity more than or equal to 99.9%.
Example 1
Scandium-containing solid waste generated by extracting titanium by a chlorination method obtained by sampling is molten salt chlorination waste salt, and the main components are shown in table 1.
TABLE 1 sampling to obtain the main components of salt chloride waste
Pulping and dissolving molten salt chloride waste salt by water, wherein the liquid-solid ratio is that the volume ml: mass g=5: 1, obtaining leaching liquid, residues and washing water, wherein the dissolution rate of molten salt chloride waste salt in the process is 92%, the leaching rate of scandium in the process is 98%, the scandium concentration in the solution is 29.4mg/L, and the pH value of the leaching liquid is 1.0; adjusting the pH value of the leaching solution to 0.4 by dilute hydrochloric acid, and fully reducing target cationic ferric iron and hexavalent chromium in the leaching solution by using a reducing agent sodium metabisulfite, wherein the adding amount of the sodium metabisulfite is 5 g/L.leaching solution; precipitating scandium ions in the leaching solution subjected to the reduction treatment by using potassium hydroxide with the mass concentration of 10%, wherein the pH of a precipitation end point is 4.1, filtering to obtain a primary scandium-rich material, and the scandium precipitation rate is 99.98%, and the precipitation rate of Mg, fe, mn, ca, zr, V, ti, si, al, cr is 0.005%, 0.6%, 0.01%, 9.32%, 9.55%, 8.12%, 13.11%, 99.77% and 99.59% respectively; leaching with 3mol/L sulfuric acid for 3h, reducing ferric iron with reduced iron powder, precipitating scandium with 10% sodium hydroxide solution, filtering to obtain scandium-rich material, and repeating acid leaching-reduction-scandium precipitation process for 3 times according to scandium enrichment condition of scandium-rich material to obtain final scandium-rich material, wherein scandium recovery rate in the process is 99.76%, and precipitation rates of Mg, fe, mn, ca, zr, V, ti, si, al, cr in the process are respectively 0.03%, 0.15%, 0.11%, 0.08%, 13.22%, 19.66%, 12.89%, 13.64%, 99.86% and 99.72%; dissolving aluminum compounds in the final scandium-rich material at room temperature by using a potassium hydroxide solution with the concentration of 25%, wherein the aluminum leaching rate in the process is 97.95%, the scandium leaching rate is only 0.08%, the filter residue obtained by filtering after the reaction is complete is the scandium-rich material after aluminum removal, and the filtrate is used for recycling aluminum; the final scandium-rich material after the alkali treatment is baked to the water content of 1.5% at 103 ℃, then baked for 4 hours at 600 ℃ to obtain converted oxide, finely ground to the grain size range below 30 microns, and leached by sulfuric acid solution with the solid ratio of 3mol/L, wherein the solid ratio of the leaching solution is the volume of the sulfuric acid solution: scandium-rich material mass=2 after roasting and fine grinding: 1, the scandium leaching rate in the process is 98%, and the leaching rate of Mg, fe, mn, ca, zr, V, ti, si, al, cr is respectively 60.5%, 30.1%, 16.5%, 76.9%, 1.1%, 1.6%, 2.5%, 1.3%, 10.6% and 15.7%, so as to obtain scandium-rich liquid with relatively low impurity content; adding reduced iron powder into scandium-rich liquid, wherein the adding amount of the reduced iron powder is 0.8 g/L.scandium-rich liquid, the temperature is controlled to be 60 ℃ in the reduction process, ferric iron is completely reduced, scandium-containing solution is obtained after filtration, and scandium is basically not lost in the process; scandium was extracted by direct countercurrent rotational flow with 10% P204, 5% TBP, 2% Cyanex572 and 83% 260# solvent of the complex extractant, ratio of organic phase to aqueous phase 2:1, extracting at 20 ℃, for 6min, and for 1min, obtaining scandium loaded organic phase and raffinate after phase separation, wherein scandium extraction rate is 99.99%, and Mg, fe, mn, ca, zr, V, ti, si, al, cr extraction rates are all less than 0.01%; eluting the residual impurities in the loaded organic phase by using a sulfuric acid-phosphoric acid-hydrogen peroxide mixed solution, wherein the concentrations of sulfuric acid, phosphoric acid and hydrogen peroxide are respectively 4mol/L, 0.5mol/L and 1.3mol/L, the elution temperature is 20 ℃, the countercurrent elution stage number is 3, the elution time is 10min, the phase separation time is 1min, and the O/A ratio during elution is 1:1, washing water is obtained after phase separation, and organic phase is loaded after impurity removal, wherein scandium recovery rate in the process is 99.99%, and removal rate of Mg, fe, mn, ca, zr, V, ti, si, al, cr is more than 95%; 3-stage countercurrent rotational flow back extraction is carried out on the loaded organic phase by using a NaOH solution with the concentration of 3mol/L, compared with 1:1, obtaining scandium hydroxide with impurity content less than 0.01%, wherein the back extraction rate of scandium in the process is about 99%; scandium hydroxide is completely dissolved by hydrochloric acid with the concentration of 2mol/L, scandium is precipitated by oxalic acid with the concentration of 0.8mol/L to obtain a precipitate, and the volume ratio of the oxalic acid solution to the scandium-containing solution is 3: and 1, centrifugally filtering, dehydrating and drying the precipitate, roasting at 850 ℃ for 5 hours, and finally obtaining scandium oxide with the purity of 99.98%.
Example 2
Scandium-containing solid waste generated by extracting titanium by a chlorination method obtained by sampling is boiling chlorination dust collection slag, and the main components are shown in table 2.
TABLE 2 obtaining the main components of the boiling chloridized dust collection slag by sampling a certain time
Pulping and dissolving boiling chloridized dust collection slag by water, wherein the liquid-solid ratio is as follows: mass g=3: 1, obtaining leaching liquid, residues and washing water, wherein the dissolution rate of the boiling chloridized dust collection residues in the process is 91%, the scandium leaching rate in the process is 97%, the scandium concentration in the solution is 97.0mg/L, and the pH value of the leaching liquid is 0.7; adjusting the pH value of the leaching solution to 0.47 by dilute hydrochloric acid, and fully reducing target cationic ferric iron and hexavalent chromium in the leaching solution by using reduced iron powder, wherein the adding amount of the reduced iron powder is 5.5 g/L.leaching solution; precipitating scandium ions in the leaching solution subjected to the reduction treatment by using potassium hydroxide with the mass concentration of 5%, wherein the pH of a precipitation end point is 4.3, filtering to obtain a primary scandium-rich material, and the scandium precipitation rate is 99.96%, and the precipitation rate of Mg, fe, mn, ca, zr, V, ti, si, al, cr is 0.01%, 0.5%, 0.01%, 0.02%, 8.21%, 7.65%, 8.11%, 8.99%, 99.89% and 99.49% respectively; leaching with 2.5mol/L sulfuric acid for 4h, reducing ferric iron with reduced iron powder, precipitating scandium with 10% sodium hydroxide solution, filtering to obtain scandium-rich material, and repeating acid leaching-reduction-scandium precipitation process for 3 times according to scandium enrichment condition of scandium-rich material to obtain final scandium-rich material, wherein scandium recovery rate in the process is 99.81%, and precipitation rates of Mg, fe, mn, ca, zr, V, ti, si, al, cr in the process are respectively 0.02%, 0.1%, 0.08%, 0.05%, 11.13%, 15.78%, 10.55%, 11.89%, 99.88% and 99.40%; dissolving aluminum compounds in the final scandium-rich material at room temperature by using a potassium hydroxide solution with the concentration of 25%, wherein the aluminum leaching rate in the process is 98.22%, the scandium leaching rate is only 0.06%, filtering residues obtained after the reaction is completed are scandium-rich materials subjected to aluminum removal, and the filtrate is used for recycling aluminum; the final scandium-rich material after the alkali treatment is baked to the water content of 1.3% at 103 ℃, then baked for 3.5 hours at 650 ℃, the converted oxide is obtained after baking, the converted oxide is finely ground to the grain size range below 30 microns, and then leached by sulfuric acid solution with the solid ratio of 2.5mol/L, wherein the volume of the leaching solution is: scandium-rich material mass=2.5 after roasting and fine grinding: 1, the scandium leaching rate in the process is 97.9%, and the leaching rate of Mg, fe, mn, ca, zr, V, ti, si, al, cr is 66.7%, 24.5%, 17.3%, 60.7%, 0.9%, 1.1%, 2.3%, 1.0%, 9.2% and 11.8% respectively, so that scandium-rich liquid with relatively low impurity content is obtained; adding reduced iron powder into scandium-rich liquid, wherein the adding amount of the reduced iron powder is 0.5g/L of scandium-rich liquid, the temperature is controlled to be 60 ℃ in the reduction process, ferric iron is completely reduced, scandium-containing solution is obtained after filtration, and scandium is basically not lost in the process; scandium was extracted by direct countercurrent rotational flow with 10% P204, 5% TBP, 2% Cyanex572 and 83% 260# solvent of the complex extractant, ratio of organic phase to aqueous phase 2:1, extracting at 20 ℃, for 6min, and for 1min, obtaining scandium loaded organic phase and raffinate after phase separation, wherein scandium extraction rate is 99.99%, and Mg, fe, mn, ca, zr, V, ti, si, al, cr extraction rates are all less than 0.01%; eluting the residual impurities in the loaded organic phase by using sulfuric acid-phosphoric acid-hydrogen peroxide mixed solution, wherein the concentrations of sulfuric acid, phosphoric acid and hydrogen peroxide are respectively 3mol/L, 0.4 mol/L and 1.1mol/L, the elution temperature is 20 ℃, the countercurrent elution stage number is 3, the elution time is 10min, the phase separation time is 1min, and the O/A ratio during elution is 1:1, washing water is obtained after phase separation, and organic phase is loaded after impurity removal, wherein scandium recovery rate in the process is 99.99%, and removal rate of Mg, fe, mn, ca, zr, V, ti, si, al, cr is more than 96%; 3-stage countercurrent rotational flow back extraction is carried out on the loaded organic phase by using a NaOH solution with the concentration of 3mol/L, compared with 1:1, back extraction temperature is 20 ℃, back extraction time is 6min, phase separation time is 5min, scandium hydroxide with impurity content less than 0.02% is obtained, and the back extraction rate of scandium in the process is about 99%; scandium hydroxide is completely dissolved by hydrochloric acid with the concentration of 2mol/L, scandium is precipitated by oxalic acid with the concentration of 0.8mol/L to obtain a precipitate, and the volume ratio of the oxalic acid solution to the scandium-containing solution is 3: and 1, centrifugally filtering, dehydrating and drying the precipitate, roasting at 850 ℃ for 5 hours, and finally obtaining scandium oxide with the purity of 99.96%.
The invention is not a matter of the known technology.
Many possible variations and modifications of the disclosed technology can be made by anyone skilled in the art without departing from the scope of the technology, or the technology can be modified to be equivalent. Therefore, any simple modification, equivalent variation and modification of the above embodiments according to the technical substance of the present invention shall still fall within the scope of the technical solution of the present invention.
Claims (10)
1. The method for recycling scandium from scandium-containing solid waste generated by extracting titanium by a chlorination method is characterized by comprising the following steps of:
s1, pulping scandium-containing solid waste with water, dissolving, washing and filtering to obtain leaching liquid, residues and washing water, wherein the washing water is circularly used for pulping and leaching;
s2, keeping the pH value of the leaching solution to be less than or equal to 0.5, and fully reducing target cations in the leaching solution by using a reducing agent;
s3, precipitating scandium ions in the leaching solution subjected to the reduction treatment by using a precipitator to prepare a primary scandium-rich material;
s4, leaching scandium in the primary scandium-rich material by using an acid solution, wherein the leaching end point pH value is less than or equal to 0.5, reducing target ions in the leaching solution by using a reducing agent, precipitating scandium ions in the leaching solution subjected to reduction treatment by using a precipitating agent, filtering to obtain a scandium-rich material, and repeating the acid leaching-reduction-scandium precipitation process according to the scandium enrichment condition of the scandium-rich material, wherein the repetition number is more than or equal to 1, so as to obtain a final scandium-rich material;
s5, countercurrent leaching of aluminum compounds in the final scandium-rich material at room temperature by using a strong alkaline solution, filtering to obtain an aluminum-containing leaching solution and a scandium-rich material after aluminum removal, washing the scandium-rich material after aluminum removal by water until the salinity is less than or equal to 0.5%, filtering to obtain filter residues and washing water, recycling the washing water to the alkali dissolution process, and preparing an alkali solution required by dissolution;
s6, drying the final scandium-rich material subjected to the alkali treatment to a water content of less than 2%, and then roasting at a high temperature, wherein the roasting temperature is more than or equal to 500 ℃, and is preferably 600-900 ℃;
finely grinding the scandium-rich material after S7 roasting, and leaching scandium by using an acid solution, wherein the leaching end point pH value is less than or equal to 0.5, so as to obtain scandium leaching solution;
s8, reducing target cations in the leaching solution by using a reducing agent, so as to ensure H in the reduced solution + The concentration is more than or equal to 0.5mol/L;
s9, performing multistage countercurrent extraction on scandium ions in the scandium-containing leaching solution subjected to reduction treatment by using a composite extractant, and separating phases to obtain a scandium-loaded organic phase and a raffinate;
s10, carrying out multistage countercurrent elution on residual metal impurities in the loaded organic phase by using a composite acid solution;
s11, carrying out multistage countercurrent back extraction on the eluted scandium loaded organic phase by using alkali liquor to obtain scandium hydroxide with the total content of impurity elements less than 0.05%, dissolving scandium hydroxide by hydrochloric acid, precipitating by oxalic acid, and roasting at high temperature, and finally obtaining scandium oxide with the purity more than or equal to 99.9%.
2. The method according to claim 1, wherein in S1 the leachate solids ratio is not less than 0.5:1ml/g, preferably 1-10:1 ml/g; scandium content in the scandium-containing solid waste is more than or equal to 10g/t, and the pH value of the leaching solution is kept to be less than or equal to 2 in the leaching process; during washing, the pH value of the washing water is kept to be less than or equal to 2, and the liquid-solid ratio during washing is the volume of the washing water: residue mass=0.5 to 5:1, washing times are more than or equal to 2 times, washing water is returned to the leaching process for use, and residues are used for comprehensive recycling of titanium, carbon and iron.
3. The method according to claim 1, wherein in S2, the target cation is ferric iron or hexavalent chromium, the reducing agent is one of metal powder or metal scraps of iron, magnesium, zinc or aluminum, sodium metabisulfite, sodium sulfate and oxalic acid, and the dosage of the reducing agent is 0.1-6 g/l.leaching solution; during reduction, the system temperature of the leaching solution is more than or equal to 40 ℃, preferably 60-90 ℃; in S3, the precipitant is one or more of sodium hydroxide, potassium hydroxide, carbonate, bicarbonate, calcium oxide, calcium hydroxide, magnesium oxide, magnesium hydroxide and ammonia water, and the pH of the precipitation end point is controlled to be 3-6.
4. The method according to claim 1, characterized in that in S4, the acid solution for leaching scandium in the primary scandium-rich material is sulfuric acid, hydrochloric acid or nitric acid, the hydrogen ion concentration of the acid solution is 1-8 mol/L, and the solid ratio of the leaching solution is the volume of the acid solution: primary scandium-rich material mass=0.5-10: 1, the leaching temperature is more than or equal to 20 ℃, preferably 50-95 ℃, the consumption of the reducing agent is 0.01-2 g/L.leaching liquid in the process of repeated acid leaching-reduction-scandium precipitation, the end pH value of the scandium precipitation process is 3-6, and the repeated acid leaching-reduction-scandium precipitation process is stopped when the scandium content of the scandium-rich material is more than or equal to 5%.
5. The method according to claim 1, wherein in S5, the strong alkaline solution is a solution containing sodium hydroxide or potassium hydroxide, the alkali concentration is 0.1-5 mol/L, and the liquid-solid ratio at the time of leaching is the volume of the alkali solution: final scandium-rich material mass=1 to 20:1, when the concentration of sodium aluminate in the aluminum-containing leaching solution obtained by filtering after leaching reaches the requirement of industrial aluminum recovery concentration, the leaching solution is used for aluminum recovery.
6. The method according to claim 1, wherein in S7, the roasted scandium-rich material is required to be ground to a particle size less than or equal to 30 microns during fine grinding, the acid solution is sulfuric acid, hydrochloric acid or nitric acid, the hydrogen ion concentration of the acid solution is 1-8 mol/L, and the solid ratio of the leaching solution is the volume of the acid solution: primary scandium-rich material mass=0.5-10: 1, the leaching temperature is more than or equal to 20 ℃, preferably 50-95 ℃.
7. The method according to claim 1, characterized in that in S8 the target cation to be reduced is ferric iron, the reducing agent is iron metal powder or metal filings, the reducing agent is used in an amount of 0.01-1 g/l.leaching solution, and the temperature of the leaching solution system during reduction is not less than 40 ℃, preferably 60-90 ℃.
8. The method according to claim 1, wherein in S9, the composite extractant comprises 5-30% by volume of P204, 5-20% by volume of TBP, 0-10% by volume of Cyanex572, and 50-90% by volume of 260# solvent oil or sulfonated kerosene; during extraction, the volume ratio of the composite extractant of the organic phase to the leaching solution after the aqueous phase is reduced is 0.1-20: 1, the extraction stage number is more than or equal to 1, the extraction temperature is more than or equal to 5 ℃, preferably 20-50 ℃, and the extraction equipment is an extraction tank, a cyclone extractor and a microchannel extraction equipment.
9. The method according to claim 1, wherein in S10, the composite acid solution comprises an acid solution and a strong oxidizer, the acid solution is a sulfuric acid-phosphoric acid mixed solution or a hydrochloric acid-sulfuric acid-phosphoric acid mixed solution, the concentration of hydrogen ions in the acid solution is equal to or more than 1mol/L, preferably 3-8 mol/L, and the concentration of phosphoric acid is equal to or more than 0.01mol/L; the strong oxidant is hydrogen peroxide, the concentration of the hydrogen peroxide is 0.5-2 mol/L, and the volume ratio of an organic phase to a water phase during elution is 0.5-5: 1, the elution level is more than or equal to 1, the elution temperature is more than or equal to 5 ℃, preferably 20-50 ℃, and the elution equipment is an extraction tank, a cyclone extractor and a microchannel extraction equipment.
10. The method according to claim 1, wherein in S11, the alkali solution is NaOH, KOH or ammonia water, the concentration of NaOH, KOH is 1-5 mol/L, the concentration of ammonia water is 1-10 mol/L, and the volume ratio of the loaded organic phase to the alkali solution during back extraction is 1: 0.5-5, the back extraction temperature is more than or equal to 5 ℃, preferably 20-50 ℃, the back extraction stage number is more than or equal to 1mol/L, preferably 2-6 mol/L, and the liquid-solid ratio of the hydrochloric acid solution is that the volume of the hydrochloric acid solution is: strip mass=1 to 10:1, a step of; the concentration of oxalic acid is 0.1-1 mol/L, and the volume of scandium-containing solution is that during scandium precipitation: oxalic acid solution volume=0.5-1: 0.5-10, the roasting temperature of the precipitate is 700-1000 ℃, and the roasting time is more than or equal to 3 hours, preferably 4-6 hours.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202311637773.6A CN117758072A (en) | 2023-12-01 | 2023-12-01 | Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202311637773.6A CN117758072A (en) | 2023-12-01 | 2023-12-01 | Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CN117758072A true CN117758072A (en) | 2024-03-26 |
Family
ID=90313754
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN202311637773.6A Pending CN117758072A (en) | 2023-12-01 | 2023-12-01 | Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN117758072A (en) |
-
2023
- 2023-12-01 CN CN202311637773.6A patent/CN117758072A/en active Pending
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| US4008076A (en) | Method for processing manganese nodules and recovering the values contained therein | |
| CN110902703A (en) | Method for producing alumina and recovering rare earth elements by using coal ash hydrochloric acid method | |
| CN113955775B (en) | Method for extracting lithium carbonate from lithium-rich clay by acid-base combined method | |
| CN111575502A (en) | Method for extracting nickel element from nickel ore | |
| CN108220631A (en) | A kind of method using aluminum-extracted pulverized fuel ash process condensing crystallizing mother liquor scandium | |
| CN115852177A (en) | Method for recycling scandium from fused salt chlorination dust collection slag | |
| CN114480882B (en) | Method for fully utilizing ferrotitanium and vanadium resources in vanadium titano-magnetite | |
| CN112391537B (en) | Method for extracting vanadium by using hydrochloric acid, sulfuric acid and vanadium-containing high-calcium high-phosphorus slag | |
| CN117758072A (en) | Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process | |
| CN110629043B (en) | Bismuth extraction method based on phase transformation of bismuth sulfide ore | |
| CN114314635A (en) | Method for extracting rare earth and recovering fluorine from bastnaesite high-grade leaching residue | |
| CN117758074A (en) | Synergistic recovery method of scandium in scandium-containing waste liquid and waste residue generated in titanium extraction process flow | |
| CN117737428A (en) | Method for recycling scandium from scandium-containing solid waste generated by extracting titanium by chlorination process | |
| CN117758076A (en) | Method for extracting scandium from scandium-containing solid waste generated by extracting titanium by chlorination process | |
| CN117758081A (en) | Method for extracting scandium from titanium white waste acid and ferric hydroxide slag in cooperation | |
| CN115786744B (en) | Method for extracting scandium by combining titanium white waste acid and fused salt chlorination dust collection slag | |
| CN117737429A (en) | Method for extracting scandium from scandium-containing waste residue generated by extracting titanium by chlorination process | |
| CN117758083A (en) | Method for recycling scandium in scandium-containing waste residues generated by titanium extraction through chlorination | |
| CN117758080A (en) | Method for extracting scandium by combining titanium white waste acid and alkali precipitation waste residue | |
| CN117758073A (en) | Scandium-containing waste liquid and waste residue synergistic scandium extraction method produced in titanium extraction process flow | |
| CN117758075A (en) | Method for extracting scandium from scandium-containing waste residue generated by extracting titanium by titanium white waste acid and chlorination method | |
| CN117758077A (en) | Method for extracting scandium from scandium-containing waste residues generated in titanium chlorination process smelting process | |
| CN115974128B (en) | Method for recycling scandium from fused salt chlorination comprehensive utilization byproducts by utilizing titanium white waste acid | |
| CN117758078A (en) | Method for recycling scandium from scandium-containing waste residues generated in titanium extraction by chlorination process | |
| CN115852179A (en) | Method for extracting scandium from fused salt chlorination dust collecting slag |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| PB01 | Publication | ||
| PB01 | Publication | ||
| SE01 | Entry into force of request for substantive examination | ||
| SE01 | Entry into force of request for substantive examination |