Method for cooperatively extracting scandium from scandium-containing waste acid and waste residue

Abstract

The application discloses a method for cooperatively extracting scandium from scandium-containing waste acid and waste residue, which comprises the following steps: adopting a sulfuric acid solution to carry out one-stage leaching on the scandium-containing waste residue, and then adopting a sulfuric acid method titanium white waste acid to carry out two-stage leaching; using a reducing agent to fully reduce target cations in the scandium-containing leaching solution, adding a phosphate solution to remove zirconium, and using a precipitant to prepare a scandium-containing precipitate; mixing the scandium-containing precipitate with sodium carbonate and calcium oxide, and then carrying out high-temperature calcination, so as to remove aluminum, and according to the scandium content, repeatedly carrying out the acid leaching-reduction-scandium precipitation process, so as to obtain a final scandium-rich material; using oxalic acid to countercurrently wash the final scandium-rich material, and then filtering to obtain oxalic acid washing water and scandium oxalate; and the scandium oxalate is subjected to high-temperature calcination, and finally, scandium oxide with a purity of greater than or equal to 99.9% is obtained. The method has the advantages of convenient operation, simple technological process, low production cost, good environmental protection benefit, easy industrialization, and efficient recovery of valuable elements scandium in the scandium-containing waste residue.

Description

Technical Field

[0001] This invention belongs to the field of valuable resource recycling technology, specifically relating to a method for the co-extraction of scandium from scandium-containing waste acid and waste residue. Background Technology

[0002] Scandium is a rare earth element and a strategic resource. Globally, scandium resources are abundant, with reserves of approximately 2 million tons. my country's scandium reserves are approximately 650,000 tons. However, more than 75% of scandium is associated with other minerals. The complex extraction process, low overall recovery rate, and high cost are the main reasons for the high price of scandium products. Scandium is known as one of the most expensive metals in the world. Domestic scandium resources are mainly distributed in bauxite and phosphate rock (including weathered leaching phosphate deposits), vanadium-titanium magnetite, tungsten, and rare earth minerals. Currently, scandium extraction raw materials mainly come from secondary resources such as waste liquid or solid waste generated during the comprehensive utilization of associated minerals. Existing secondary resources that can be used as scandium extraction raw materials include by-products of uranium ore, tungsten smelting slag, hydrolyzed acidic waste liquid generated from the sulfuric acid process for titanium dioxide production, chlorinated dust generated from boiling chlorination for titanium extraction, titanium-containing blast furnace slag, red mud, ion-adsorption type rare earth minerals, and Bayan Obo tailings. Different scandium-containing raw materials have different physicochemical properties, especially in terms of chemical composition and solubility in different leaching agents, which leads to different scandium extraction processes. Existing traditional scandium element extraction processes have technical defects such as complex processes, low recovery rates, and high costs. Summary of the Invention

[0003] To address the bottlenecks in traditional scandium extraction processes, this invention aims to provide a method for the co-extraction of scandium from scandium-containing waste acid and waste residue. The raw materials for scandium extraction are scandium-containing waste acid and a primary scandium-containing waste residue. The waste residue relates to the iron-manganese slag mentioned in a Chinese patent application (CN105883911A) regarding a comprehensive utilization method for molten salt chlorination waste. Investigations have shown that scandium in molten salt chlorination waste is primarily concentrated in the aforementioned iron-manganese slag. This iron-manganese slag has a simpler composition and higher scandium content compared to molten salt chlorination waste, making it a high-quality raw material for scandium extraction. However, existing technologies do not efficiently utilize this iron-manganese slag, resulting in resource waste. This technical solution can effectively recover scandium from scandium-containing waste residue generated during titanium extraction via the chlorination process. Furthermore, this method is convenient to operate, has a simple process flow, high efficiency, low production cost, good environmental benefits, and is easily industrialized.

[0004] To achieve the above-mentioned invention, the present invention provides a method for co-extracting scandium from scandium-containing waste acid and waste residue, the method comprising the following steps:

[0005] S1 first undergoes a first-stage leaching process, using sulfuric acid solution to rapidly leach iron, manganese, magnesium, aluminum, titanium, and chromium elements from the scandium-containing waste residue. The leaching time is strictly controlled to be ≤30 min, preferably 20–30 min. After filtration, a leachate and scandium-enriched residue are obtained. Then, a second-stage leaching process is performed, using sulfuric acid from titanium dioxide production to leach the scandium-enriched residue. After filtration, a scandium-containing leachate and filter residue are obtained. The filter residue is returned to the second-stage leaching process for further leaching.

[0006] S2 uses a reducing agent to fully reduce the target cation in the scandium-containing leachate, filters to obtain the reduced scandium-containing solution and residue, returns the residue to the reduction process, adds phosphate solution to the scandium-containing solution to remove zirconium, filters to obtain zirconium phosphate precipitate and the scandium-containing solution after zirconium removal;

[0007] S3 uses a precipitant to precipitate scandium ions in the scandium-containing solution after zirconium removal, thus obtaining a scandium-containing precipitate;

[0008] S4 The above scandium-containing precipitate is thoroughly mixed with sodium or potassium carbonate and calcium oxide, and the mixture is calcined at high temperature. The calcined product is washed with water, filtered to obtain chromium-containing wash water and filter residue. The filter residue is then washed with dilute acid, filtered to obtain aluminum-containing solution and scandium-enriched filter residue.

[0009] S5 uses an acid solution to leach the scandium enriched in the filter residue. After filtration, a scandium-containing filtrate and leaching residue are obtained. The leaching residue is returned to the acid solution scandium leaching process. Then, a reducing agent is used to fully reduce the target cation in the scandium-containing leaching solution. A precipitating agent is used to precipitate the scandium ions in the reduced leaching solution. After filtration, a filtrate and primary scandium-rich material are obtained. The above acid leaching-reduction-scandium precipitation process is repeated according to the scandium content of the scandium-rich material. The number of repetitions is ≥1 to obtain the final scandium-rich material.

[0010] S6 uses oxalic acid to countercurrently wash the above-mentioned ultimate scandium-rich material, and after filtration, oxalic acid wash water and scandium oxalate are obtained. The scandium oxalate is then calcined at high temperature to finally obtain scandium oxide with a purity of ≥99.9%.

[0011] In the above technical solution, further, in S1, the concentration of the sulfuric acid solution used for the first-stage leaching is ≥0.5mol / L, preferably 1~3.5mol / L; the scandium content in the scandium-containing waste residue is ≥50g / t, the leaching temperature is controlled at 20~50℃, the leaching time is strictly controlled at ≤30min, preferably 20~30min, and the liquid-solid ratio during leaching is sulfuric acid solution volume (ml): scandium-containing waste residue mass (g) = 1~10:1; the scandium content of the scandium-containing waste acid (sulphuric acid process titanium dioxide waste acid) in the second-stage leaching process is ≥5mg / L, the leaching temperature is ≥20℃, preferably 50~98℃, the leaching time is ≥1h, preferably 3~6h, and the liquid-solid ratio during leaching is sulfuric acid process titanium dioxide waste acid volume (ml): scandium enrichment residue mass (g) = 1~5:1.

[0012] Furthermore, in S2, the target cation to be reduced is ferric iron, and the reducing agent is one of iron, magnesium, zinc or aluminum metal powder or metal scrap, sodium metabisulfite, or sodium sulfate. The amount of reducing agent used is 1-20 g / L of scandium-containing solution, and the reduction temperature is ≥40℃, preferably 60-95℃. The soluble phosphate used for zirconium removal is sodium phosphate, sodium monohydrogen phosphate, sodium dihydrogen phosphate, potassium phosphate, potassium monohydrogen phosphate, or potassium dihydrogen phosphate. The amount of soluble phosphate added per liter of scandium-containing leachate is 0.06-2 g, and it is prepared as a solution before addition.

[0013] Furthermore, 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 at the precipitation endpoint is controlled between 3 and 6.

[0014] Furthermore, in S4, the mass ratio of the scandium-containing precipitate, sodium carbonate, and calcium oxide is 0.5–2:0.5–5:0.25–3; the scandium-containing precipitate is pre-dried to a moisture content ≤3%, and after drying, it is crushed to a particle size ≤2 cm; during high-temperature calcination, the temperature is increased to ≥800℃ at a rate of 10–40℃ / min, and held at this temperature for ≥3 hours, preferably at 900–1300℃ for 5–8 hours; during water washing, the liquid-to-solid ratio is water volume (ml): calcined mixture (g) = 1–10:1; during dilute acid washing of the filter residue, the dilute acid is an aqueous solution of sulfuric acid or hydrochloric acid with a pH of 0.5–2, and the liquid-to-solid ratio during washing is dilute acid volume (ml): filter residue mass (g) = 1–30:1.

[0015] Furthermore, in S5, the acid solution is one of sulfuric acid, hydrochloric acid, and nitric acid, with a hydrogen ion concentration ≥1 mol / L, preferably 2-7 mol / L. During leaching, the liquid-to-solid ratio is acid solution volume: scandium enriched filter residue mass (dry basis) = 1-10:1 (ml:g). The target cation to be reduced is ferric iron, and the reducing agent is one of iron, magnesium, zinc, or aluminum metal powder or metal scraps, sodium metabisulfite, or sodium sulfate. The amount of reducing agent is 0.1-3 g / L of leaching solution, and the temperature of the leaching solution system during reduction is ≥40℃, preferably 60-95℃. The precipitant is one or more of sodium hydroxide, potassium hydroxide, carbonate, bicarbonate, calcium oxide, calcium hydroxide, magnesium oxide, magnesium hydroxide, and ammonia water. The pH at the precipitation endpoint is controlled at 3-6. When the scandium content of the scandium-containing precipitate is ≥5%, the repeated acid leaching-reduction-scandium precipitation process is stopped to obtain a final scandium-rich material with a higher scandium content.

[0016] Furthermore, in S6, the concentration of oxalic acid used for countercurrent washing is ≥0.1mol / L, preferably 0.3~1mol / L. The liquid-solid ratio during washing is oxalic acid solution volume (ml): final scandium-rich material (dry basis) weight (g) = 1~10:1. The number of countercurrent washing stages is ≥4, preferably 5~10 stages. After washing and filtration, the material is calcined. During calcination, the temperature is increased to 105~125℃ at a rate of 5~10℃ / min and kept at this temperature to dry the precipitate until its moisture content is ≤3%. The temperature is then increased to ≥700℃ at a rate of 10~40℃ / min and kept at this temperature for ≥3h. The preferred calcination temperature is 800~1000℃ and the preferred holding time is 5~8h.

[0017] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0018] This invention utilizes the synergistic extraction of scandium from scandium-containing waste acid and waste residue to ultimately obtain scandium oxide products with a purity ≥ 99.9%. This method is convenient to operate, has a simple process flow, low production cost, good environmental benefits, and is easy to industrialize. It can efficiently recover valuable elements such as scandium from scandium-containing waste residue. Detailed Implementation

[0019] The present invention will be further described below with reference to specific embodiments, but this does not limit the invention in any way. To avoid redundancy, unless otherwise specified, the raw materials used in the following embodiments are commercially available products, and the methods used are conventional methods unless otherwise specified. Unless otherwise specified, the specific processing conditions, operating steps, equipment, and instruments involved in this invention are all things that those skilled in the art can learn from the prior art without creative effort.

[0020] A method for co-extracting scandium from scandium-containing waste acid and waste residue, the method comprising the following steps:

[0021] S1 first undergoes a first-stage leaching process, using sulfuric acid solution to rapidly leach iron, manganese, magnesium, aluminum, titanium, and chromium elements from the scandium-containing waste residue. The leaching time is strictly controlled to be ≤30 min, preferably 20–30 min. After filtration, a leachate and scandium-enriched residue are obtained. Then, a second-stage leaching process is performed, using sulfuric acid from titanium dioxide production to leach the scandium-enriched residue. After filtration, a scandium-containing leachate and filter residue are obtained. The filter residue is returned to the second-stage leaching process for further leaching.

[0022] S2 uses a reducing agent to fully reduce the target cation in the scandium-containing leachate, filters to obtain the reduced scandium-containing solution and residue, returns the residue to the reduction process, adds phosphate solution to the scandium-containing solution to remove zirconium, filters to obtain zirconium phosphate precipitate and the scandium-containing solution after zirconium removal;

[0023] S3 uses a precipitant to precipitate scandium ions in the scandium-containing solution after zirconium removal, thus obtaining a scandium-containing precipitate;

[0024] S4 The above scandium-containing precipitate is thoroughly mixed with sodium or potassium carbonate and calcium oxide, and the mixture is calcined at high temperature. The calcined product is washed with water, filtered to obtain chromium-containing wash water and filter residue. The filter residue is then washed with dilute acid, filtered to obtain aluminum-containing solution and scandium-enriched filter residue.

[0025] S5 uses an acid solution to leach the scandium in the scandium enrichment filter residue. After filtration, a scandium-containing filtrate and leaching residue are obtained. The leaching residue is returned to the acid solution scandium leaching process. Then, a reducing agent is used to fully reduce the target cation in the scandium-containing leaching solution. A precipitating agent is used to precipitate the scandium ions in the reduced leaching solution. After filtration, a filtrate and primary scandium-rich material can be obtained. The above acid leaching-reduction-scandium precipitation process is repeated according to the scandium content of the scandium-rich material. The number of repetitions is ≥1 to obtain the final scandium-rich material.

[0026] S6 uses oxalic acid to countercurrently wash the above-mentioned ultimate scandium-rich material, and after filtration, oxalic acid wash water and scandium oxalate are obtained. The scandium oxalate is then calcined at high temperature to finally obtain scandium oxide with a purity of ≥99.9%.

[0027] Example

[0028] The main components of the titanium dioxide waste acid and residue obtained from a certain sampling are shown in Table 1 and Table 2.

[0029] Table 1. Composition of Waste Acid from Titanium Dioxide Production

[0030]

[0031] Table 2. Waste residue composition table

[0032]

[0033] S1 first undergoes a leaching process: Iron, manganese, magnesium, aluminum, titanium, and chromium elements in waste residue with a scandium content of 160 g / t are rapidly leached using a 20% sulfuric acid solution. The leaching time is strictly controlled at 20 min, the leaching temperature at 30℃, and the liquid-to-solid ratio during leaching is sulfuric acid solution volume (ml): scandium-containing waste residue mass (g) = 2.5:1. During the leaching process, stirring is intensified until no solid sediment remains at the bottom of the reactor. After leaching, the scandium leaching rate is 5%. The leaching of Mg, Fe, Mn, Ca, Zr, V, Ti, Si, Al, Cr (iron, manganese, magnesium, aluminum, titanium, and chromium) is also carried out. The extraction rates reached 91%, 95%, 90%, 3%, 5%, 6%, 5%, 6%, 90%, and 92%, respectively. After filtration, leachate and scandium-enriched residue were obtained. Then, a second-stage leaching was performed: the scandium-enriched residue was leached with sulfuric acid waste acid containing 10 mg / L of scandium from the titanium dioxide process. The leaching temperature was 90℃, the leaching time was 5 hours, and the liquid-to-solid ratio during leaching was 2:1 (volume of sulfuric acid waste acid in ml : mass of scandium-enriched residue in g). After filtration, scandium-containing leachate and filter residue were obtained. The filter residue was returned to the second-stage leaching process for further leaching. The scandium extraction rate in this process was 99.69%.

[0034] S2 uses reduced iron powder to fully reduce the ferric ions in the filtrate. The amount of iron powder used is 11 g / L of scandium-containing leachate. The temperature of the scandium-containing leachate system during reduction is 65°C. During the reduction process, the mixture is stirred vigorously until no solids are deposited at the bottom of the reactor. The filtrate and residue are obtained after filtration. The residue is returned to the reduction process for recycling. Disodium hydrogen phosphate solution is added to the filtrate to remove zirconium. The amount of disodium hydrogen phosphate added is 1.1 g. Before adding, it is prepared as a solution. After filtration, zirconium phosphate precipitate and scandium-containing solution after zirconium removal are obtained. The zirconium precipitation rate is 99.5% and the scandium precipitation rate is <0.01%.

[0035] S3 uses a 10% NaOH solution to adjust the pH of the scandium-containing solution after zirconium removal to 3.9. The system temperature is 50°C. During the pH adjustment process, stirring is intensified until no solids are deposited at the bottom of the reactor. After filtration, scandium-containing precipitate and residual liquid are obtained. The scandium precipitation rate in this process is 99.99%.

[0036] S4 dried the scandium-containing precipitate at 105℃ until the moisture content was 0.7%, then crushed it to a particle size ≤1mm. The crushed scandium-containing precipitate was mixed with sodium carbonate and calcium oxide solids at a mass ratio of 1:2:1 and calcined in air. The temperature was increased to 1000℃ at a rate of 30℃ / min and maintained at that temperature for 5 hours. The calcined mixture was then washed with water. During countercurrent washing, the liquid-to-solid ratio was water volume:calcined mixture = 1:1. The washing process involved two stages. After filtration, a chromium-containing solution and filter residue were obtained. The chromium leaching rate was 99.99%, and the scandium leaching rate was <0.01%. The filter residue was then countercurrently washed with a dilute acid solution with a pH adjusted to 1.5 using hydrochloric acid. The liquid-to-solid ratio during washing was 3:1 (volume of dilute acid solution in ml: mass of filter residue in g). The washing process involved three stages. After filtration, an aluminum-containing solution and scandium-enriched filter residue were obtained. The aluminum leaching rate was 99.37%, and the scandium leaching rate was <0.01%.

[0037] S5 dissolves the scandium enrichment filter residue in a 6 mol / L hydrochloric acid solution. The liquid-to-solid ratio is 2.5:1 (hydrochloric acid solution volume in ml : scandium precipitate mass in g (50% water content)). The ferric ions in the obtained solution are fully reduced using reduced iron powder at a concentration of 0.2 g / L solution. After complete reduction, the solution and residue are obtained by filtration. The residue is returned to the reduction process. The pH of the filtrate is adjusted to 4.0 using a 10% NaOH solution. The residue and scandium-rich material are then obtained by filtration. The acid leaching-reduction-scandium precipitation process is repeated four times based on the scandium content of the scandium-rich material. The process is then stopped, and the final scandium-rich material is obtained. The scandium recovery rate in this process is 99.52%.

[0038] S6 uses an oxalic acid solution with a mass concentration of 0.8 mol / L to countercurrently wash the above-mentioned ultimate scandium-enriched material. The liquid-to-solid ratio during washing is 2:1 (volume of oxalic acid solution in ml) : weight of ultimate scandium-enriched material (dry basis) in g). The countercurrent washing stage is 5. After washing and filtration, the material is calcined. During calcination, the temperature is increased to 105°C at a rate of 10°C / min and kept at this temperature until the moisture content is 1.5%. The temperature is then increased to 800°C at a rate of 30°C / min and kept at this temperature for 6 hours to obtain scandium oxide with a purity ≥99.95%.

[0039] Matters not covered in this invention are common knowledge.

[0040] For anyone skilled in the art, many possible variations and modifications can be made to the technical solutions of this invention, or equivalent embodiments can be modified based on the disclosed technical content, without departing from the scope of the technical solutions of this invention. Therefore, any simple modifications, equivalent changes, and modifications made to the above embodiments based on the technical essence of this invention without departing from the content of the technical solutions of this invention should still fall within the protection scope of the technical solutions of this invention.

Claims

1. A method for co-extracting scandium from scandium-containing waste acid and waste residue, characterized in that, The method includes the following steps: S1 first undergoes a first-stage leaching process, using sulfuric acid solution to leach iron, manganese, magnesium, aluminum, titanium, and chromium elements from the scandium-containing waste residue. The leaching time is controlled to be ≤30 min. After filtration, a leachate and scandium-enriched residue are obtained. Then, a second-stage leaching process is performed, using sulfuric acid from titanium dioxide production to leach the scandium-enriched residue. After filtration, a scandium-containing leachate and filter residue are obtained. The filter residue is returned to the second-stage leaching process for further leaching. S2 uses a reducing agent to fully reduce the target cation in the scandium-containing leachate, filters to obtain the reduced scandium-containing solution and residue, the residue is returned to the reduction process, and then a phosphate solution is added to the scandium-containing solution to remove zirconium, filters to obtain zirconium phosphate precipitate and the scandium-containing solution after zirconium removal; S3 uses a precipitant to precipitate scandium ions in the scandium-containing solution after zirconium removal, thus obtaining a scandium-containing precipitate; S4 mixes the above scandium-containing precipitate with sodium carbonate and calcium oxide, and calcines the mixture at high temperature. The calcined product is washed with water, filtered to obtain chromium-containing wash water and filter residue. The filter residue is then washed with dilute acid, filtered to obtain aluminum-containing solution and scandium-enriched filter residue. S5 uses an acid solution to leach the scandium in the scandium enrichment filter residue. After filtration, a scandium-containing filtrate and leaching residue are obtained. The leaching residue is returned to the acid solution scandium leaching process. Then, a reducing agent is used to fully reduce the target cation in the scandium-containing leaching solution. A precipitating agent is used to precipitate the scandium ions in the reduced leaching solution. After filtration, a filtrate and primary scandium-rich material are obtained. The above acid leaching-reduction-scandium precipitation process is repeated according to the scandium content of the scandium-rich material. The number of repetitions is ≥1 to obtain the final scandium-rich material. S6 uses oxalic acid to countercurrently wash the above-mentioned ultimate scandium-rich material, and after filtration, oxalic acid wash water and scandium oxalate are obtained. The scandium oxalate is calcined at high temperature to finally obtain scandium oxide with a purity of ≥99.9%. In S2, the target cation to be reduced is ferric iron; In S3, the final pH of precipitation is controlled at 3-6.

2. The method according to claim 1, characterized in that, In S1, the concentration of the sulfuric acid solution used in the first leaching stage is ≥0.5mol / L; the scandium content in the scandium-containing waste residue is ≥50g / t; the leaching temperature is controlled at 20~50℃; and the liquid-solid ratio during leaching is sulfuric acid solution volume: scandium-containing waste residue mass = 1~10:

1. In the second leaching process, the scandium content of the sulfuric acid waste acid from the titanium dioxide process is ≥5mg / L; the leaching temperature is ≥20℃; the leaching time is ≥1h; and the liquid-solid ratio during leaching is sulfuric acid waste acid volume: scandium-enriched residue mass = 1~5:

1.

3. The method according to claim 2, characterized in that, The concentration of the sulfuric acid solution used for leaching in the first stage is 1~3.5 mol / L, and the leaching time is 20~30 min; The two-stage leaching process involves leaching at a temperature of 50-98℃ for 3-6 hours.

4. The method according to claim 1, characterized in that, In S2, the reducing agent is one of the following: iron, magnesium, zinc or aluminum metal powder, sodium metabisulfite, or sodium sulfate. The amount of reducing agent used is 1~20 g / L of scandium-containing solution, and the reduction temperature is ≥40℃. The soluble phosphate used for zirconium removal is sodium phosphate, sodium monohydrogen phosphate, sodium dihydrogen phosphate, potassium phosphate, potassium monohydrogen phosphate, or potassium dihydrogen phosphate. The amount of soluble phosphate added per liter of scandium-containing leachate is 0.06~2 g, and it is prepared as a solution before addition.

5. The method according to claim 4, characterized in that, The reduction temperature is 60~95℃.

6. The method according to claim 1, characterized in that, 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.

7. The method according to claim 1, characterized in that, In S4, the mass ratio of the scandium-containing precipitate, sodium carbonate, and calcium oxide is 0.5~2:0.5~5:0.25~3; the scandium-containing precipitate is dried to a moisture content ≤3%, then crushed to a particle size ≤2cm, and calcined at high temperature at a rate of 10~40℃ / min to a temperature ≥800℃, and held at this temperature for ≥3h; the liquid-to-solid ratio during water washing is water volume: calcined mixture = 1~10:1; the dilute acid used to wash the filter residue is an aqueous solution of sulfuric acid or hydrochloric acid with a pH of 0.5~2, and the liquid-to-solid ratio during washing is dilute acid volume: filter residue mass = 1~30:

1.

8. The method according to claim 7, characterized in that, The roasting temperature is 900~1300℃ and the roasting time is 5~8h.

9. The method according to claim 1, characterized in that, In step S5, the acid solution is one of sulfuric acid, hydrochloric acid, and nitric acid, with a hydrogen ion concentration ≥1 mol / L. During leaching, the liquid-to-solid ratio is acid solution volume (ml): scandium enrichment filter residue dry basis mass (g) = 1~10:

1. The target cation to be reduced is ferric iron, and the reducing agent is one of iron, magnesium, zinc, or aluminum metal powder, sodium metabisulfite, or sodium sulfate. The amount of reducing agent used is 0.1~3 g / L of leaching solution, and the temperature of the leaching solution system during reduction is ≥40℃. The precipitant is one or more of sodium hydroxide, potassium hydroxide, carbonate, bicarbonate, calcium oxide, calcium hydroxide, magnesium oxide, magnesium hydroxide, and ammonia water. The pH at the precipitation endpoint is controlled at 3~6. When the scandium content of the scandium-containing precipitate is ≥5%, the repeated acid leaching-reduction-scandium precipitation process is stopped to obtain a final scandium-rich material with a higher scandium content.

10. The method according to claim 9, characterized in that, The hydrogen ion concentration is 2~7 mol / L; The temperature of the leachate system during reduction is 60~95℃.

11. The method according to claim 1, characterized in that, In S6, the oxalic acid concentration used for countercurrent washing is ≥0.1mol / LL, the liquid-to-solid ratio during washing is oxalic acid solution volume: final scandium-rich material dry weight = 1~10:1, the number of countercurrent washing stages is ≥4, after washing and filtration, the precipitate is calcined, the temperature is raised to 105~125℃ at a rate of 5~10℃ / min to dry the precipitate until its moisture content is ≤3%, and the temperature is further raised to ≥700℃ at a rate of 10~40℃ / min, and held at this temperature for ≥3h.

12. The method according to claim 11, characterized in that, The concentration of oxalic acid used for countercurrent washing is 0.3~1mol / L, and the number of countercurrent washing stages is 5~10. The roasting temperature is 800~1000℃, and the holding time is 5~8h.

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