CN116497229A - Method for selectively separating useful elements from anatase - Google Patents
Method for selectively separating useful elements from anatase Download PDFInfo
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- CN116497229A CN116497229A CN202310505508.6A CN202310505508A CN116497229A CN 116497229 A CN116497229 A CN 116497229A CN 202310505508 A CN202310505508 A CN 202310505508A CN 116497229 A CN116497229 A CN 116497229A
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- anatase
- leaching
- filtrate
- scandium
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- GWEVSGVZZGPLCZ-UHFFFAOYSA-N Titan oxide Chemical compound O=[Ti]=O GWEVSGVZZGPLCZ-UHFFFAOYSA-N 0.000 title claims abstract description 91
- 238000000034 method Methods 0.000 title claims abstract description 28
- 238000002386 leaching Methods 0.000 claims abstract description 80
- 239000000706 filtrate Substances 0.000 claims abstract description 69
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 64
- 239000000463 material Substances 0.000 claims abstract description 54
- 229910052706 scandium Inorganic materials 0.000 claims abstract description 54
- SIXSYDAISGFNSX-UHFFFAOYSA-N scandium atom Chemical compound [Sc] SIXSYDAISGFNSX-UHFFFAOYSA-N 0.000 claims abstract description 51
- 239000007788 liquid Substances 0.000 claims abstract description 42
- RTAQQCXQSZGOHL-UHFFFAOYSA-N Titanium Chemical compound [Ti] RTAQQCXQSZGOHL-UHFFFAOYSA-N 0.000 claims abstract description 39
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims abstract description 39
- 229910052782 aluminium Inorganic materials 0.000 claims abstract description 39
- 229910052719 titanium Inorganic materials 0.000 claims abstract description 39
- 239000010936 titanium Substances 0.000 claims abstract description 39
- 238000000926 separation method Methods 0.000 claims abstract description 33
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims abstract description 32
- 229910052742 iron Inorganic materials 0.000 claims abstract description 32
- JTNCEQNHURODLX-UHFFFAOYSA-N 2-phenylethanimidamide Chemical compound NC(=N)CC1=CC=CC=C1 JTNCEQNHURODLX-UHFFFAOYSA-N 0.000 claims abstract description 23
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 23
- 229910000343 potassium bisulfate Inorganic materials 0.000 claims abstract description 23
- 229910052710 silicon Inorganic materials 0.000 claims abstract description 17
- 239000010703 silicon Substances 0.000 claims abstract description 17
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims abstract description 16
- 229910017604 nitric acid Inorganic materials 0.000 claims abstract description 16
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 14
- 239000004408 titanium dioxide Substances 0.000 claims abstract description 11
- 238000000605 extraction Methods 0.000 claims abstract description 10
- 238000002156 mixing Methods 0.000 claims abstract description 10
- 230000004913 activation Effects 0.000 claims abstract description 9
- 238000001816 cooling Methods 0.000 claims abstract description 9
- 238000001704 evaporation Methods 0.000 claims abstract description 8
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 claims abstract description 6
- 229910021578 Iron(III) chloride Inorganic materials 0.000 claims abstract description 4
- RBTARNINKXHZNM-UHFFFAOYSA-K iron trichloride Chemical compound Cl[Fe](Cl)Cl RBTARNINKXHZNM-UHFFFAOYSA-K 0.000 claims abstract description 4
- 230000004927 fusion Effects 0.000 claims abstract description 3
- 230000007062 hydrolysis Effects 0.000 claims abstract description 3
- 238000006460 hydrolysis reaction Methods 0.000 claims abstract description 3
- 238000006386 neutralization reaction Methods 0.000 claims abstract description 3
- 239000000843 powder Substances 0.000 claims abstract description 3
- 239000000377 silicon dioxide Substances 0.000 claims abstract description 3
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical group [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 21
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 claims description 18
- 238000004137 mechanical activation Methods 0.000 claims description 14
- 239000007787 solid Substances 0.000 claims description 11
- 239000012074 organic phase Substances 0.000 claims description 7
- 239000012071 phase Substances 0.000 claims description 6
- 229910018072 Al 2 O 3 Inorganic materials 0.000 claims description 3
- 229910004298 SiO 2 Inorganic materials 0.000 claims description 3
- 239000004566 building material Substances 0.000 claims description 3
- 230000008021 deposition Effects 0.000 claims description 3
- 239000003337 fertilizer Substances 0.000 claims description 3
- 230000008020 evaporation Effects 0.000 claims description 2
- 239000011863 silicon-based powder Substances 0.000 claims description 2
- 239000004411 aluminium Substances 0.000 claims 2
- 238000007654 immersion Methods 0.000 claims 2
- 230000008901 benefit Effects 0.000 abstract description 6
- 239000000047 product Substances 0.000 abstract description 3
- 238000006243 chemical reaction Methods 0.000 description 31
- 239000002002 slurry Substances 0.000 description 11
- 230000000052 comparative effect Effects 0.000 description 6
- 239000012141 concentrate Substances 0.000 description 5
- 238000002474 experimental method Methods 0.000 description 5
- 238000002844 melting Methods 0.000 description 4
- 230000008018 melting Effects 0.000 description 4
- 230000000694 effects Effects 0.000 description 3
- 238000000227 grinding Methods 0.000 description 3
- 239000002245 particle Substances 0.000 description 3
- 238000011084 recovery Methods 0.000 description 3
- 238000004064 recycling Methods 0.000 description 3
- 238000003723 Smelting Methods 0.000 description 2
- 229910052751 metal Inorganic materials 0.000 description 2
- LUKDNTKUBVKBMZ-UHFFFAOYSA-N aluminum scandium Chemical compound [Al].[Sc] LUKDNTKUBVKBMZ-UHFFFAOYSA-N 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 230000003631 expected effect Effects 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- 239000007769 metal material Substances 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 230000001180 sulfating effect Effects 0.000 description 1
Classifications
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B21/00—Obtaining aluminium
- C22B21/0015—Obtaining aluminium by wet processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B1/00—Preliminary treatment of ores or scrap
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B21/00—Obtaining aluminium
- C22B21/0007—Preliminary treatment of ores or scrap or any other metal source
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/065—Nitric acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B34/00—Obtaining refractory metals
- C22B34/10—Obtaining titanium, zirconium or hafnium
- C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
- C22B34/1218—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by dry processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B34/00—Obtaining refractory metals
- C22B34/10—Obtaining titanium, zirconium or hafnium
- C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
- C22B34/1236—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B59/00—Obtaining rare earth metals
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Abstract
The invention provides a method for selectively separating useful elements from anatase, which comprises the following steps: s1, sequentially carrying out mechanical breaking activation and nitric acid pressure leaching on anatase, and then carrying out solid-liquid separation to obtain scandium-containing aluminum liquid and a first filter material, and carrying out neutralization extraction and back extraction on the scandium-containing aluminum liquid to obtain aluminum liquid and scandium-rich materials; s2, adding hydrochloric acid solution into the first filter material to leach iron, and carrying out solid-liquid separation to obtain ferric chloride solution and a second filter material; the main element in the second filter material is silicon and titanium; s3, carrying out solid-liquid separation on the second filter material mixed potassium bisulfate after fusion roasting-water leaching to obtain a first filtrate and active silica powder, and carrying out solid-liquid separation on the first filtrate after high-temperature hydrolysis to obtain titanium dioxide and a second filtrate; and S4, mixing sulfuric acid with the second filtrate, evaporating, concentrating, cooling and crystallizing to obtain potassium bisulfate, and returning the potassium bisulfate to the step S3. The method of the invention has the advantages of high comprehensive utilization rate of resources, high product purity, short process flow, good economic benefit and the like.
Description
Technical Field
The invention relates to the technical field of comprehensive recycling of titanium-containing resources, in particular to a method for selectively separating useful elements from anatase.
Background
Titanium is a widely applied metal material, and is mainly obtained by ore dressing-smelting of titanium-containing ores, and the titanium-containing ores can be roughly divided into: magma type ore, sedimentary type ore, metamorphic ore, residual ore and placer ore. Because of the unique cause of the deposition anatase, the deposition anatase has the characteristics of fine mineral embedding granularity and accompanying element impurities, has large dressing and smelting difficulty, and does not form a set of technical proposal with high resource utilization rate and good economic benefit.
Jiang Peng et al studied the recycling of precipitated anatase in a paper entitled "chemical beneficiation Process of certain precipitated anatase in Sichuan", compared with two processes of sulfating roasting-water leaching and roasting-sulfuric acid leaching, and finally the leaching rates of aluminum, iron and titanium in the leaching solution reach 90%, 90% and 80%, and the expected results are obtained, but the process dissolves all valuable metal elements in the solution, and the problem of separation of each element is faced in the later stage, so that the whole process flow is longer.
In other publications, there is no report on comprehensive recycling of deposited anatase resources.
Disclosure of Invention
Aiming at the problems of long process flow, difficult multi-element separation in leaching solution, low comprehensive recovery rate and the like in the prior art, the invention provides a method for selectively separating aluminum, scandium, iron, titanium and silicon in anatase, and according to the property differences of different ore phases in ores, the method can realize the selective separation and the comprehensive recovery utilization of the elements such as aluminum, scandium, iron, titanium and silicon in anatase, and has the advantages of high comprehensive utilization rate of resources, high product purity, short process flow, good economic benefit and the like.
In order to solve the technical problems, the invention provides the following technical scheme:
a method for selectively separating useful elements from anatase comprising:
s1, sequentially carrying out mechanical breaking activation and nitric acid pressure leaching on anatase, and then carrying out solid-liquid separation to obtain scandium-containing aluminum liquid and a first filter material, and carrying out neutralization extraction and back extraction on the scandium-containing aluminum liquid to obtain aluminum liquid and scandium-rich materials; mechanical activation is adopted, so that the anatase reaction activity is improved, aluminum and scandium are leached, nitric acid is adopted to selectively leach aluminum and scandium, the iron content in the leaching liquid is low, and the scandium extraction selectivity and separation effect are improved.
S2, adding hydrochloric acid solution into the first filter material to leach iron, and carrying out solid-liquid separation to obtain ferric chloride solution and a second filter material; the main element in the second filter material is silicon and titanium; the hydrochloric acid is used for leaching the iron, and the pure ferric chloride solution is obtained through separation and can be sold as a product.
S3, carrying out solid-liquid separation on the second filter material mixed potassium bisulfate after fusion roasting-water leaching to obtain a first filtrate and active silica powder, and carrying out solid-liquid separation on the first filtrate after high-temperature hydrolysis to obtain titanium dioxide and a second filtrate; and potassium bisulfate is added for melting and extracting titanium, so that compared with the titanium extraction by a sulfuric acid method, the leaching effect is better, and meanwhile, the potassium bisulfate can be recycled, so that the consumption of auxiliary materials is reduced.
And S4, mixing sulfuric acid with the second filtrate, evaporating, concentrating, cooling and crystallizing to obtain potassium bisulfate, and returning the potassium bisulfate to the step S3.
Optionally, in step S1, the anatase is a deposited scandium-containing anatase ore, and the mass content of main elements thereof is: tiO (titanium dioxide) 2 ：5.0％～7.0％、Al 2 O 3 :20% -30%, TFe (total iron content): 15.0 to 20.0 percent of SiO 2 ：25％～30％、MgO：1.0％～1.5％、CaO：0.1％～0.5％、MnO：0.1％～0.2％、V 2 O 5 ：0.1％～0.2％、Sc 2 O 3 ：40～60g/t。
Optionally, in step S1, the mechanical activation of anatase is performed by at least one of a disc mill, a ball mill and a rod mill, and the fineness of anatase after mechanical activation is more than 95wt% in proportion to-325 mesh.
Optionally, in step S1, the nitric acid pressure leaching conditions are: the concentration of nitric acid is 20-35 wt%, the solid mass ratio of the leaching solution is 2:1-4:1, the leaching time is 2-5 h, and the leaching temperature is 150-220 ℃.
Optionally, in step S1, the scandium-containing aluminum liquid is extracted by using a mixed extractant of P204 and TBP, the raffinate is aluminum liquid, and the inorganic phase is scandium-rich material after NaOH back extraction of the organic phase extract.
Optionally, in step S2, the conditions of hydrochloric acid leaching of iron are: hydrochloric acid concentration 10-30wt%, leaching liquid solid ratio 2:1-4:1, leaching time 0.5-2 h, leaching temperature 80-100deg.C.
Optionally, in step S3, the apparatus for extracting titanium by melting and roasting is a vacuum furnace, and the conditions are that: the addition amount of potassium bisulfate is 2-3 times (molar mass) of the titanium dioxide content in anatase, the roasting temperature is 400-800 ℃, the roasting time is 2-4 hours, and the mixture is naturally cooled to the room temperature after the roasting is completed.
Optionally, in step S3, before leaching at normal temperature, the roasted material needs to be ground to-325 mesh with a ratio of more than 90%, and the leaching conditions are as follows: the leaching temperature is 10-30 ℃, the leaching time is 1-3 h, and the solid mass ratio of the leaching solution is 2:1-4:1.
Optionally, the active silicon powder has high content of effective silicon, and can be used for preparing fertilizer, building materials and water treatment.
Optionally, in step S3, the evaporation concentration is performed by adding 0.5 to 0.6 times (molar mass) of sulfuric acid.
The technical scheme provided by the invention has the beneficial effects that at least:
1) The invention fully utilizes the property difference of different ore phases in the ore, adopts different leaches to realize the fractional selective separation of multi-metal elements, and separates aluminum scandium in step S1, wherein the total leaching rate of iron, titanium and silicon is less than 0.5 percent; step S2, separating out iron, wherein the leaching rate of titanium and silicon is less than 0.1 percent; and (3) separating titanium in the step (S3), wherein the leaching rate of silicon is less than 0.1%.
2) The silicon-containing ore phase in the ore is subjected to mechanical activation, pressurized nitric acid leaching and roasting mixed treatment, so that the silicon-containing ore phase has high content of effective silicon, can be used for preparing fertilizer, building materials, water treatment and other industries, and has wide application.
3) The utilization rate of each element in the ore is high, and in the step S1 of nitric acid pressure leaching, the leaching rate of aluminum and scandium reaches 93 percent and more than 98 percent respectively; in the step S2 of leaching iron by hydrochloric acid, the leaching rate of the iron can reach more than 98 percent; in the step S3 of melting roasting-water leaching titanium, the leaching rate of titanium can reach more than 95%.
4) Adopts potassium bisulfate to carry out melting extraction of titanium, so that the leaching rate of the titanium is higher, and the potassium bisulfate can be reused.
Drawings
In order to more clearly illustrate the technical solutions of the embodiments of the present invention, the drawings required for the description of the embodiments will be briefly described below, and it is apparent that the drawings in the following description are only some embodiments of the present invention, and other drawings may be obtained according to these drawings without inventive effort for a person skilled in the art.
FIG. 1 is a process flow diagram of a method of the present invention for selectively separating aluminum, scandium, iron, titanium, and silicon from anatase.
Detailed Description
In order to make the objects, technical solutions and advantages of the present invention more apparent, the technical solutions of the present invention will be described in detail with reference to the accompanying drawings and specific embodiments.
Anatase titanium used in comparative examples and examplesThe main element content of the ore is: tiO (titanium dioxide) 2 ：5.4％、Al 2 O 3 ：23.6％、TFe：16.8％、SiO 2 ：28.3％、MgO：1.3％、CaO：0.27％、MnO：0.16％、V2O 5 ：0.11％、Sc 2 O 3 ：45.5g/t。
Comparative example 1
S1, taking 100g of coarse crushed anatase with the average particle size of 0.1mm, pouring the coarse crushed anatase into a disc grinder for mechanical activation for 6min (the fineness of the activated anatase is minus 325 meshes and the proportion of the activated anatase is 95%), pouring 400g of 20% concentration hydrochloric acid into a beaker, setting the reaction temperature to be 80 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter residues, wherein the aluminum content of the filtrate is 15.41g/L, the scandium content of the filtrate is 0.003g/L, the iron content of the filtrate is 45.5g/L, and the leaching rates of aluminum and scandium are 70.1%, 85.2% and 89.65% respectively. Hydrochloric acid is adopted for direct leaching, the purpose of selectively separating aluminum, iron and scandium is not achieved, and the next experiment is not carried out.
Comparative example 2
S1, taking 100g of coarse crushed anatase with the average particle size of 0.1mm, pouring the coarse crushed anatase into a disc grinder for mechanical activation for 6min (the fineness of the activated anatase is minus 325 meshes and the proportion of the activated anatase is 95%), pouring 100g of the activated anatase and 400g of 30% concentration sulfuric acid into a beaker, setting the reaction temperature to 120 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter residues, wherein the aluminum content of the filtrate is 20.41g/L, the scandium content of the filtrate is 0.003g/L, the iron content of the filtrate is 45.5g/L, the titanium content of the filtrate is 8.21g/L, and the leaching rates of aluminum, scandium, iron and titanium are 87.21%, 84.51%, 90.65% and 74.11% respectively. Sulfuric acid pressure leaching is adopted, the purpose of selectively separating aluminum, scandium, iron and titanium is not achieved, and the next experiment is not carried out.
Comparative example 3
S1, taking 100g of anatase with the average grain diameter of 0.1mm after coarse crushing, pouring the anatase into a disc grinder for mechanical activation for 3min (the fineness of the anatase after activation is-200 meshes and the proportion of the anatase to 95%), pouring 400g of nitric acid with the concentration of 20 percent and 100g of the anatase into a pressurized reaction kettle, setting the reaction temperature to 180 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter material 1, wherein the aluminum content of the filtrate is 15.41g/L, the scandium content of the filtrate is 0.003g/L, and the leaching rates of aluminum and scandium are 70.1 percent and 85.2 percent respectively. The leaching rate of aluminum and scandium does not reach the expected effect, and the next experiment is not carried out. Indicating that the recovery of useful elements in anatase cannot be performed when the fineness of the anatase after activation does not meet the conditions.
Comparative example 4
S1, taking 100g of anatase with the average grain diameter of 0.1mm after coarse crushing, pouring the anatase into a disc grinder for mechanical activation for 5min (the fineness of the anatase after activation is minus 325 meshes and the proportion of the anatase to 95%), pouring 400g of nitric acid with the concentration of 30 percent and 100g of the anatase into a pressurized reaction kettle, setting the reaction temperature to 180 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter material 1, wherein the aluminum content of the filtrate is 27.42g/L, the scandium content of the filtrate is 0.004g/L, and the leaching rates of aluminum and scandium are 94.5 percent and 98.4 percent respectively. The filtrate is extracted by P204+TBP, and the organic phase is back extracted by sodium hydroxide solution, thus obtaining scandium concentrate with scandium content of 0.5 percent.
S2, adding 200g of 20% hydrochloric acid into the filter material 1 obtained in the step S1, wherein the solid mass ratio of the leaching solution is 2.3:1, reacting for 2 hours at the temperature of 30 ℃, obtaining filtrate and filter material 2 after the reaction is finished, wherein the iron content in the filtrate is 25g/L, and the iron leaching rate is only 82%. The iron leaching rate is not expected, and the next experiment is not performed.
Comparative example 5
S1, taking 100g of anatase with the average grain diameter of 0.1mm after coarse crushing, pouring the anatase into a disc grinder for mechanical activation for 5min (the fineness of the anatase after activation is minus 325 meshes and the proportion of the anatase to 95%), pouring 400g of nitric acid with the concentration of 30 percent and 100g of the anatase into a pressurized reaction kettle, setting the reaction temperature to 180 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter material 1, wherein the aluminum content of the filtrate is 26.55g/L, the scandium content of the filtrate is 0.0038g/L, and the leaching rates of aluminum and scandium are 94.11 percent and 97.93 percent respectively. The filtrate is extracted by P204+TBP, and the organic phase is back extracted by sodium hydroxide solution, thus obtaining scandium concentrate with scandium content of 0.6 percent.
S2, adding 200g of 30% hydrochloric acid into the filter material 1 obtained in the step S1, wherein the solid mass ratio of the leaching solution is 2.3:1, reacting for 2 hours at the temperature of 80 ℃, obtaining filtrate and filter material 2 after the reaction is finished, wherein the iron content in the filtrate is 65g/L, and the iron leaching rate is 96%.
S3, directly mixing and adding 30g of potassium bisulfate according to the molar mass which is 2.5 times that of the titanium in the filter material 2 obtained in the step S2, roasting at 300 ℃ for 2 hours, pouring the material into 400ml of water when the roasting is finished, leaching for 2 hours, and carrying out solid-liquid separation after the leaching is finished to obtain a filter liquor 1 and the filter material, wherein the titanium content in the filter liquor 1 is 2g/L, the titanium leaching rate is only 10.1%, the titanium leaching rate is not expected, and the next experiment is not carried out.
Example 1
S1, taking 100g of anatase with the average grain diameter of 0.1mm after coarse crushing, pouring the anatase into a disc grinder for mechanical activation for 5min (the fineness of the anatase after activation is minus 325 meshes and the proportion of the anatase to 95%), pouring 400g of nitric acid with the concentration of 30 percent and 100g of the anatase into a pressurized reaction kettle, setting the reaction temperature to 180 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter material 1, wherein the aluminum content of the filtrate is 26.55g/L, the scandium content of the filtrate is 0.0038g/L, and the leaching rates of aluminum and scandium are 94.11 percent and 97.93 percent respectively. The filtrate is extracted by P204+TBP, and the organic phase is back extracted by sodium hydroxide solution, thus obtaining scandium concentrate with scandium content of 0.6 percent.
S2, adding 200g of 30% hydrochloric acid into the filter material 1 obtained in the step S1, reacting for 2 hours at the temperature of 80 ℃ with the solid mass ratio of the leaching solution being 2.3:1, and obtaining filtrate and filter material 2 after the reaction is finished, wherein the iron content in the filtrate is 65g/L, and the iron leaching rate is 96%.
S3, directly mixing and adding 30g of potassium bisulfate into the filter material 2 obtained in the step S2 according to the molar mass which is 2.5 times that of titanium, roasting for 2 hours at the temperature of 500 ℃, grinding until the fineness of the roasted material is minus 325 meshes and the proportion is 90% after naturally cooling to the room temperature, mixing 200ml of water, leaching for 1 hour at the temperature of 25 ℃, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain a filtrate 1 and filter residues, wherein the titanium content in the filtrate 1 is 12.56g/L, and the leaching rate of titanium is 94.21%; the silicon content in the filter residue is 41.23%, the filtrate 1 is hydrolyzed for 1h at the temperature of 100 ℃, then the filtrate 2 and the filter material are obtained through solid-liquid separation, and the obtained filter material is titanium dioxide with the content of 98.5%.
S4, adding 25g of 98% sulfuric acid into the filtrate 3 in the step S3, and evaporating, concentrating, cooling and crystallizing to obtain the potassium bisulfate with the purity of 98%.
Example 2
S1, taking 100g of coarse crushed anatase with the average particle size of 0.1mm, pouring the coarse crushed anatase into a disc grinder for mechanical activation for 5min (the fineness of the activated anatase is minus 325 meshes and the proportion of the activated anatase is 95%), pouring 300g of 30% concentration nitric acid into a pressurized reaction kettle, setting the reaction temperature to 180 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter material 1, wherein the aluminum content of the filtrate is 23.55g/L, the scandium content of the filtrate is 0.0038g/L, and the leaching rates of aluminum and scandium are 89.45% and 90.61% respectively. The filtrate is extracted by P204+TBP, and the organic phase is back extracted by sodium hydroxide solution, thus obtaining scandium concentrate with scandium content of 0.55 percent.
S2, adding 200g of 30% hydrochloric acid into the filter material 1 obtained in the step S1, wherein the solid mass ratio of the leaching solution is 2.3:1, reacting for 2 hours at the temperature of 80 ℃, obtaining filtrate and filter material 2 after the reaction is finished, wherein the iron content in the filtrate is 63.51g/L, and the iron leaching rate is 94.57%.
S3, directly mixing and adding 25g of potassium bisulfate into the filter material 2 obtained in the step S2 according to the molar mass which is 2 times of the titanium content, roasting for 2 hours at the temperature of 500 ℃, grinding until the fineness of the roasted material is minus 325 meshes and accounts for 90 percent after naturally cooling to the room temperature, mixing 200ml of water, leaching for 1 hour at the temperature of 25 ℃, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate 1 and filter residues, wherein the titanium content in the filtrate 1 is 11.25g/L, and the titanium leaching rate is 91.21 percent. The silicon content in the filter residue is 40.55%, the filtrate 1 is hydrolyzed for 1h at the temperature of 120 ℃, then the filtrate 2 and the filter material are obtained through solid-liquid separation, and the obtained filter material is titanium dioxide with the content of 99.10%.
S4, adding 25g of 98% sulfuric acid into the filtrate 3 in the step S3, and evaporating, concentrating, cooling and crystallizing to obtain the potassium bisulfate with the purity of 98.5%.
Example 3
S1, taking 100g of anatase with the average grain diameter of 0.1mm after coarse crushing, pouring the anatase into a disc grinder for mechanical activation for 5min (the fineness of the anatase after activation is minus 325 meshes and the proportion of the anatase to 95%), pouring 400g of nitric acid with the concentration of 30 percent and 100g of the anatase into a pressurized reaction kettle, setting the reaction temperature to 180 ℃, reacting for 4 hours, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain filtrate and filter material 1, wherein the aluminum content of the filtrate is 27.11g/L, the scandium content of the filtrate is 0.0039g/L, and the leaching rates of aluminum and scandium are 95.05 percent and 99.3 percent respectively. The filtrate is extracted by P204+TBP, and the organic phase is back extracted by sodium hydroxide solution, thus obtaining scandium concentrate with scandium content of 0.58 percent.
S2, adding 200g of 30% hydrochloric acid into the filter material 1 obtained in the step S1, reacting for 2 hours at 90 ℃ with the solid mass ratio of the leaching solution being 2.3:1, and obtaining filtrate and filter material 2 after the reaction is finished, wherein the iron content in the filtrate is 72g/L, and the iron leaching rate is 99.1%.
S3, directly mixing and adding 30g of potassium bisulfate into the filter material 2 obtained in the step S2 according to the molar mass which is 2.5 times that of titanium, roasting for 2 hours at the temperature of 600 ℃, grinding until the fineness of the roasted material is minus 325 meshes and the proportion is 90% after naturally cooling to the room temperature, mixing 200ml of water, leaching for 1 hour at the temperature of 25 ℃, pouring out slurry after the reaction is finished, and carrying out solid-liquid separation to obtain a filtrate 1 and filter residues, wherein the titanium content in the filtrate 1 is 15.01g/L, and the leaching rate of titanium is 95.34%; the silicon content in the filter residue is 43.11%, the filtrate 1 is hydrolyzed for 1h at the temperature of 100 ℃, then the filtrate 2 and the filter material are obtained through solid-liquid separation, and the obtained filter material is titanium dioxide with the content of 99.10%.
S4, adding 25g of 98% sulfuric acid into the filtrate 3 in the step S3, and evaporating, concentrating, cooling and crystallizing to obtain the potassium bisulfate with the purity of 98%.
The foregoing is merely illustrative of the present invention, and the present invention is not limited thereto, and any person skilled in the art will readily recognize that variations or substitutions are within the scope of the present invention. Therefore, the protection scope of the present invention shall be subject to the protection scope of the claims.
Claims (10)
1. A method for selectively separating useful elements from anatase comprising:
s1, sequentially carrying out mechanical breaking activation and nitric acid pressure leaching on anatase, and then carrying out solid-liquid separation to obtain scandium-containing aluminum liquid and a first filter material, and carrying out neutralization extraction and back extraction on the scandium-containing aluminum liquid to obtain aluminum liquid and scandium-rich materials;
s2, adding hydrochloric acid solution into the first filter material to leach iron, and carrying out solid-liquid separation to obtain ferric chloride solution and a second filter material; the main element in the second filter material is silicon and titanium;
s3, carrying out solid-liquid separation on the second filter material mixed potassium bisulfate after fusion roasting-water leaching to obtain a first filtrate and active silica powder, and carrying out solid-liquid separation on the first filtrate after high-temperature hydrolysis to obtain titanium dioxide and a second filtrate;
and S4, mixing sulfuric acid with the second filtrate, evaporating, concentrating, cooling and crystallizing to obtain potassium bisulfate, and returning the potassium bisulfate to the step S3.
2. The method according to claim 1, wherein in the step S1, the anatase is a deposition type scandium-containing anatase ore, and the mass contents of main elements thereof are: tiO (titanium dioxide) 2 :5.0％～7.0％、Al 2 O 3 :20％～30％、TFe:15.0％～20.0％、SiO 2 :25％～30％、MgO:1.0％～1.5％、CaO:0.1％～0.5％、MnO:0.1％～0.2％、V 2 O 5 :0.1％～0.2％、Sc 2 O 3 :40～60g/t。
3. The method according to claim 1, wherein in the step S1, the mechanical activation of anatase is performed by at least one of a disc mill, a ball mill and a rod mill, and the fineness of anatase after the mechanical activation is more than 95% in a proportion of-325 mesh.
4. The method according to claim 1, characterized in that in step S1, the nitric acid pressure leaching conditions are: nitric acid concentration is 20% -35%, the solid ratio of the leaching solution is 2:1-4:1, leaching time is 2-5 h, and leaching temperature is 150-220 ℃.
5. The method according to claim 1, wherein in step S1, the scandium-containing aluminium liquid is extracted with a mixed extractant of P204 and TBP, the raffinate is aluminium liquid, and the organic phase extract is NaOH back extracted to obtain the inorganic phase scandium-rich material.
6. The method according to claim 1, wherein in step S2, the hydrochloric acid leached iron conditions are: hydrochloric acid concentration 10% -30%, leaching solution solid ratio 2:1-4:1, leaching time 0.5-2 h, leaching temperature 80-100 ℃.
7. The method according to claim 1, wherein in step S3, the apparatus for extracting titanium by melt-roasting is a vacuum furnace, provided that: the addition amount of the potassium bisulfate is 2-3 times that of the anatase, the titanium dioxide content is 2-4 hours after the roasting is finished, and the temperature is 400-800 ℃ and the roasting time is 2-4 hours, and the potassium bisulfate is naturally cooled to the room temperature state after the roasting is finished.
8. The method according to claim 1, wherein in step S3, the baked material is ground to-325 mesh with a ratio of more than 90% before water immersion at normal temperature, and water immersion conditions are: leaching temperature is 10-30 ℃, leaching time is 1-3 h, and leaching solution solid ratio is 2:1-4:1.
9. The method of claim 1, wherein the active silicon powder has a high content of available silicon and can be used for preparing fertilizers, building materials and water treatment.
10. The method according to claim 1, wherein in step S3, the evaporation concentration is performed by adding 0.5 to 0.6 times of sulfuric acid.
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