WO2016108344A1 - Method for collecting scandium from residue generated during refining of low-grade nickel ore - Google Patents

Abstract

The present invention relates to a method for collecting scandium from residue generated during the collection of nickel from low-grade nickel ore, through a series of procedures, including leaching, neutralization, dissolving, solvent extraction, neutralization-precipitation, and firing. The method of the present invention allows an effective collection of scandium from residue discarded at the time of nickel collection, and thus can reutilize the residue generated at the time of nickel collection and obtain scandium at lower costs.

Description

 Specification

 FIELD OF THE INVENTION The present invention relates to a rare water method of scandium from residues generated during smelting low quality nickel ores, and more particularly, in the process of recovering nickel from low quality nickel ores. The present invention relates to a method for recovering scandium through a series of processes including leaching, neutralization, dissolution, solvent extraction, neutralization precipitation, and calcination from the resulting residue. Background of the Invention Scandium (Sc) is a silver white transition metal belonging to Group 3 (Group 3B) in the periodic table and is one of rare earth elements.

Scandium is a lightweight metal material that is as light as aluminum (density, 2.7 g / cm ³ ; melting point 66CTC), is less corrosive, and has a higher melting point. Scandium is used in sports equipment such as baseball bats, racing bicycle skeletons and parts, golf clubs, guns, and cell phone cases because it adds to aluminum and allows crystals to be finely and uniformly mixed, greatly improving strength, elasticity, and weldability. In addition, zirconium oxide (Ζ) 2 to which scandium is added is used as a highly efficient electrolyte in a solid oxide fuel cell. Furthermore, scandium is also used for very bright (high-brightness) metal halide lamps, and scandium oxide (Sc ₂ 0 ₃ ) is used for surface coating of laser crystals and high power ultraviolet lasers.

Scandium is present in small amounts in most rare earth ores and in uranium and tungsten ores. The scandium-rich ores contain tortveitite ((Sc, Y) ₂ Si ₂ 0 ₇ , about 30-40% Sc. ₂ 0 ₃ included) and callback kites (kolbeckite: ScP0 to ₄ .2H ₂ 0) which are usually buried in a specific area of Norway. Scandium itself is obtained as a by-product of uranium and tungsten production, rather than directly from the ore, and its annual output is only about 2 tonnes, calculated as the output of Sc ₂ 0 ₃ .

 As such, despite various applications of scandium, there have been efforts to improve scandium production or to recover scandium from various materials due to its insufficient production. As an example of a method for recovering nickel from a low-grade nickel ore containing nickel in the prior art, Korean Patent No. 10-1281367 discloses a reduction, leaching / pH control and precipitation from low-grade nickel ore, for example limonite. A method of recovering nickel is disclosed. This document discloses a method of removing Si by adjusting the pH before precipitation to 2.5 to 5.5 in order to prevent silicon (Si) in limonite from being dissolved and precipitated together in the leaching step to lower the deposition rate of nickel. . In addition, Korean Patent Publication No. 2007-107761 discloses a method for recovering nickel and cobalt from laterite using pressure leaching. Furthermore, Korean Patent Publication No. 2006-52817 discloses a method for recovering nickel and cobalt by dummy leaching of low quality nickel or cobalt containing material.

It is known that a large amount of scandium is contained in the residue generated during the recovery of nickel from the above-described low-grade nickel ore, but no attempt or recovery method for recovering scandal from the residue is known at all. SUMMARY OF THE INVENTION Accordingly, it is an object of the present invention to provide a series of processes for efficiently recovering scandium from residues generated in the process of recovering nickel from low quality nickel ores. In order to achieve the above object, the present invention comprises the steps of (1) obtaining a scandium leaching liquid by adjusting the pH of the residue generated in the process of recovering nickel from low-grade nickel ore in the range of -1 to 2; (2) neutralizing the scandium leachate to a range of pH 4 to 7 to obtain a scandium concentrated cake; (3) dissolving the scandium concentrated cake in acid; (4) selectively extracting scandium by adding a cation extractant to the dissolution solution; (5) neutralizing the extract after the precipitation in the range of pH 6 to 8 to precipitate the scandium; And (6) calcining the precipitate to obtain Sc ₂ 0 ₃ , thereby providing a method for recovering scandium from residues generated in the process of recovering nickel from low quality nickel ores.

According to the method for recovering disintegrated scandium according to the present invention, since the scandium can be concentrated primarily from the low concentration scandium in the residue that is discarded, it is possible to enjoy the scale of equipment in the step of acid dissolution, solvent extraction and neutralization precipitation, which is a post-process. Operating costs can be lowered, resulting in economic and fair recovery rates. In addition, according to the present invention, since the residue generated during nickel recovery is recycled, scandium can be obtained at a lower cost since the raw material cost is not required. Figure 1 shows a flow chart of the scan number recovery method according to the present invention. The flow chart shows (1) Sc leaching, (2) neutralization, (3) acid dissolution, and (4) solvent extraction from residues from the leaching / pH step during the recovery of Ni concentrate from low-grade nickel ore (limonite). , (5) neutral precipitation and (6) calcining to obtain Sc ₂ 0 ₃ .

2 shows an Eh ᅳ pH diagram of scandium. Detailed description of the invention Hereinafter, the present invention will be described in more detail. The present invention comprises the steps of: (1) adjusting the pH of the residue generated in the process of recovering nickel from a low-grade nickel ore to the range of -1 to 2 to obtain a scandium leaching liquor; (2) neutralizing the scandium leachate to a range of pH 4-7 to obtain a scandium concentrated cake; (3) dissolving the scandium concentrated cake in acid; (4) selectively extracting scandium by adding a cation extractant to the dissolution solution; (5) neutralizing the extract after the precipitation in the range of pH 6 to 8 to precipitate the scandium; And (6) calcining the precipitate to obtain Sc ₂ 0 ₃ , thereby providing a method for recovering scandium from residues generated in the process of recovering nickel from low quality nickel ores.

 According to the method of the present invention, scandium can be concentrated by leaching and neutralizing the residue to obtain a scandium concentrated cake. As a result, it is possible to reduce the size of facilities in the post-dissolution, dissolution, solvent extraction, and neutralization stages, resulting in lower investment and operating costs, and higher recovery rate. In addition, since scandium is recovered from the discarded residue, scandium can be recovered at low cost. Each process of the scandium recovery method according to the present invention is as shown in FIG. Scandium recovery method according to the present invention, (1) Sc leaching, (2) neutralization, (3) acid dissolution,

It includes a series of processes including (4) solvent extraction, (5) neutralization precipitation, and (6) firing. Each process is explained in full detail.

0) Sc leaching

In the Sc leaching step, the pH of the residue generated in the process of recovering nickel from the low quality nickel ore is adjusted to a range of -1 to 2 to obtain a scandium leaching liquor. As used herein, the term 'low grade nickel ore' refers to nickel oxide (Laterite), which can be largely divided into limonite and saprolite according to the nickel content and properties. The low quality nickel ore may preferably be limonite, but is not limited thereto. _,

 As used herein, the term 'residue' refers to process waste that occurs during the recovery of nickel from the low quality nickel ores. In one embodiment of the present invention, the residue may be produced, for example, when recovering nickel through a series of processes including a step of reducing, leaching / pH adjustment and precipitation from low-grade nickel ore. In another embodiment of the present invention, the residue is produced in the extraction / pH adjustment step, for example, when recovering nickel through a series of processes including, for example, reduction, leaching / pH control and precipitation from low quality nickel ores. It may be a discarded residue. See, for example, Korean Patent No. 10-1281367 for a detailed description of the process of recovering nickel from the low quality nickel ore. The residue is 0.001 to 0.05% Sc, 5 to 40% Fe, 3 to 15% A1, 5 to 20% Si, 0.1 to 3% Ca, 0.001 to 0.1% Co and 0.001 5% Ni It may contain.

 The Sc leaching step may be performed by dissolving the residue into a slurry (la), adjusting the pH of the slurry (lb), leaching Sc (lc), and obtaining a liquid by solid-liquid separation (Id). It can include a series of substeps, including).

 In step (la) above, the residue is dissolved in a suitable solvent, for example water, or a slightly acidic solution, preferably water, in order to make the residue into a slurry. The solvent may be used at a ratio of 1 L of solvent per 0.5 to 3.0 kg of the residue, and the dissolution may be performed at room temperature for 10 to 60 minutes.

In step (lb), the pH of the slurry can be adjusted to -1 to 2, preferably 0 to 1, more preferably 0.3 to 0.8, and most preferably 0.5. The pH adjustment can be carried out using an inorganic acid selected from the group consisting of hydrochloric acid, sulfuric acid and nitric acid, and preferably can be adjusted using hydrochloric acid. As shown in FIG. 2, since the scan rhythm is present in an ionized state (Sc ^{+ 3} ) in the range of pH 0 to 4, the scandium is concentrated in the leachate by adjusting the pH of the sultry to leach the scandium in the residue in an ion state. You can.

In the step (lc), the slurry of the sulfuric state is stirred by stirring the pH-adjusted slurry ^{at a} temperature of 30 to 70 ^° C., preferably at 50 ° C. for 2 to 6 hours, preferably 4 hours. Leak in liquid state.

 In the above step (Id), a liquid (also referred to as 'leaching liquor') is obtained by solid-liquid separation, in which the obtained liquid contains a large amount of leached scandium. The remaining solids except the liquid are discarded. The solid-liquid separation method may include filter press centrifugation, centrifugal filtration, or reduced pressure filtration, but is not limited thereto.

 The Sc leaching rate in the aforementioned Sc leaching step is 80% 'or more, as calculated by the following equation.

 Leaching rate (%) = (scandium content in the leaching liquor / scandium content in the residue) * 100

(2) China

 In this neutralization step, the scandium leachate obtained in step (1) is neutralized to a range of pH 4 to 7 to prepare a concentrated concentrate cake.

As shown in FIG. 2, scandium is present in a white precipitate state (Sc (OH) ₃ ) in the range of pH 4 to 8, so that scandium may be precipitated by neutralizing the scandium leachate.

 The neutralization step may include a series of substeps including adjusting the pH of the scandium leachate (2a) and obtaining a solid by solid-liquid separation (2b).

 In the step (2a), the pH of the scandium leachate is 4 to 7, preferably

5 can be adjusted. The pH adjustment is performed from the group consisting of calcium hydroxide, sodium hydroxide, magnesium hydroxide, iron hydroxide, sodium carbonate, calcium carbonate, calcium oxide and the like. It can be carried out using the base selected. The base used may be used in a dilution amount of 10 to 30% by weight, preferably 20% by weight, but is not limited thereto. The conditions of the neutralization process may be performed for 1 to 3 hours, preferably 2 hours at phase silver.

 In the step (2b), a solid is obtained by solid-liquid separation. The solid-liquid separation method is as described in step (Id). The solid obtained is taken to the next stage and the liquid can be reused in the process according to the invention.

 The Sc neutralization rate (precipitation) in the aforementioned neutralization step is 100% as calculated by the following equation.

 Neutralization rate (%) = 100- (Scandium amount in neutralized precipitated liquid / Scandium amount in leached liquid) * 100

(3) acid dissolution

 In this acid dissolving step, the scandium concentrate cake listed in step (2) is dissolved in acid.

 The acid dissolving step includes dissolving the scandium concentrated cake to make a sultry (3a), adding acid to the slurry for further dissolution (3b) and obtaining liquid by solid-liquid separation (3c) It can include a series of substeps.

 In step (3a), a suitable solvent, for example water, or a slightly acidic solution, preferably water, may be used to dissolve the scandium concentrated cake. The scandium concentrated cake and the solvent may be used in a ratio of 1.0 to 3.0 kg of the scandium concentrated cake per 1 L of solvent, and the dissolution conditions may be performed at room temperature for 10 to 60 minutes.

The acid used in step (3b) may be an inorganic acid, preferably hydrochloric acid, selected from the group consisting of hydrochloric acid, sulfuric acid, and nitric acid. The acid can be used in an amount to adjust the pH of the sultry to 0-1. The dissolution condition is It may be carried out at room temperature for 10 minutes to 3 hours, preferably 1 hour. In the step (3c), the liquid is obtained by solid-liquid separation. The solid-liquid separation method is as described in step (Id). The obtained liquid (also called 'future solution') proceeds to the next step and the solid is discarded.

 The Sc dissolution rate in the hydrochloric acid dissolution step described above is calculated by the following equation.

95%.

 Dissolution rate (%) = (Scandium amount in acid dissolving liquid / Scandium amount in scandium concentrate cake cake) * 100 (4) Solvent extraction

 In this solvent extraction step, scandium is selectively extracted by adding a cation extractant to the dissolution solution obtained in step (3).

 The solvent extraction step may comprise a series of substeps including an extraction step 4a, a washing step 4b and a back extraction step 4c.

 In step (4a), scandium is extracted using a suitable solvent. The solvent is di-2-ethylnuclear phosphonic acid (D2EHPA), 2-ethylnuclear phosphonic acid mono-2-ethylnucleyl ester (PC88A),

It may be a cationic extractor (or acidic extractant) such as bis (2,4,4-trimethylpentyl) phosphinic acid (CYANEX272), which is used alone or in combination with tri-butyl phosphate (TBP). It can be used in the form of water. In the dissolving solution obtained in step (3), scandium is present in a cationic form (Sc ³⁺ ), and the cation extractant serves to selectively extract only scandium as a cation in the dissolving solution, as shown in Equation 1 below, and TBP is It serves to facilitate solvent separation.

 <Equation 1>

Sc ³⁺ + 3 [H ₂ R ₂ ] → ScR ₃ * 3HR + 3H ⁺

The solvent may preferably comprise 5 to 10 volume% cationic extractant and 1 to 5 volume% TBP based on the total amount of the total solvent, more Preferably it may comprise 5 volume% D2EHPA and 2.5 volume% TBP based on the total amount of the total solvent. The solvent may be used in an amount of 10 to 30% by volume, preferably 20% by volume, based on the volume of the solution after dissolution.

 Step (4a) may be performed once, but may be performed two or more times according to the extracted scandium concentration.

 Step (4a) may further include adjusting the pH to 0 to 1.5 by adding hydrochloric acid prior to extraction, in order to adjust the pH of the dissolution solution obtained in step (3) to an appropriate pH for extracting scandium. Can be.

 In step (4b), a small amount of impurities (e.g., Fe, Al, Co, Ni, Si, etc.) extracted together with the scandium in step (4a) are removed using a suitable detergent. Examples of the cleaning agent include oxalic acid or hydrochloric acid solution, and the cleaning agent may be used in an amount for adjusting the pH of the mixture of the extraction solvent and the cleaning solution in the range of 0 to 1 in the washing step.

 The step (4b) may be performed one or more times, preferably four times. In step (4b) the volume ratio of the extract after the detergent to the detergent may be 1: 2.

 In the step (4c), using a suitable back extraction solution to obtain a solution containing scandium. Examples of the reverse extraction solution include, but are not limited to, sodium hydroxide, ammonia water, sodium carbonate, and the like. The step (4c) may be performed one or more times, preferably two times.

 Sc extraction rate in the solvent extraction step described above is calculated by the following formula

100% and Sc purity is 3N or more as calculated by the following equation.

 Extraction rate (%) = [100-(quantity of scandium in extract after extract / amount of scandium in acid solution)] * 100

 Purity = 100-amount of impurities other than scandium (based on back extraction)

(5) Chinese precipitation

In this neutralization precipitation step, the back extract obtained in step (5) is pH 6 to Neutralize to 8 to precipitate scandium.

 The thickening precipitation step may comprise a series of substeps including a pH adjusting step (5a), a solid-liquid separation step (5b), and a washing step (5c).

The pH is adjusted to 6-8 by addition of a suitable acid in step (5a), wherein Sc (OH) ₃ is produced as a white precipitate. The neutralization can be carried out using an inorganic acid selected from the group consisting of hydrochloric acid, sulfuric acid and nitric acid. The neutralization process may be performed at room temperature for 0.5 to 3 hours.

 Then, in step (5b), a solid is obtained by solid-liquid separation. The solid-liquid separation method is as described in step (Id). The obtained solid is passed to the next step, the liquid is discarded and part is reused in the process.

 Thereafter, in step (5c), the obtained solid is washed with water. At this time, alkali elements, such as Na, are removed. The washed solid can be dried.

 The Sc precipitation rate in the aforementioned neutral precipitation step is 99% or more as calculated by the following formula.

 Sedimentation rate (%) = [100-(amount of scandium in the thickening precipitate thickening / the amount of scandium in the back extract thickening)] * 100

(6) Firing In the firing step, the precipitate (Sc (OH) ₃ ; solid) is calcined to obtain Sc ₂ 0 ₃ . The firing can be carried out under air or 0 ₂ atmosphere at a temperature of 400 to 800 ° C., preferably 600 ° C., for 2 to 6 hours, preferably 4 hours. Through the calcination step Sc ₂ 0 ₃ can be obtained. As described above, according to the scandium recovery method according to the present invention, scandium can be concentrated first, and thus the scale of equipment in the acid dissolution, solvent extraction, and neutralization and precipitation steps, which are post-processes, can be reduced, and the recovery rate can be increased. In addition, according to the method of the present invention, Since scandium is recovered from the residue that is discarded, scandium can be recovered at low cost. Hereinafter, the present invention will be described in detail with reference to Examples, but the following Examples are only intended to illustrate the present invention, which is not intended to limit the scope of the present invention. Example 1 Recovery of Scandium from Residue <!-!> Sc Leaching

 1.5 kg of the residue was added to 1 L distilled water and stirred at room temperature for 30 minutes to obtain a slurry. HC1 was added to the slurry to adjust the pH to 0.5, and the reaction temperature was adjusted to 5 CTC, followed by stirring for 4 hours, thereby leaching Sc. Then, the leachate was subjected to solid-liquid separation to discard the solid and to obtain a leachate.

<1-2> Chinese

20% Ca (OH) ₂ in the slurry state was added to the leachate after elution, and the pH was adjusted to 5.0 by stirring. Then, the mixture was neutralized by further stirring for 2 hours at silver phase, followed by solid-liquid separation, and the liquid was reused into the process to obtain a solid Sc concentrated cake.

<1-3> acid dissolution

2 kg of the Sc concentrated cake was added to 1 L of distilled water, and stirred at room temperature for 30 minutes to obtain a slurry. HC1 was added to the slurry to adjust the pH to 0.5, and when the pH was stabilized, the solution was further stirred for 1 hour at silver to dissolve. After the reaction was completed, the liquid was separated and the residue was discarded to obtain a liquid. <1-4> Solvent Extraction

 HC1 was added to the hydrochloric acid solution to adjust the pH to 0 to 0.5. After the solution was added to the extraction stage, 5 vol% of di-2-ethylnucleic acid phosphoric acid (D2EHPA) and 2.5 vol% of tri-butyl phosphate (TBP) were added as a solvent to extract Sc toward the solvent. When the scandium was extracted into the solvent, the volume ratio of the solvent and the hydrochloric acid solution was adjusted to 1: 5 to concentrate the scandium in the solvent, and the extraction stage was extracted in one stage. The extracted solvent contains impurities such as Fe, Al, Si, Ca, Co, and Ni in addition to scandium. Therefore, after the solvent was transferred to the washing stage, the pH was adjusted to 0 to 0.2 by adding oxalic acid solution to remove impurities. At this time, the volume ratio of the solvent and the solution was adjusted to 1: 2, and impurities were removed through four stages. The solvent from which the impurities were removed was transferred to a back extraction stage, and then back extracted by adding sodium hydroxide solution. The volume ratio of the solvent and the sodium hydroxide solution from which the impurities were removed was adjusted to 1: 1, and the scandium in the solvent was back extracted through two stages. <1-5> neutralization precipitation

After adding HC1 to the Sc-containing solution from the back extraction stage, the pH was adjusted to 6-8. Thereafter, the mixture was further stirred for 1 hour at room temperature, followed by solid-liquid separation to obtain Sc) H) ₃ as a solid. The solid was washed with water to remove alkali elements such as Na and dried.

<1-6> firing

The dried Sc (OH) ₃ was calcined at 60 CTC for 4 hours under an air atmosphere to obtain Sc ₂ 0 ₃ . Experimental Example 2: Component Content by Processes The contents of Sc, Fe, Al, Si, Ca, Co, and Ni were analyzed for each process. Table 1 shows the results of the residues from low-grade nickel ores and the solutions, concentrated cakes, or products from each of the processes of leaching, neutralization, acid dissolution, solvent extraction, neutralization and firing.

 The results are shown in Table 1 below.

 Table 1

As shown in Table 1 above, as the content of other impurities is lowered through each step, the scandium content is increased, and finally SC20 ₃ can be obtained in very high yield and purity.

 Specifically, the residues generated after recovering nickel from low-grade nickel ores are Sc 123ppm, Fe 20.3%, Al 7.4%, and Si 10.7%, which contains impurities such as low concentrations of Sc and high concentrations of Fe, Al, and Si. have. When leaching this, the scandium in the leaching liquor is 68 mg / L, and when it is substituted into the following formula (1), it can be seen that the Sc leaching rate is 82%.

(1) leaching rate (%) = [scandium amount I in the leaching liquor amount of scandium in the residue] * 100 and then, to neutralize the scandium in the leached scandium solution in the form of a cake. In the neutralization process through the following formula (2), Sc is 100%, and Fe 60%, Al 100%, Co 50%, and Ni 55% are precipitated and caked as other impurities. At this time, it can be seen that the scandium content, which was 123 ppm (0.0123%) in the initial residue, was concentrated about 10 times to 1000 ppm (0.1%) in the Sc concentrate cake through a neutralization process. (2) Neutralization rate (%) = 100- (Scandium amount in neutralizing solution / Scandium amount in leaching solution) * 100

 The process enables economical recovery of scandium by reducing the scale and increasing the recovery rate of the post-acid dissolution, solvent extraction and neutralization precipitation steps. Thereafter, the concentrated scandium cake was dissolved in solution by adjusting the pH of water and HC1 in an acid dissolution process. At this time, the scandium concentration was 380 ppm and other impurities were Fe 8651 lppm, Al 23243 ppm, and Si 1890 ppm. It was possible to dissolve the scandium in the concentrated scandium cake through.

 Subsequently, the Sc dissolution solution was reacted with a solvent to extract Sc into the solvent. Scandium content in the extract after-treatment is <lppm, it can be seen that the scandium extraction rate is 100% by substituting this in the following formula (3).

 (3) Extraction rate (%) = [100- (scandium concentration in extraction solution / scandium concentration in solution after dissolution)] * 100

 When scandium is extracted, small amounts of other impurities are also extracted.

By calculating the A1 extraction rate in the same equation, it can be seen that Fe 2%, Al 2.6%, Si 2.9%, and Ca 0.98% are extracted together. The impurities extracted together in the extracted solvent were removed through a washing process, back extracted with sodium hydroxide solution, and neutralized and precipitated in the form of Sc (OH) ₃ to recover scandium as a product. After neutralizing by adding HC1 solution to the sodium hydroxide solution, it was calculated through the following equation (4), the scandium precipitation was found to be 99.5%.

 (4) Sedimentation rate (%) = [10 (scandium amount in sedimentation liquid / scandium amount in back extraction liquid)] * 100

Sc (OH) ₃ obtained through neutralization precipitation was commercialized as Sc ₂ 0 ₃ through a calcination process, and as a result, Sc ₂ 0 ₃ having a purity of 3N or higher was obtained.

Claims

Patent Claims: Claim 1. (1) obtaining a post-scandium leaching solution by adjusting the pH of the residue generated in the process of recovering nickel from low-grade nickel ore to a range of -1 to 2; (2) neutralizing the scandium leached solution to a pH range of 4 to 7 to obtain a scandium-enriched cake; (3) dissolving the scandium enriched cake in acid; (4) selectively extracting scandium by adding a cationic extractant to the post-dissolution solution; (5) neutralizing the extracted liquid to a pH range of 6 to 8 to precipitate scandium; (6) Calcining the precipitate to obtain Sc ₂ 0 ₃ A method of recovering scandium from residue generated in the process of recovering nickel from low-grade nickel ore, including a method. Claim 2. The method according to claim 1, wherein the low-grade nickel ore is limonite. Claim 3. The method according to claim 1, wherein the pH adjustment in step (1) is performed using an inorganic acid selected from the group consisting of hydrochloric acid, sulfuric acid, and nitric acid. 4. The method of claim 1, wherein the neutralization in step (2) is performed using a base selected from the group consisting of calcium hydroxide, sodium hydroxide, magnesium hydroxide, iron hydroxide, sodium carbonate, calcium carbonate, and calcium oxide. . 5. The method of claim 1, wherein the acid in step (3) is hydrochloric acid, sulfuric acid, and nitric acid. A method characterized in that it is an inorganic acid selected from the group consisting of Claim 6. The method of claim 1, wherein the cationic extractant in step (4) is di-2-ethylhexyl phosphonic acid, 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester, and bis (2,4,4- A method characterized in that it is any one selected from the group consisting of trimethylpentyl)phosphinic acid or a mixture of the selected cationic extractant and tri-butyl phosphate (TBP). 7. The method of claim 1, wherein step (4) further comprises adjusting the pH to 0 to 1.5 before extraction. 8. The method according to claim 1, wherein the neutralization in step (5) is carried out using an inorganic acid selected from the group consisting of hydrochloric acid, sulfuric acid and nitric acid. Claim 9. The method according to claim 1, wherein the calcination in step (6) is performed in a temperature range of 400 to 800 ^° C in air or 02 atmosphere.