Process for recovering molybdenum from high-acid high-ion-concentration white smoke leachate

Abstract

The invention discloses a process for recovering molybdenum from high-acid high-ion-concentration white smoke leachate, and belongs to the technical field of metal smelting, the recovery efficiency of molybdenum is improved by adding a self-synthesized modified flocculant in the process, the modified flocculant is of a non-ionic main net structure, acrylamide does not have strong charge groups after being polymerized, and the high-acid high-ion-concentration white smoke leachate can be used for recycling molybdenum from the high-acid high-ion-concentration white smoke leachate. Sulfonic acid groups of sulfonated starch are only hydrophilic and do not adsorb molybdate radicals, fine solid impurities are agglomerated only through bridging action, electrostatic adsorption or chemical reaction with negatively charged molybdate radicals is avoided, and molybdenum loss is avoided from the source; iron chloride in the raw materials provides iron ions, the iron ions react with 2, 5-diamino terephthalic acid to generate an iron-based metal organic framework, the iron-based metal organic framework has a porous structure and can enhance subsequent impurity adsorption, PVDF is a typical acid-resistant and corrosion-resistant high polymer material, and after PVDF is compounded with the iron-based metal organic framework, the adsorption capacity of the PVDF is greatly improved. The iron-based metal organic framework can be prevented from being dissolved in strong acid, and meanwhile, the powder is endowed with excellent mechanical stability.

Description

Technical Field

[0001] This invention belongs to the field of metal smelting technology, specifically a process for recovering molybdenum from high-acid, high-ion-concentration white dust leachate. Background Technology

[0002] Molybdenum is a very important rare metal, widely used in metallurgy, chemical industry, military, electronics, biomedicine, agriculture, environmental protection, and aerospace due to its excellent physical and chemical properties such as high strength, high melting point, abrasion resistance, and corrosion resistance. It has become an important raw material and an irreplaceable strategic resource in the national economy. Molybdenite is one of the important ore sources for molybdenum production, commonly found in the oxidation zone of lead-zinc deposits. It can coexist with other secondary lead-zinc minerals such as cerussite, arsenite, phosphogypsum, vanadium pyrite, smithsonite, hemimorphite, as well as limonite, pyrolusite, and permeate.

[0003] Traditionally, molybdenum-lead ore leaching has employed alkaline processes, specifically sodium hydroxide decomposition and sodium sulfide decomposition. Sodium hydroxide decomposition uses the strong inorganic base sodium hydroxide as the leaching agent. During leaching, molybdenum and lead are leached into soluble sodium molybdate and sodium lead oxide, respectively, which simultaneously enter the solution. However, grade zero molybdenum products have extremely strict requirements regarding lead content, necessitating complex deep lead removal processes. To simplify the molybdenum-lead separation process in sodium hydroxide decomposition, sodium sulfide leaching was developed. This method leaches molybdenum while lead precipitates as lead sulfide, remaining in the residue. However, sodium sulfide leaching typically operates at temperatures above 90°C, resulting in high energy consumption. Furthermore, the amount of sodium sulfide used is strictly controlled; insufficient sodium sulfide leads to poor molybdenum-lead separation, while excessive sodium sulfide increases solution viscosity, complicates solid-liquid separation, and causes discoloration due to the formation of thiomolybdates, leading to a decline in product quality. Furthermore, sodium sulfide is easily hydrolyzed at high temperatures, producing hydrogen sulfide gas, and is also easily oxidized to elemental sulfur. In the subsequent acidification and precipitation of ammonium molybdate to adjust the pH, a large amount of hydrogen sulfide will escape, posing safety and environmental risks. Therefore, there is an urgent need to develop a smelting process that is both environmentally friendly and facilitates deep separation of lead and molybdenum. Summary of the Invention

[0004] The purpose of this invention is to provide a process for recovering molybdenum from high-acid, high-ion-concentration white dust leachate. By adding a self-synthesized modified flocculant to the process, the molybdenum recovery efficiency is improved. The modified flocculant of this invention has a non-ionic, predominantly network structure. After acrylamide polymerization, it has no strongly charged groups. The sulfonate groups of sulfonated starch are only hydrophilic and do not adsorb molybdate ions. It only agglomerates fine solid impurities through bridging and will not undergo electrostatic adsorption or chemical reaction with negatively charged molybdate ions, thus avoiding molybdenum loss from the source.

[0005] The objective of this invention can be achieved through the following technical solutions: A process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate includes the following steps: Step 1: Using molybdenum-lead ore as raw material, oxidative roasting is performed to obtain lead molybdate roasted sand with a molybdenum content of 12.32% and a lead content of 42%. The lead molybdate roasted sand and a sulfuric acid solution with a concentration of 100g / L are added to a reaction vessel and stirred for 30-40 minutes at 40-50℃ and 400-500r / min. Then, phosphoric acid is added as a leaching aid and leached for 2-3 hours. Then, a modified flocculant is added and stirring is continued for 40-50 minutes. After standing for 10-12 minutes, a mixed solution is obtained.

[0006] Step 2: Mix 30-32% P507 (volume fraction) and 70-72% sulfonated kerosene (volume fraction) to obtain an extractant. Extract for 1-2 hours under the conditions of an extraction ratio of O / A=2:1, countercurrent extraction, and 5 extraction stages to obtain molybdenum-loaded P507. Back-extract the molybdenum-loaded P507 with 10-12% hydrogen peroxide (mass fraction) under the conditions of an extraction ratio of O / A=5:1, countercurrent extraction, and 3 extraction stages to obtain molybdenum by evaporation and crystallization. This completes a process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate.

[0007] Furthermore, the ratio of lead molybdate calcined sand, sulfuric acid solution, phosphoric acid and modified flocculant is 4-5g: 20-30mL: 0.8-0.9g: 0.2-0.3g.

[0008] Furthermore, the specific preparation steps of the modified flocculant are as follows: Sulfonated starch and deionized water were added to a reaction vessel and stirred for 30-40 minutes at 70-75℃ and 400-500 r / min. Then Ti / PVDF powder was added, heated to 60-70℃, and ultrasonically dispersed for 30-40 minutes. Acrylamide, N,N'-methylenebisacrylamide, and potassium persulfate were then added, and stirring was continued until a gel was formed. The gel was cut into small pieces and washed in distilled water. After washing, the gel was dried at 60-70℃ for 20-22 hours and then pulverized through a 200-220 mesh sieve to obtain the modified flocculant.

[0009] Furthermore, the ratio of sulfonated starch, deionized water, Ti / PVDF powder, acrylamide, N,N'-methylenebisacrylamide and potassium persulfate is 10-12g:100-120mL:3-4g:30-40g:4-5g:0.8-0.9g.

[0010] Furthermore, the specific preparation steps of sulfonated starch are as follows: Add aminosulfonic acid and urea to a reaction vessel and stir for 30-40 minutes at 70-75℃ and 400-500 r / min. Then add potato starch, heat to 90-100℃, and continue the reaction for 5-6 hours. After natural cooling, filter and collect the precipitate. Wash the precipitate 2-4 times with deionized water and anhydrous ethanol, and dry it under vacuum at 60-70℃ for 1-2 hours to obtain sulfonated starch.

[0011] Furthermore, the ratio of aminosulfonic acid, urea, and potato starch is 100-120g: 120-130g: 15-20g.

[0012] Furthermore, the specific preparation steps for Ti / PVDF powder are as follows: A solution of 2,5-diaminoterephthalic acid and N,N-dimethylformamide was added to a polytetrafluoroethylene-lined autoclave and stirred for 30-40 minutes at 20-25°C and 500-600 r / min. Then, ferric chloride and PVDF powder were added, and the mixture was heated to 120-130°C and reacted for 24-26 hours. The mixture was then allowed to cool naturally to room temperature, filtered, and the filter cake was washed 2-4 times with methanol solution and deionized water, respectively. The cake was then vacuum dried at 60-70°C for 1-2 hours to obtain Ti / PVDF powder.

[0013] Furthermore, the ratio of 2,5-diaminoterephthalic acid, N,N-dimethylformamide solution, ferric chloride and PVDF powder is 50-60g: 550-600mL: 35-50g: 30-40g.

[0014] The beneficial effects of this invention are: 1. The present invention provides a process for recovering molybdenum from high-acid, high-ion-concentration white dust leachate, which improves the molybdenum recovery efficiency by adding a self-synthesized modified flocculant to the process.

[0015] 2. In this invention, ferric chloride provides iron ions, which react with 2,5-diaminoterephthalic acid to generate an iron-based metal-organic framework. The iron-based metal-organic framework has a porous structure, which can enhance the subsequent adsorption of impurities. PVDF (polyvinylidene fluoride) is a typical acid-resistant and corrosion-resistant polymer material. After being combined with the iron-based metal-organic framework, it can prevent the iron-based metal-organic framework from dissolving in strong acids. At the same time, it gives the powder excellent mechanical stability, ensuring that the microporous / mesoporous structure of the iron-based metal-organic framework is not destroyed by strong acids when it is combined with sulfonated starch and acrylamide in S3. It can then work synergistically with the network structure of the flocculant to increase the adsorption capacity for fine impurities in the leachate and improve the impurity removal efficiency. Through the acid-resistant framework of Ti / PVDF, the sulfonic acid groups of sulfonated starch, and the carbon chain structure of polyacrylamide, the flocculant can exist stably in 100g / L sulfuric acid without dissolving or decomposing.

[0016] 3. The modified flocculant of the present invention has a non-ionic network structure. After acrylamide polymerization, it has no strongly charged groups. The sulfonate groups of sulfonated starch are only hydrophilic and do not adsorb molybdate. It only aggregates fine solid impurities through bridging and will not undergo electrostatic adsorption or chemical reaction with negatively charged molybdate, thus avoiding molybdenum loss from the source. Detailed Implementation

[0017] The technical solutions of the present invention will be clearly and completely described below with reference to the embodiments of the present invention. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those of ordinary skill in the art without creative effort are within the scope of protection of the present invention.

[0018] Example 1: A process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate, comprising the following steps: S1: Add 50g of 2,5-diaminoterephthalic acid and 550mL of N,N-dimethylformamide solution to a polytetrafluoroethylene-lined autoclave. Stir for 30min at 20℃ and 500r / min. Then add 35g of ferric chloride and 30g of PVDF powder. Heat to 120℃ and continue the reaction for 24h. Cool naturally to room temperature, filter, wash the filter cake twice with methanol solution and deionized water, and dry under vacuum at 60℃ for 1h to obtain Ti / PVDF powder.

[0019] S2: Add 100g of aminosulfonic acid and 120g of urea to the reaction vessel, stir for 30min at 70℃ and 400r / min, then add 15g of potato starch, heat to 90℃, continue to react for 5h, cool naturally, filter, collect the precipitate, wash the precipitate twice with deionized water and anhydrous ethanol, and dry under vacuum at 60℃ for 1h to obtain sulfonated starch.

[0020] S3: Add 10g of sulfonated starch and 100mL of deionized water to a reaction vessel and stir for 30min at 70℃ and 400r / min. Then add 3g of Ti / PVDF powder, heat to 60℃, and ultrasonically disperse for 30min. Then add 30g of acrylamide, 4g of N,N'-methylenebisacrylamide, and 0.8g of potassium persulfate and continue stirring until a gel is formed. Cut the gel into small pieces and wash it in distilled water. After washing the gel, dry it at 60℃ for 20h, pulverize it and pass it through a 200-mesh sieve to obtain the modified flocculant.

[0021] S4: Lead molybdate calcined sand with a molybdenum content of 12.32% and a lead content of 42% was obtained by oxidative roasting using lead molybdate ore as raw material; 4g of lead molybdate calcined sand and 20mL of sulfuric acid solution with a concentration of 100g / L were added to the reaction vessel and stirred for 30min at 40℃ and 400r / min. Then, 0.8g of phosphoric acid as a leaching aid was added and leached for 2h. Then, 0.2g of modified flocculant was added and stirred for another 40min. After standing for 10min, a mixed solution was obtained.

[0022] S5: 30% by volume of P507 and 70% by volume of sulfonated kerosene are stirred and mixed to obtain an extractant. The extractant is extracted for 1 hour under the conditions of an extraction ratio of O / A=2:1, countercurrent extraction, and 5 extraction stages to obtain molybdenum-loaded P507. The molybdenum-loaded P507 is back-extracted with 10% by mass of hydrogen peroxide under the conditions of an extraction ratio of O / A=5:1, countercurrent extraction, and 3 extraction stages to obtain molybdenum by evaporation and crystallization. This completes a process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate.

[0023] Example 2: A process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate, comprising the following steps: S1: Add 55g of 2,5-diaminoterephthalic acid and 575mL of N,N-dimethylformamide solution to a polytetrafluoroethylene-lined autoclave. Stir for 35min at 22.5℃ and 550r / min. Then add 42.5g of ferric chloride and 35g of PVDF powder. Heat to 125℃ and continue the reaction for 25h. Cool naturally to room temperature, filter, and wash the filter cake three times with methanol solution and deionized water respectively. Dry under vacuum at 65℃ for 1.5h to obtain Ti / PVDF powder.

[0024] S2: Add 110g of aminosulfonic acid and 125g of urea to the reaction vessel and stir for 35min at 72.5℃ and 450r / min. Then add 17.5g of potato starch, heat to 95℃, and continue the reaction for 5.5h. After natural cooling, filter and collect the precipitate. Wash the precipitate three times with deionized water and anhydrous ethanol, and dry it under vacuum at 65℃ for 1.5h to obtain sulfonated starch.

[0025] S3: Add 11g of sulfonated starch and 110mL of deionized water to a reaction vessel and stir for 35min at 72.5℃ and 450r / min. Then add 3.5g of Ti / PVDF powder, heat to 65℃, and ultrasonically disperse for 35min. Then add 35g of acrylamide, 4.5g of N,N'-methylenebisacrylamide, and 0.85g of potassium persulfate and continue stirring until a gel is formed. Cut the gel into small pieces and wash it in distilled water. After washing the gel, dry it at 65℃ for 21h, pulverize it and pass it through a 210-mesh sieve to obtain the modified flocculant.

[0026] S4: Lead molybdate calcined sand with a molybdenum content of 12.32% and a lead content of 42% was obtained by oxidative roasting using lead molybdate ore as raw material; 4.5g of lead molybdate calcined sand and 25mL of sulfuric acid solution with a concentration of 100g / L were added to the reaction vessel and stirred for 35min at 45℃ and 450r / min. Then, 0.85g of phosphoric acid as a leaching aid was added and leached for 2.5h. Then, 0.25g of modified flocculant was added and stirred for another 45min. After standing for 11min, a mixed solution was obtained.

[0027] S5: 31% P507 (volume fraction) and 71% sulfonated kerosene (volume fraction) were stirred and mixed to obtain an extractant. The extractant was extracted for 1.5 h under the conditions of an extraction ratio of O / A=2:1, countercurrent extraction, and 5 extraction stages to obtain molybdenum-loaded P507. The molybdenum-loaded P507 was back-extracted with 11% hydrogen peroxide (mass fraction) under the conditions of an extraction ratio of O / A=5:1, countercurrent extraction, and 3 extraction stages to obtain molybdenum by evaporation and crystallization. This completes a process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate.

[0028] Example 3: A process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate, comprising the following steps: S1: Add 60g of 2,5-diaminoterephthalic acid and 600mL of N,N-dimethylformamide solution to a polytetrafluoroethylene-lined autoclave. Stir for 40min at 25℃ and 600r / min. Then add 50g of ferric chloride and 40g of PVDF powder. Heat to 130℃ and continue the reaction for 26h. Cool naturally to room temperature, filter, and wash the filter cake four times with methanol solution and deionized water, respectively. Dry under vacuum at 70℃ for 2h to obtain Ti / PVDF powder.

[0029] S2: Add 120g of aminosulfonic acid and 130g of urea to the reaction vessel, stir for 40min at 75℃ and 500r / min, then add 20g of potato starch, heat to 100℃, continue to react for 6h, cool naturally, filter, collect the precipitate, wash the precipitate 4 times with deionized water and anhydrous ethanol, and dry under vacuum at 70℃ for 2h to obtain sulfonated starch.

[0030] S3: Add 12g of sulfonated starch and 120mL of deionized water to a reaction vessel and stir for 40min at 75℃ and 500r / min. Then add 4g of Ti / PVDF powder, heat to 70℃, and ultrasonically disperse for 40min. Then add 40g of acrylamide, 5g of N,N'-methylenebisacrylamide, and 0.9g of potassium persulfate and continue stirring until a gel is formed. Cut the gel into small pieces and wash them in distilled water. After washing the gel, dry it at 70℃ for 22h, pulverize it through a 220-mesh sieve, and obtain the modified flocculant.

[0031] S4: Lead molybdate calcined sand with a molybdenum content of 12.32% and a lead content of 42% was obtained by oxidative roasting using molybdenum-lead ore as raw material; 5g of lead molybdate calcined sand and 30mL of sulfuric acid solution with a concentration of 100g / L were added to the reaction vessel and stirred for 40min at 50℃ and 500r / min. Then, 0.9g of phosphoric acid as a leaching aid was added and leached for 3h. Then, 0.3g of modified flocculant was added and stirred for another 50min. After standing for 12min, a mixed solution was obtained.

[0032] S5: 32% P507 (volume fraction) and 72% sulfonated kerosene (volume fraction) were stirred and mixed to obtain an extractant. The extractant was extracted for 2 hours under the conditions of an extraction ratio of O / A=2:1, countercurrent extraction, and 5 extraction stages to obtain molybdenum-loaded P507. The molybdenum-loaded P507 was back-extracted with 12% hydrogen peroxide (mass fraction) under the conditions of an extraction ratio of O / A=5:1, countercurrent extraction, and 3 extraction stages to obtain molybdenum by evaporation and crystallization. This completes a process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate.

[0033] Comparative Example 1: Based on Example 3, the Ti / PVDF powder in step S3 was replaced with PVDF powder.

[0034] Comparative Example 2: Based on Example 3, the sulfonated starch in step S3 was omitted.

[0035] Comparative Example 3: Based on Example 3, the modified flocculant in step S4 was omitted.

[0036] The concentrations of recovered molybdenum provided in Examples 1-3 and Comparative Examples 1-3 were tested, and the results are shown in Table 1: Table 1

[0037] In Comparative Example 1, when Ti / PVDF powder was replaced with PVDF powder, fine impurities remaining in the leachate entered the extraction stage with the solution, contaminating the P507 extractant and causing a decrease in molybdenum extraction rate and a decrease in the structural stability of the flocculant. The combination of iron-based metal-organic framework and PVDF can enhance the binding force between the powder and sulfonated starch and polyacrylamide. Pure PVDF has poor compatibility, and the flocculant network structure is easy to loosen, further reducing the impurity removal efficiency.

[0038] In Comparative Example 2, the sulfonated starch in step S3 was omitted. PVDF and polyacrylamide alone could not remain stable in 100 g / L sulfuric acid. The polyacrylamide chains are prone to protonation under strong acidity, causing the flocculant network structure to disintegrate and failing to form effective flocs, resulting in a sharp drop in impurity removal efficiency. Furthermore, the lack of hydrophilic groups from sulfonated starch caused the flocculant to easily agglomerate into clumps in the leachate, preventing sufficient contact with fine impurities and further reducing the impurity removal effect. The absence of sulfonated starch also caused an imbalance in the surface charge of the flocculant. The small amount of residual amide groups in the polyacrylamide may carry a weak positive charge, which can electrostatically adsorb with the negatively charged molybdate ions, leading to the precipitation and loss of some molybdenum along with the impurities.

[0039] In Comparative Example 3, the modified flocculant in step S4 was omitted, causing impurities to directly enter the extraction system. The fine impurities in the leachate formed a stable emulsion layer with the P507 extractant, clogging the extraction equipment. Furthermore, the impurities competitively adsorbed the active sites of P507. Even if some molybdenum was extracted, the impurities would still enter the back-extraction liquid with the extractant, resulting in the molybdenum product purity not meeting the standards during evaporation and crystallization. Moreover, the molybdenum concentration in the back-extraction liquid decreased significantly due to the low extraction efficiency.

[0040] Although embodiments of the invention have been shown and described, it will be understood by those skilled in the art that various changes, modifications, substitutions and variations can be made to these embodiments without departing from the principles and spirit of the invention.

Claims

1. A process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate, characterized in that, Includes the following steps: Step 1: Using molybdenum-lead ore as raw material, oxidative roasting is performed to obtain lead molybdate roasted sand with a molybdenum content of 12.32% and a lead content of 42%. The lead molybdate roasted sand and a sulfuric acid solution with a concentration of 100g / L are added to a reaction vessel and stirred for 30-40 minutes at 40-50℃ and 400-500r / min. Then, phosphoric acid is added as a leaching aid and leached for 2-3 hours. Then, a modified flocculant is added and stirring is continued for 40-50 minutes. After standing for 10-12 minutes, a mixed solution is obtained.

2. Step Two: Mix 30-32% (v / v) of P507 and 70-72% (v / v) of sulfonated kerosene to obtain an extractant. Extract for 1-2 hours under the conditions of an extraction ratio of O / A=2:1, countercurrent extraction, and 5 extraction stages to obtain molybdenum-loaded P507. Back-extract the molybdenum-loaded P507 with 10-12% (w / w) of hydrogen peroxide under the conditions of an extraction ratio of O / A=5:1, countercurrent extraction, and 3 extraction stages to obtain molybdenum by evaporation and crystallization. This completes a process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate.

3. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 1, characterized in that, The ratio of lead molybdate calcined sand, sulfuric acid solution, phosphoric acid and modified flocculant is 4-5g: 20-30mL: 0.8-0.9g: 0.2-0.3g.

4. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 1, characterized in that, The specific preparation steps of the modified flocculant are as follows: Sulfonated starch and deionized water were added to a reaction vessel and stirred for 30-40 minutes at 70-75℃ and 400-500 r / min. Then Ti / PVDF powder was added, heated to 60-70℃, and ultrasonically dispersed for 30-40 minutes. Acrylamide, N,N'-methylenebisacrylamide, and potassium persulfate were then added, and stirring was continued until a gel was formed. The gel was cut into small pieces and washed in distilled water. After washing, the gel was dried at 60-70℃ for 20-22 hours and then pulverized through a 200-220 mesh sieve to obtain the modified flocculant.

5. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 2, characterized in that, The ratio of sulfonated starch, deionized water, Ti / PVDF powder, acrylamide, N,N'-methylenebisacrylamide and potassium persulfate is 10-12g:100-120mL:3-4g:30-40g:4-5g:0.8-0.9g.

6. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 1, characterized in that, The specific preparation steps for the sulfonated starch are as follows: Add aminosulfonic acid and urea to a reaction vessel and stir for 30-40 minutes at 70-75℃ and 400-500 r / min. Then add potato starch, heat to 90-100℃, and continue the reaction for 5-6 hours. After natural cooling, filter and collect the precipitate. Wash the precipitate 2-4 times with deionized water and anhydrous ethanol, and dry it under vacuum at 60-70℃ for 1-2 hours to obtain sulfonated starch.

7. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 2, characterized in that, The ratio of aminosulfonic acid, urea and potato starch used is 100-120g: 120-130g: 15-20g.

8. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 1, characterized in that, The specific preparation steps for the Ti / PVDF powder are as follows: A solution of 2,5-diaminoterephthalic acid and N,N-dimethylformamide was added to a polytetrafluoroethylene-lined autoclave and stirred for 30-40 minutes at 20-25°C and 500-600 r / min. Then, ferric chloride and PVDF powder were added, and the mixture was heated to 120-130°C and reacted for 24-26 hours. The mixture was then allowed to cool naturally to room temperature, filtered, and the filter cake was washed 2-4 times with methanol solution and deionized water, respectively. The cake was then vacuum dried at 60-70°C for 1-2 hours to obtain Ti / PVDF powder.

9. The process for recovering molybdenum from high-acid, high-ion-concentration white flue dust leachate according to claim 7, characterized in that, The ratio of 2,5-diaminoterephthalic acid, N,N-dimethylformamide solution, ferric chloride and PVDF powder is 50-60g: 550-600mL: 35-50g: 30-40g.

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