A method for recovering molybdenum from coal direct liquefaction oil residue acid system leach liquor

Abstract

The application discloses a method for recovering molybdenum from coal direct liquefaction oil residue acid system leaching liquor, and belongs to the technical field of hydrometallurgy and catalyst recovery. The method comprises the following steps: S1, using a mixed extractant containing a compound extractant and a diluent, extracting the coal direct liquefaction oil residue acid system leaching liquor, stirring, standing and separating, to obtain a mixed extractant loaded with molybdenum and a raffinate water phase, wherein the coal direct liquefaction oil residue acid system leaching liquor comprises Fe elements, Mo elements, Al elements, Ca elements and sulfate radicals; S2, using alkali to back-extract the mixed extractant loaded with molybdenum, stirring, standing and separating, to obtain an ammonium molybdate solution and a regenerated mixed extractant, and the regenerated mixed extractant is used in S1; S3, evaporating and concentrating the ammonium molybdate solution, cooling and crystallizing, and drying, to obtain a high-purity ammonium molybdate product. The method can realize a molybdenum recovery rate of not less than 98%, and the operation process is relatively simple.

Description

Technical Field

[0001] This application relates to a method for recovering molybdenum from the leachate of an acidic system for direct coal liquefaction oil residue, belonging to the field of hydrometallurgy and catalyst recovery technology. Background Technology

[0002] In the direct coal liquefaction process, pulverized coal undergoes hydrocracking under high temperature and pressure conditions using a molybdenum-iron composite catalyst. The molybdenum-iron catalyst is enriched in the coal liquefaction residue. This residue is then acid-leached to obtain a molybdenum-containing acid leachate. Molybdenum, as a high-value non-ferrous metal, has significant recovery value. Currently, for molybdenum recovery technologies from acid-based leachates of various solid wastes, extraction is a relatively mature and widely used separation technology. For example, CN110817944B uses trioctylamine or di-(2-hexylhexyl)phosphoric acid extractants to simultaneously extract vanadium, tungsten, and molybdenum, followed by stepwise back-extraction of each element. However, this technology's extractants cannot selectively extract molybdenum. CN118957277 A discloses a method for separating molybdenum from copper, zinc, and iron. This method involves extracting and separating molybdenum from copper flue dust leachate using the molybdenum extraction reagent MP136. MP136 is prepared by reacting sodium cellulose, glycerol, and 1-ethyl-(3-dimethylaminopropyl)carbodiimide (EDC) in a formulated ratio under pressure in an oxygen atmosphere. This method is protected by patent. Patent CN120591557A uses a trioctyldecyl tertiary amine extractant to extract the acidic leachate from coal direct liquefaction oil residue. Fe and Mo are adsorbed by the extractant and separated from the aqueous phase. Then, the extractant loaded with Fe and Mo is back-extracted with ammonia water, causing Fe to precipitate as Fe(OH)3, while Mo remains in the solution as ammonium molybdate, thus completing the separation of Fe and Mo. Finally, the aqueous phase containing ammonium molybdate is evaporated and crystallized to obtain ammonium molybdate, with a molybdenum recovery rate of not less than 95%.

[0003] The above techniques have the limitation that the extractant cannot selectively adsorb molybdenum. In general, multiple elements are adsorbed simultaneously before subsequent separation, making the operation process relatively complex. Summary of the Invention

[0004] To address the aforementioned technical problems, this application provides a method for recovering molybdenum from the leachate of a coal direct liquefaction oil residue acid system. The method includes the following steps: S1, using a mixed extractant containing a compound extractant and a diluent to extract the coal direct liquefaction oil residue acid system leachate, stirring, allowing it to stand and separate, obtaining a molybdenum-loaded mixed extractant and a raffinate aqueous phase, wherein the coal direct liquefaction oil residue acid system leachate includes Fe, Mo, Al, Ca, and sulfate elements; S2, back-extracting the molybdenum-loaded mixed extractant with alkali, stirring, allowing it to stand and separate, obtaining an ammonium molybdate solution and a regenerated mixed extractant, the regenerated mixed extractant being reused in S1; S3, evaporating and concentrating the ammonium molybdate solution, cooling and crystallizing, and drying to obtain a high-purity ammonium molybdate product. This method can achieve a molybdenum recovery rate of not less than 98%, and the operation process is relatively simple.

[0005] According to one aspect of this application, a method for recovering molybdenum from the leachate of an acidic system for direct coal liquefaction residue is provided.

[0006] A method for recovering molybdenum from the leachate of an acidic system for direct coal liquefaction residue, the method comprising the following steps: S1. Using a mixed extractant containing compound extractant and diluent, extract the acid leaching solution of the coal direct liquefaction oil residue system, stir, let stand to separate the layers, and separate to obtain a mixed extractant loaded with molybdenum and a raffinate aqueous phase, wherein the acid leaching solution of the coal direct liquefaction oil residue system includes Fe, Mo, Al, Ca and sulfate elements. S2. The molybdenum-loaded mixed extractant is back-extracted with alkali, stirred, allowed to stand and separate into layers, and then separated to obtain an ammonium molybdate solution and a regenerated mixed extractant, which is then reused in S1. S3. The ammonium molybdate solution is evaporated, concentrated, cooled, crystallized, and dried to obtain a high-purity ammonium molybdate product.

[0007] Optionally, in step S1, the compound extractant includes trioctyldecyl tertiary amine and 2-ethylhexyl phosphate mono-2-ethylhexyl ester.

[0008] Optionally, in step S1, the diluent is white oil and / or kerosene.

[0009] Optionally, in step S1, the volume fraction of the trioctyldecyl tertiary amine in the mixed extractant is 5-15%, and the volume fraction of the 2-ethylhexyl phosphate mono-2-ethylhexyl ester in the mixed extractant is 10-25%.

[0010] Optionally, the volume fraction of the trioctyldecyl tertiary amine in the mixed extractant is any value from 5%, 6%, 8%, 9%, 10%, 12%, 14%, 15%, or any range between both.

[0011] Optionally, the volume fraction of the 2-ethylhexyl phosphate mono-2-ethylhexyl ester in the mixed extractant is any value from 10%, 12%, 14%, 15%, 18%, 20%, 22%, 25%, or any range between both.

[0012] Optionally, in step S1, the volume ratio of the mixed extractant to the leachate from the coal direct liquefaction oil residue acid system is 0.3 to 1:1.

[0013] Optionally, the volume ratio of the mixed extractant to the leachate from the coal direct liquefaction oil residue acid system is any value from 0.3:1, 0.4:1, 0.5:1, 0.6:1, 0.7:1, 0.8:1, 0.9:1, 1:1, or any value between the two.

[0014] Optionally, in step S1, the extraction temperature is 25–45°C.

[0015] Optionally, in step S1, the stirring speed is 120–550 r / min and the stirring time is 5–20 min.

[0016] Optionally, in step S1, the settling time is 5 to 20 minutes.

[0017] Optionally, in step S1, the extraction is performed at least once.

[0018] Optionally, in step S2, the alkali is ammonia water, and the concentration of the ammonia water is 5-15 wt%.

[0019] Optionally, the concentration of the ammonia water is any value among 5wt%, 6wt%, 8wt%, 10wt%, 12wt%, 13wt%, 14wt%, and 15wt%, or a range between any two.

[0020] Optionally, in step S2, the volume ratio of the molybdenum-loaded mixed extractant to the ammonia water is 1~2:1.

[0021] Optionally, the volume ratio of the molybdenum-loaded mixed extractant to the ammonia water is any value or a range between 1:1, 1.2:1, 1.3:1, 1.5:1, 1.6:1, 1.8:1, 1.9:1, and 2:1.

[0022] Optionally, in step S2, the temperature of the back-extraction is 30–50°C.

[0023] Optionally, in step S2, the stirring speed is 60-200 r / min and the stirring time is 2-10 min.

[0024] Optionally, in step S2, the settling time is 10 to 50 minutes.

[0025] Optionally, in step S2, the back-extraction is performed at least once.

[0026] Optionally, in step S2, the regenerated mixed extractant is used 4 to 6 times.

[0027] Optionally, in step S3, the evaporation and concentration temperature is 80-110°C, and the solution concentration is concentrated to 20-30 wt%.

[0028] Optionally, in step S3, the temperature for cooling crystallization is 20–30°C.

[0029] Optionally, in step S3, the drying temperature is 60–80°C and the time is 2–4 hours.

[0030] Optionally, in step S3, the purity of the ammonium molybdate product is ≥99.5%.

[0031] The beneficial effects that this application can produce include: The mixed extractant in this application exhibits minimal iron adsorption but a very high molybdenum adsorption capacity, enabling selective molybdenum adsorption without the need for additional iron separation. Furthermore, due to the minimal iron adsorption, the mixed extractant does not prematurely saturate due to iron adsorption, extending its lifespan and reducing consumption. This makes it suitable for large-scale continuous industrial production, balancing high efficiency and economy. Simultaneously, after back-extraction with ammonia, the saturated mixed extractant retains molybdenum in the aqueous phase as ammonium molybdate, which can be evaporated and crystallized to obtain high-purity ammonium molybdate. The regenerated mixed extractant can be recycled. This method achieves a molybdenum recovery rate of at least 98%, and the operation process is relatively simple. Attached Figure Description

[0032] Figure 1 This is a flowchart of the method for recovering molybdenum from the acid leaching solution of coal direct liquefaction residue in Examples 1-4. Detailed Implementation

[0033] The present application is described in detail below with reference to the embodiments, but the present application is not limited to these embodiments.

[0034] Unless otherwise specified, the raw materials and reagents used in the embodiments of this application were all purchased commercially.

[0035] The leaching solution of the coal direct liquefaction oil residue acid system in this application is the leaching solution obtained in step S3 of patent CN120591557A.

[0036] Unless otherwise specified, all test methods are standard and all instrument settings are those recommended by the manufacturer.

[0037] In the examples, N235 refers to trioctyldecyl tertiary amine, and P507 refers to 2-ethylhexyl phosphate mono-2-ethylhexyl ester.

[0038] Example 1 Methods for recovering molybdenum from the leachate of coal direct liquefaction residue acid system, such as... Figure 1 As shown: S1. Preparation of mixed extractant: Mix N235, P507 and white oil in a volume ratio of 10%, 20% and 70% to obtain a mixed extractant.

[0039] S2. Extraction: Take 300 mL of the acid leaching solution from the coal direct liquefaction residue system (Mo concentration 538 mg / L, Fe concentration 20.5 g / L) and place it in a 1000 mL Erlenmeyer flask. Add 150 mL of the above mixed extractant (extractant to acid leaching solution volume ratio O / A = 0.5:1). Place the flask in a constant temperature water bath shaker, control the temperature at 35℃, and the stirring speed at 500 r / min. Shake and stir for 10 min to complete the selective adsorption of molybdenum. After removal, place the flask in a separatory funnel and let it stand for 10 min to separate the layers, obtaining the molybdenum-loaded mixed extractant and the raffinate aqueous phase. ICP analysis showed that the Fe concentration in the raffinate aqueous phase was 19.9 g / L and the Mo concentration was 0.31 mg / L.

[0040] S3, Back-extraction: Place the molybdenum-loaded mixed extractant (approximately 150 mL) collected in S2 into a 500 mL Erlenmeyer flask, slowly add 150 mL of 10 wt% ammonia solution (extractant to ammonia volume ratio O / A = 1:1), place it in a constant temperature water bath shaker, control the temperature at 40℃, stir at 100 r / min, shake and stir for 5 min, remove the separatory funnel and let it stand for 30 min to separate the layers, obtaining ammonium molybdate solution and regenerated mixed extractant; use two-stage extraction to completely back-extract Mo into the aqueous phase.

[0041] S4. Evaporation and crystallization: The ammonium molybdate solution (about 150 mL) obtained in S3 was transferred to a 250 mL beaker and placed on an electric thermostatic water bath. The evaporation temperature was controlled at 95 °C, and the solution was concentrated to a concentration of 25 wt%. Then, it was cooled to 25 °C for crystallization. After centrifugation and dehydration, it was placed in an electric thermostatic drying oven and dried at 70 °C for 3 h to obtain 0.262 g of ammonium tetramolybdate product.

[0042] Molybdenum recovery calculation: Total Mo in the original acid leaching solution = 300 mL × 538 mg / L = 161.4 mg; Mo content in the final ammonium tetramolybdate product = 0.262 g × 61.13% = 160.2 mg; Molybdenum recovery rate = (160.2 mg ÷ 161.4 mg) × 100% = 99.2%.

[0043] Example 2 Methods for recovering molybdenum from the leachate of coal direct liquefaction residue acid system, such as... Figure 1 As shown: S1. Preparation of mixed extractant: Mix N235, P507 and kerosene in a volume fraction of 5%, 10% and 85% to obtain a mixed extractant.

[0044] S2. Extraction: Take 300 mL of the acid leaching solution from the coal direct liquefaction oil residue system (Mo concentration 480 mg / L, Fe concentration 18.2 g / L) and place it in a 1000 mL Erlenmeyer flask. Add 90 mL of the above mixed extractant (extractant to acid leaching solution volume ratio O / A = 0.3:1). Place the flask in a constant temperature water bath shaker, control the temperature at 25℃, and the stirring speed at 120 r / min. Shake and stir for 5 min to complete the selective adsorption of molybdenum. After removal, place the flask in a separatory funnel and let it stand for 5 min to separate the layers, obtaining the molybdenum-loaded mixed extractant and the raffinate aqueous phase. ICP analysis showed that the Fe concentration in the raffinate aqueous phase was 18.1 g / L and the Mo concentration was 0.75 mg / L.

[0045] S3, Back-extraction: Place the molybdenum-loaded mixed extractant (approximately 90 mL) collected in S2 into a 500 mL Erlenmeyer flask, slowly add 90 mL of 5 wt% ammonia water (extractant to ammonia water volume ratio O / A = 1:1), place it in a constant temperature water bath shaker, control the temperature at 30℃, stir at 60 r / min, shake and stir for 2 min, remove the separatory funnel and let it stand for 10 min to separate the layers, obtaining ammonium molybdate solution and regenerated mixed extractant; use two-stage extraction to completely back-extract Mo into the aqueous phase.

[0046] S4. Evaporation and crystallization: The ammonium molybdate solution (about 90 mL) obtained in S3 was transferred to a 250 mL beaker and placed on an electric thermostatic water bath. The evaporation temperature was controlled at 80 °C, and the solution was concentrated to a concentration of 20 wt%. Then, it was cooled to 20 °C for crystallization. After centrifugation and dehydration, it was placed in an electric thermostatic drying oven and dried at 60 °C for 4 h to obtain 0.228 g of ammonium tetramolybdate product.

[0047] Molybdenum recovery calculation: Total Mo in the original acid leaching solution = 300 mL × 480 mg / L = 144.0 mg; Mo content in the final ammonium tetramolybdate product = 0.232 g × 61.13% = 141.82 mg; Molybdenum recovery rate = (141.82 mg ÷ 144.0 mg) × 100% = 98.5%.

[0048] Example 3 Methods for recovering molybdenum from the leachate of coal direct liquefaction residue acid system, such as... Figure 1 As shown: S1. Preparation of mixed extractant: Mix N235, P507 and white oil in a volume ratio of 15%, 25% and 60%, and stir to obtain mixed extractant.

[0049] S2. Extraction: Take 300 mL of the acid leaching solution from the coal direct liquefaction oil residue system (Mo concentration 580 mg / L, Fe concentration 22.3 g / L) and place it in a 1000 mL Erlenmeyer flask. Add 300 mL of the above mixed extractant (extractant to acid leaching solution volume ratio O / A = 1:1). Place the flask in a constant temperature water bath shaker, control the temperature at 45℃, and the stirring speed at 550 r / min. Shake and stir for 20 min to complete the selective adsorption of molybdenum. After removal, place the flask in a separatory funnel and let it stand for 20 min to separate the molybdenum-loaded mixed extractant and the raffinate aqueous phase. ICP analysis showed that the Fe concentration in the raffinate aqueous phase was 22.2 g / L and the Mo concentration was 0.2 mg / L.

[0050] S3, Back-extraction: The molybdenum-loaded mixed extractant (approximately 300 mL) collected in S2 was placed in a 500 mL Erlenmeyer flask, and 150 mL of 15 wt% ammonia solution was slowly added (extractant to ammonia solution volume ratio O / A = 2:1). The flask was placed in a constant temperature water bath shaker, and the temperature was controlled at 50 °C. The stirring speed was 200 r / min, and the mixture was shaken for 10 min. The separatory funnel was then removed and allowed to stand for 50 min to separate the layers, thus obtaining the ammonium molybdate solution and the regenerated mixed extractant. Mo was completely back-extracted into the aqueous phase using a two-stage extraction process.

[0051] S4. Evaporation and crystallization: The ammonium molybdate solution (about 150 mL) obtained in S3 was transferred to a 250 mL beaker and placed on an electric thermostatic water bath. The evaporation temperature was controlled at 110 °C and the solution was concentrated to a concentration of 30 wt%. Then, it was cooled to 30 °C for crystallization. After centrifugation and dehydration, it was placed in an electric thermostatic drying oven and dried at 80 °C for 2 hours to obtain 0.272 g of ammonium tetramolybdate product.

[0052] Molybdenum recovery calculation: Total Mo in the original acid leaching solution = 300mL × 580mg / L = 174.0mg; Mo content in the final ammonium tetramolybdate product = 0.281g × 61.13% = 171.8mg; Molybdenum recovery rate = (171.8mg ÷ 174.0mg) × 100% = 98.8%.

[0053] Example 4 Methods for recovering molybdenum from the leachate of coal direct liquefaction residue acid system, such as... Figure 1 As shown: S1. Preparation of mixed extractant: Mix N235, P507 and kerosene in a volume fraction of 7%, 18% and 75% to obtain a mixed extractant.

[0054] S2. Extraction: Take 300 mL of the acid leaching solution from the coal direct liquefaction oil residue system (Mo concentration 510 mg / L, Fe concentration 19.6 g / L) and place it in a 1000 mL Erlenmeyer flask. Add 180 mL of the above mixed extractant (extractant to acid leaching solution volume ratio O / A = 0.6:1). Place the flask in a constant temperature water bath shaker, control the temperature at 30℃, and the stirring speed at 300 r / min. Shake and stir for 12 min to complete the selective adsorption of molybdenum. After removal, place the flask in a separatory funnel and let it stand for 12 min to separate the molybdenum-loaded mixed extractant and the raffinate aqueous phase. ICP analysis showed that the Fe concentration in the raffinate aqueous phase was 19.5 g / L and the Mo concentration was 0.42 mg / L.

[0055] S3, Back-extraction: Place the molybdenum-loaded mixed extractant (approximately 180 mL) collected in S2 into a 500 mL Erlenmeyer flask, slowly add 180 mL of 8 wt% ammonia solution (extractant to ammonia volume ratio O / A = 1:1), place it in a constant temperature water bath shaker, control the temperature at 35℃, stir at 120 r / min, shake and stir for 6 min, remove the separatory funnel and let it stand for 20 min to separate the layers, obtaining ammonium molybdate solution and regenerated mixed extractant; use two-stage extraction to completely back-extract Mo into the aqueous phase.

[0056] S4. Evaporation and crystallization: The ammonium molybdate solution (about 180 mL) obtained in S3 was transferred to a 250 mL beaker and placed on an electric thermostatic water bath. The evaporation temperature was controlled at 90 °C, and the solution was concentrated to a concentration of 22 wt%. Then, it was cooled to 22 °C for crystallization. After centrifugation and dehydration, it was placed in an electric thermostatic drying oven and dried at 65 °C for 3 h to obtain 0.251 g of ammonium tetramolybdate product.

[0057] Molybdenum recovery calculation: Total Mo in the original acid leaching solution = 300 mL × 510 mg / L = 153.0 mg; Mo content in the final ammonium tetramolybdate product = 0.251 g × 61.13% = 153.4 mg; Molybdenum recovery rate = (153.4 mg ÷ 153.0 mg) × 100% = 98.9%.

[0058] Comparative Example 1 S1. Preparation of extractant: P507 and white oil are mixed evenly at a volume fraction of 20% and 80% to obtain a single P507 extractant.

[0059] S2. Extraction: Take 300 mL of the acid leaching solution from the coal direct liquefaction oil residue system (Mo concentration 538 mg / L, Fe concentration 20.5 g / L) and place it in a 1000 mL Erlenmeyer flask. Add 150 mL of the above-mentioned single P507 extractant (extractant to acid leaching solution volume ratio O / A = 0.5:1). Place the flask in a constant temperature water bath shaker, control the temperature at 35℃, and the stirring speed at 500 r / min. Shake and stir for 10 min to complete the adsorption. After removal, place the flask in a separatory funnel and let it stand for 10 min to separate the Fe and Mo loaded extractant and the raffinate aqueous phase. ICP detection showed that the Fe concentration in the raffinate aqueous phase was 5.2 g / L and the Mo concentration was 4.8 mg / L.

[0060] S3. Back-extraction: The Fe and Mo loaded extractant (approximately 150 mL) collected in S2 was placed in a 500 mL Erlenmeyer flask. 150 mL of 10 wt% ammonia solution was slowly added (extractant to ammonia solution volume ratio O / A = 1:1). The flask was placed in a constant temperature water bath shaker, with the temperature controlled at 40℃ and the stirring speed at 100 r / min. The shaking was maintained for 5 min, and the stirring time was extended to 20 min. The back-extraction was repeated twice. The separatory funnel was removed and allowed to stand for 30 min to separate the layers, resulting in the separation of ammonium molybdate solution and extractant still loaded with a large amount of Fe. After testing, the Fe residue in the extractant after back-extraction was still as high as 8.6 g / L, and the iron was difficult to wash off, making regeneration and recycling impossible.

[0061] S4. Evaporation and crystallization: The ammonium molybdate solution (about 150 mL) obtained in S3 was transferred to a 250 mL beaker and placed on an electric thermostatic water bath. The evaporation temperature was controlled at 95 °C, and the solution was concentrated to a concentration of 25 wt%. Then, it was cooled to 25 °C for crystallization. After centrifugation and dehydration, it was placed in an electric thermostatic drying oven and dried at 70 °C for 3 h to obtain 0.253 g of ammonium tetramolybdate product.

[0062] Molybdenum recovery calculation: Total Mo in the original acid leaching solution = 300 mL × 538 mg / L = 161.4 mg; Mo content in the final ammonium tetramolybdate product = 0.253 g × 61.13% = 154.9 mg; Molybdenum recovery rate = (154.9 mg ÷ 161.4 mg) × 100% = 96.0%.

[0063] Comparative Example 2 S1. Preparation of mixed extractant: Mix N235, P507 and white oil in a volume ratio of 10%, 20% and 70% to obtain a mixed extractant.

[0064] S2. Extraction: Take 300 mL of the acid leaching solution from the coal direct liquefaction oil residue system (Mo concentration 538 mg / L, Fe concentration 20.5 g / L) and place it in a 1000 mL Erlenmeyer flask. Add 150 mL of the above mixed extractant (extractant to acid leaching solution volume ratio O / A = 0.5:1). Place the flask in a constant temperature water bath shaker, control the temperature at 35℃, and the stirring speed at 500 r / min. Shake and stir for 10 min to complete the selective adsorption of molybdenum. After removal, place the flask in a separatory funnel and let it stand for 10 min to separate the layers, obtaining the molybdenum-loaded mixed extractant and the raffinate aqueous phase. ICP analysis showed that the Fe concentration in the raffinate aqueous phase was 19.8 g / L and the Mo concentration was 0.35 mg / L.

[0065] S3. Back-extraction: The molybdenum-loaded mixed extractant collected in S2 (about 150 mL) was placed in a 500 mL Erlenmeyer flask, and 150 mL of 3 wt% ammonia water was slowly added (the ammonia water concentration is lower than 5-15 wt% in Examples 1-4). The flask was placed in a constant temperature water bath shaker, the temperature was controlled at 40°C, the stirring speed was 100 r / min, and the mixture was shaken for 5 min. The separatory funnel was then removed and allowed to stand for 30 min to separate the layers, thus obtaining the ammonium molybdate solution and the regenerated mixed extractant. Since secondary extraction was not used, the molybdenum back-extraction was incomplete.

[0066] S4. Evaporation and crystallization: The ammonium molybdate solution (about 150 mL) obtained in S3 was transferred to a 250 mL beaker and placed on an electric thermostatic water bath. The evaporation temperature was controlled at 95 °C, and the solution was concentrated to a concentration of 25 wt%. Then, it was cooled to 25 °C for crystallization. After centrifugation and dehydration, it was placed in an electric thermostatic drying oven and dried at 70 °C for 3 h to obtain 0.238 g of ammonium tetramolybdate product.

[0067] Molybdenum recovery calculation: Total Mo in the original acid leaching solution = 300 mL × 538 mg / L = 161.4 mg; Mo content in the final ammonium tetramolybdate product = 0.238 g × 61.13% = 145.5 mg; Molybdenum recovery rate = (145.5 mg ÷ 161.4 mg) × 100% = 90.2%.

[0068] The above description is merely a few embodiments of this application and is not intended to limit this application in any way. Although this application discloses preferred embodiments as described above, it is not intended to limit this application. Any changes or modifications made by those skilled in the art without departing from the scope of the technical solution of this application using the disclosed technical content are equivalent to equivalent implementation cases and fall within the scope of the technical solution.

Claims

1. A method for recovering molybdenum from the leachate of an acidic system for direct coal liquefaction residue, characterized in that, The method includes the following steps: S1. Using a mixed extractant containing compound extractant and diluent, extract the acid leaching solution of the coal direct liquefaction oil residue system, stir, let stand to separate the layers, and separate to obtain a mixed extractant loaded with molybdenum and a raffinate aqueous phase, wherein the acid leaching solution of the coal direct liquefaction oil residue system includes Fe, Mo, Al, Ca and sulfate elements. S2. The molybdenum-loaded mixed extractant is back-extracted with alkali, stirred, allowed to stand and separate into layers, and then separated to obtain an ammonium molybdate solution and a regenerated mixed extractant, which is then reused in S1. S3. The ammonium molybdate solution is evaporated, concentrated, cooled, crystallized, and dried to obtain a high-purity ammonium molybdate product.

2. The method according to claim 1, characterized in that, In step S1, the compound extractant includes trioctyldecyl tertiary amine and 2-ethylhexyl phosphate mono-2-ethylhexyl ester; Preferably, in step S1, the diluent is white oil and / or kerosene.

3. The method according to claim 2, characterized in that, In step S1, the volume fraction of the trioctyldecyl tertiary amine in the mixed extractant is 5-15%, and the volume fraction of the 2-ethylhexyl phosphate mono-2-ethylhexyl ester in the mixed extractant is 10-25%.

4. The method according to claim 1, characterized in that, In step S1, the volume ratio of the mixed extractant to the leachate from the coal direct liquefaction oil residue acid system is 0.3 to 1:

1.

5. The method according to claim 1, characterized in that, In step S1, the extraction temperature is 25–45°C; Preferably, in step S1, the stirring speed is 120-550 r / min and the stirring time is 5-20 min; Preferably, in step S1, the settling time is 5 to 20 minutes; Preferably, in step S1, the extraction is performed at least once.

6. The method according to claim 1, characterized in that, In step S2, the alkali is ammonia water, and the concentration of the ammonia water is 5-15 wt%.

7. The method according to claim 6, characterized in that, In step S2, the volume ratio of the molybdenum-loaded mixed extractant to the ammonia water is 1~2:

1.

8. The method according to claim 1, characterized in that, In step S2, the temperature of the back-extraction is 30–50°C; Preferably, in step S2, the stirring speed is 60-200 r / min and the stirring time is 2-10 min; Preferably, in step S2, the settling time is 10 to 50 minutes; Preferably, in step S2, the back-extraction is performed at least once; Preferably, in step S2, the regenerated mixed extractant is used 4 to 6 times.

9. The method according to claim 1, characterized in that, In step S3, the evaporation and concentration temperature is 80–110°C, and the solution concentration is reduced to 20–30 wt%. Preferably, in step S3, the temperature for cooling crystallization is 20–30°C; Preferably, in step S3, the drying temperature is 60-80°C and the time is 2-4 hours.

10. The method according to claim 1, characterized in that, In step S3, the purity of the ammonium molybdate product is ≥99.5%.

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