Method for recycling vanadium, iron and aluminum in vanadium shale

Abstract

The invention provides a method for recycling vanadium, iron and aluminum in vanadium shale, and relates to the technical field of non-ferrous metallurgy. The method comprises the following steps: step 1, ultrasonic leaching; step 2, iron precipitation; step 3, aluminum precipitation; and 4, vanadium enrichment and precipitation. According to the method, an ultrasonic enhanced leaching technology is utilized, crystal lattices of vanadium-containing minerals are efficiently destroyed under the condition that high-corrosivity fluoride / oxidizing agents are not added, and synchronous dissolution of vanadium, iron and part of aluminum is achieved. Compared with a current vanadium shale full-wet extraction technology, the method has the advantages that efficient leaching of the vanadium shale can be achieved under the mild condition, energy consumption and reagent consumption in the leaching process are reduced, and remarkable environmental and economic benefits are achieved.

Description

Technical Field

[0001] This invention relates to the field of non-ferrous metal metallurgy, and more particularly to a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale. Background Technology

[0002] Vanadium shale is a unique and advantageous vanadium resource in my country, with total reserves of approximately 61.88 billion tons, of which vanadium reserves account for about 87% of my country's total vanadium reserves. Green extraction and metallurgy of vanadium shale has become a top priority in the development and utilization of my country's vanadium resources. The development and utilization of vanadium shale is progressing in depth towards low-carbon, high-efficiency, green, and high-end directions.

[0003] The all-wet vanadium extraction technology has attracted much attention due to its low-carbon and clean characteristics, enabling the extraction of vanadium from vanadium-containing resources without high-temperature roasting. Since vanadium mainly exists in the mica mineral lattice as an isomorphous substitute for aluminum, the lattice of vanadium-containing mica minerals must be disrupted to extract vanadium. Mica minerals have stable lattices, and current all-wet extraction processes for vanadium shale mainly employ fluorine-based leaching aids and oxidant leaching techniques to disrupt the vanadium-containing mica mineral lattice and enhance vanadium release and dissolution. However, all-wet extraction technologies have the following common problems: 1) Low vanadium leaching efficiency, i.e., low vanadium leaching rate and long leaching cycle; 2) A large number of impurity ions, such as iron and aluminum, are dissolved along with vanadium during the enhanced leaching process, seriously affecting subsequent separation processes and limiting the purity of the final vanadium product; 3) The use of fluorine-based leaching aids and oxidants makes the treatment of acidic wastewater difficult. Meanwhile, current vanadium extraction processes from vanadium shale mainly focus on vanadium recovery, with very little attention paid to the recovery of associated iron and aluminum, resulting in a serious waste of valuable metal resources.

[0004] Therefore, it is urgent to solve the common problems existing in the traditional wet extraction process of vanadium shale and develop a new wet extraction technology to achieve efficient and green recovery of vanadium, while synergistically utilizing the associated iron and aluminum resources to avoid their waste. Summary of the Invention

[0005] In view of this, the present invention proposes a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale. The technical solution of the present invention is implemented as follows: This invention provides a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale, comprising the following steps: Step 1: Crush and grind the vanadium shale ore until the proportion of particles smaller than 200 mesh is not less than 75%, then mix it with sulfuric acid solution, leach it with ultrasonic waves, and separate the solid and liquid to obtain leachate. Step 2: Add hydrogen peroxide to the leachate, and after the oxidation reaction, add phosphate. Then adjust the pH of the solution to 1.0~2.0, and then react at a set temperature. Then perform solid-liquid separation to obtain crude iron phosphate and precipitated liquid A. Then purify the crude iron phosphate to prepare battery-grade iron phosphate. Step 3: Add phosphate to the precipitated solution A, adjust the pH of the mixed solution to 2.4~3.0, then react at a set temperature, and then perform solid-liquid separation to obtain industrial-grade aluminum phosphate and precipitated solution B; Step 4: Adjust the pH value of the precipitated liquid B, then extract and separate to obtain the loaded organic phase. The loaded organic phase is back-extracted to obtain a vanadium-rich solution, which is then treated with ammonium salt precipitation to prepare vanadium pentoxide product.

[0006] Further, in step one, the chemical composition of the vanadium shale ore, by mass percentage, includes: V2O5 0.5%~1.2%, Al2O3 6%~10%, Fe2O3 2%~5%, SiO2 45%~73%, K2O 1%~3%, MgO 0.5~2%, and CaO 0.3~5%.

[0007] Furthermore, in step one, the concentration of the sulfuric acid solution is 10~30 vol.%.

[0008] Furthermore, in step one, the liquid-to-solid ratio of the sulfuric acid solution to the vanadium shale ore is 1~5 mL / g.

[0009] Furthermore, in step one, during the ultrasonic leaching process, the ultrasonic power is set to 100~500 W, the leaching temperature is 25~100°C, and the leaching time is 0.5~4 h.

[0010] Furthermore, in step two, the amount of hydrogen peroxide added is 1 to 4 times the amount of ferrous ions in the leachate.

[0011] Furthermore, in step two, the oxidation reaction takes 10 to 60 minutes and the oxidation reaction temperature is 25 to 50°C.

[0012] Furthermore, in step two, the phosphate is selected from one or more of sodium phosphate, potassium phosphate, and ammonium phosphate; after adding the phosphate, the phosphorus-iron molar ratio in the solution is 1.0~1.6.

[0013] Furthermore, in step two, the set temperature is 25~120°C, and the reaction time is 0.1~3 h.

[0014] Further, in step two, the purification of the crude ferric phosphate specifically includes: The crude ferric phosphate and acid solution are mixed, followed by solid-liquid separation and drying to obtain refined ferric phosphate; wherein the pH value of the acid solution is 0.2~1.2; and / or, the liquid-solid ratio of the acid solution to the crude ferric phosphate is 1~10 mL / g; and / or, the temperature for preparing battery-grade ferric phosphate is 140~200°C and the time is 2~4 h.

[0015] Further, in step three, the phosphorus-aluminum molar ratio in the mixed solution is 1.5~2.0; the set temperature is 25~120°C; and the reaction time is 0.1~3 h.

[0016] Further, in step four, the extractant is selected from di(2-ethylhexyl) phosphate and / or 2-ethylhexylphosphonic acid mono(2-ethylhexyl) ester.

[0017] Furthermore, in step four, during the extraction process, the volume ratio of the organic phase containing the extractant to the precipitated liquid B is 0.25~1, and the extraction time is 5~20 min.

[0018] Further, in step four, the stripping agent is selected from dilute hydrochloric acid and / or dilute sulfuric acid; during the stripping process, the volume ratio of the supported organic phase to the aqueous phase containing the stripping agent is 1-4, and the extraction time is 5-20 min.

[0019] Furthermore, in step four, the preparation of vanadium pentoxide product from the vanadium-rich solution specifically includes: After adjusting the pH of the vanadium-rich solution to 2-3, ammonium salt is added, and after precipitation, solid-liquid separation is performed to obtain ammonium polyvanadate, which is then calcined to obtain vanadium pentoxide product. The ammonium salt is selected from ammonium chloride and / or ammonium sulfate, and the molar ratio of ammonium ions to vanadium ions in the ammonium salt is 2 to 4. The precipitation temperature is 20~30℃, and the precipitation time is 1~2 h; The calcination temperature is 500~600°C, and the calcination time is 30~120 min.

[0020] The present invention provides a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale, which has the following advantages over the prior art: (1) Green and efficient leaching: Utilizing ultrasonic-enhanced leaching technology, the crystal lattice of vanadium-containing minerals is efficiently destroyed without the addition of highly corrosive fluorides / oxidants, achieving simultaneous leaching of vanadium, iron, and some aluminum. Compared to current wet extraction techniques for vanadium shale, the method of this invention can achieve efficient leaching of vanadium shale under "mild" conditions, reducing energy and reagent consumption during the leaching process, and has significant environmental and economic benefits.

[0021] (2) Precise Separation and Enrichment of Multimetals: An innovative stepwise directional precipitation technology based on precise control of solution components is employed. Iron ions in the leachate are preferentially converted into crude ferric phosphate, and aluminum ions are recovered as industrial-grade aluminum phosphate. Finally, vanadium separation and precipitation are achieved through a simple extraction-precipitation process. This invention solves the problems of difficulty in preparing high-purity vanadium products and waste of associated iron and aluminum resources in the wet extraction process of vanadium shale, achieving precise gradient separation and enrichment of iron, aluminum, and vanadium.

[0022] (3) Resource utilization and high-value product preparation: Using the method of this invention, battery-grade iron phosphate, industrial-grade aluminum phosphate, and high-purity vanadium pentoxide products can be finally prepared. The entire process significantly improves the comprehensive utilization rate of vanadium shale resources and the added value of products. Attached Figure Description

[0023] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0024] Figure 1 This is a process flow diagram of the resource recovery of vanadium, iron, and aluminum from vanadium shale provided by the present invention. Detailed Implementation

[0025] 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 a part of the embodiments of the present invention, and not all of the 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.

[0026] like Figure 1 As shown, the present invention provides a method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale, specifically including the following steps: Step 1: Crush and grind the vanadium shale ore to a density of over 75% -200 mesh, then mix it with sulfuric acid solution at a certain liquid-solid ratio. After ultrasonic leaching at a certain temperature for a certain time, the solid and liquid are separated to obtain the leachate. In step one of the present invention, the chemical composition of the vanadium shale ore by mass percentage is: V2O5 0.5%~1.2%, Al2O3 6%~10%, Fe2O3 2%~5%, SiO2 45%~73%, K2O 1%~3%, MgO 0.5~2%, CaO 0.3~5%.

[0027] In step one of the present invention, the ultrasonic power during the ultrasonic leaching process is 100~500 W, more preferably 200~300 W.

[0028] In step one of this invention, the sulfuric acid concentration during ultrasonic leaching is 10-30 vol.%, more preferably 15-25 vol.%.

[0029] In step one of the present invention, the liquid-to-solid ratio during ultrasonic leaching is 1~5 mL / g, more preferably 3~4 mL / g.

[0030] In step one of the present invention, the temperature of ultrasonic immersion is 25~100℃, and more preferably 85~95℃.

[0031] In step one of the present invention, the leaching time is 0.5 to 4 hours, more preferably 1.5 to 2 hours.

[0032] Step 2: Add a certain amount of hydrogen peroxide to the leachate, and after the oxidation reaction has been carried out for a certain period of time, add a certain amount of phosphate to the mixed solution, adjust the pH value of the solution, and after the reaction has been carried out at a certain temperature for a certain period of time, the solid and liquid are separated to obtain crude iron phosphate and precipitated liquid A. Then, the crude iron phosphate is purified to prepare battery-grade iron phosphate. By employing the above-mentioned technical solution, based on the difference in solubility products between iron phosphate (FePO4) and aluminum phosphate (AlPO4) and their precipitation behavior under different acidity conditions, efficient and selective separation of iron and aluminum in solution can be achieved. From a thermodynamic perspective, the precipitation-dissolution equilibrium of sparingly soluble substances is significantly affected by the pH value of the solution. The solubility of both FePO4 and AlPO4 in acidic solutions decreases with increasing pH, but the pH ranges at which they reach minimum solubility differ. Through thermodynamic calculations and experimental verification, FePO4 can precipitate at relatively low pH conditions, while AlPO4 requires a relatively high pH environment to form stably. The precipitation of vanadium and phosphate is even more difficult; therefore, separating iron, aluminum, and vanadium using a precisely controlled stepwise directional precipitation technique with phosphate is feasible.

[0033] In the iron precipitation process of step two of this invention, the amount of hydrogen peroxide added before the iron precipitation process is 1 to 4 times the amount of ferrous ions, and more preferably 1.5 to 2.5 times.

[0034] In the iron precipitation process of step two of the present invention, the oxidation time before the iron precipitation process is 10~60 min, and more preferably 20~30 min.

[0035] In the iron precipitation process of step two of the present invention, the oxidation temperature before the iron precipitation process is 25~50°C, and more preferably 25~30°C.

[0036] In step two of the iron precipitation process of this invention, the added phosphate can be sodium phosphate, potassium phosphate, ammonium phosphate, etc.

[0037] In the iron precipitation process of step two of the present invention, the phosphorus-iron molar ratio of the solution before the iron precipitation process is controlled to be 1.0~1.6, and more preferably 1.2~1.4.

[0038] In the iron precipitation process of step two of the present invention, the pH of the iron precipitation process is adjusted to 1.0~2.0 using ammonia or sodium hydroxide solution, and more preferably 1.5~1.8.

[0039] In the iron precipitation process of step two of the present invention, the iron precipitation temperature is 25~120°C, and more preferably 90~120°C.

[0040] In step two of the present invention, the iron precipitation time is 0.1 to 3 hours, more preferably 0.2 to 1 hour.

[0041] In the iron precipitation process of step two of this invention, the purification process of crude iron phosphate includes: mixing crude iron phosphate and acid solution for reaction, followed by solid-liquid separation and drying to obtain refined iron phosphate.

[0042] In any embodiment of the present invention, the acid solution may be hydrochloric acid, sulfuric acid, nitric acid, etc., preferably hydrochloric acid; the initial pH value of the solution is controlled to be 0.2~1.2, more preferably 0.6~0.9.

[0043] In any embodiment of the present invention, the liquid-to-solid ratio of the acid solution to the crude ferric phosphate is 1~10 mL / g, more preferably 1~3 mL / g.

[0044] In any embodiment of the present invention, the temperature of the mixing reaction is 140~200℃, more preferably 150~180℃; the mixing reaction time is 2~4 h, more preferably 3~3.5 h. During this process, if the temperature is too high, energy consumption will increase; if the temperature is too low, a good impurity removal effect cannot be guaranteed.

[0045] Step 3: Adjust the pH value of the precipitated liquid A, and after reacting for a certain time at a certain temperature, separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated liquid B; In the aluminum precipitation process of step three of the present invention, the phosphorus-aluminum molar ratio of the solution before the aluminum precipitation process is controlled to be 1.5~2.0, and more preferably 1.6~1.8; In the aluminum precipitation process of step three of the present invention, the pH is adjusted to 2.4~3.0 using ammonia or sodium hydroxide solution, and more preferably 2.4~2.6.

[0046] In the aluminum precipitation process of step three of the present invention, the aluminum precipitation temperature is 25~120°C, and more preferably 90~120°C.

[0047] In the aluminum precipitation process of step three of the present invention, the aluminum precipitation time is 0.1~3 h, more preferably 0.2~1 h.

[0048] Step 4: Adjust the pH value of the precipitated liquid B, and use an extractant to extract and separate the vanadium in the precipitated liquid B to obtain a loaded organic phase. The loaded organic phase is back-extracted to obtain a vanadium-rich solution, and then the vanadium-rich solution is treated by ammonium salt precipitation to prepare vanadium pentoxide product.

[0049] The present invention will be further described below with reference to specific embodiments. The scope of protection of the present invention is not limited by the following embodiments. Unless otherwise specified, the materials mainly involved in the following examples are all conventional commercially available products.

[0050] The vanadium shale ore of this invention was obtained from vanadium shale in a certain area of ​​Jiangxi Province, and its chemical analysis results are shown in Table 1.

[0051] Table 1. Chemical composition of vanadium shale (wt.%)

[0052] Example 1 This embodiment provides a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale, specifically including: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-to-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0053] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0054] Example 2 This embodiment provides a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale, specifically including: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 90°C and an ultrasonic power of 250 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-to-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0055] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0056] Example 3 This embodiment provides a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale, specifically including: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.2. Adjust the pH of the solution to 1.6 using ammonia and react at 100°C for 0.5 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0057] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the vanadium extraction residue and the vanadium-loaded organic phase. Mix the vanadium-loaded organic phase with a 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0058] Example 4 This embodiment provides a method for the resource-based recovery of vanadium, iron, and aluminum from vanadium shale, specifically including: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (0.25 mol / L) with an initial pH of 0.6 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 160°C for 3 h to obtain refined ferric phosphate.

[0059] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.6 using sodium phosphate, and adjust the pH of the precipitated solution A to 2.4 using ammonia. React at 90°C for 1 hour. Separate the solid and liquid phases to obtain industrial-grade aluminum phosphate and precipitated solution B.

[0060] Step 4: Adjust the pH of the precipitate (B value) to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0061] Comparative Example 1 Compared to Example 1, the difference lies in the use of a traditional process to treat vanadium shale, namely, vanadium is directly extracted and enriched from the leachate after mechanical stirring leaching. Specifically: (1) Mechanical stirring leaching: The vanadium shale ore is crushed and ground to -200 mesh (75% or more), then mixed with 25 vol.% sulfuric acid solution at a liquid-solid ratio of 3 mL / g. The leaching is carried out at a leaching temperature of 95°C and mechanical stirring at a speed of 500 r / min for 120 min. The solid and liquid are then separated to obtain the leachate. (2) Vanadium enrichment and precipitation: The pH of the leachate was adjusted to 2.0 using sodium hydroxide. A mixture of 15 vol.% di(2-ethylhexyl) phosphate (P2O4) and 85 vol.% sulfonated kerosene was used as the extractant, and extraction was performed for 10 min at an extraction ratio of 0.5 (O:A). After phase separation, raffinate and loaded organic phase were obtained. The loaded organic phase was mixed with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extracted for 10 min to obtain vanadium-rich solution with a vanadium concentration of 18 g / L. The pH was adjusted to 2.5, and ammonium chloride was added to the vanadium-rich solution with an ammonium addition coefficient of 3. Precipitation was carried out at room temperature for 2 h. After solid-liquid separation, ammonium polyvanadate was obtained. Vanadium pentoxide was prepared by calcining it at 500°C for 1 h.

[0062] Comparative Example 2 Compared to Example 1, the only difference is that the pH value of the solution during the iron precipitation process is 1, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 minutes. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1 using ammonia and react at 120°C for 1 hour. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 hours to obtain refined ferric phosphate.

[0063] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the loaded organic phase. Mix the vanadium-loaded organic phase with a 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0064] Comparative Example 3 Compared to Example 1, the only difference is that the pH value of the solution during the iron precipitation process is 2.5, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 2.5 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0065] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the loaded organic phase. Mix the vanadium-loaded organic phase with a 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0066] Comparative Example 4 Compared to Example 1, the only difference is that the phosphorus-iron molar ratio in the solution during the iron precipitation process is set to 0.8, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 minutes. Then add sodium phosphate to the mixed solution to control the phosphorus-to-iron molar ratio of 0.8. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 hour. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 hours to obtain refined ferric phosphate.

[0067] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0068] Comparative Example 5 Compared to Example 1, the only difference is that the phosphorus-iron molar ratio in the solution during the iron precipitation process is set to 2.0, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 minutes. Then add sodium phosphate to the mixed solution to control the phosphorus-to-iron molar ratio at 2.0. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 hour. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 hours to obtain refined ferric phosphate.

[0069] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0070] Comparative Example 6 Compared to Example 1, the difference lies in the purification process of crude ferric phosphate, which uses a hydrochloric acid solution (approximately 1 mol / L) with an initial pH of 0. Specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with hydrochloric acid solution (approximately 0.1 mol / L) with the initial pH of 0 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0071] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0072] Comparative Example 7 Compared to Example 1, the difference lies in the purification process of crude ferric phosphate, which uses a hydrochloric acid solution (approximately 0.01 mol / L) with an initial pH of 2. Specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution with an initial pH of 2 (approximately 0.01 mol / L) at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0073] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1.8 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0074] Comparative Example 8 Compared to Example 1, the difference lies in that the pH of the aluminum precipitation process is set to 2, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0075] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1 using sodium phosphate, adjust the pH of the precipitated solution A to 2 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0076] Comparative Example 9 Compared to Example 1, the difference lies in the pH setting of the aluminum precipitation process to 3, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0077] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1 using sodium phosphate, adjust the pH of the precipitated solution A to 3 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0078] Comparative Example 10 Compared with Example 1, the difference lies in that the phosphorus-aluminum molar ratio in the aluminum precipitation process is set to 1, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0079] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 1 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0080] Comparative Example 11 Compared to Example 1, the difference lies in the phosphorus-aluminum molar ratio being set to 2.2 during the aluminum precipitation process, specifically: Step 1: Crush and grind the vanadium shale ore to a density of -200 mesh or higher (75%), then mix it with a 25 vol.% sulfuric acid solution at a liquid-to-solid ratio of 3 mL / g. Leach at a leaching temperature of 95°C and an ultrasonic power of 300 W for 120 min, and then separate the solid and liquid to obtain the leachate. Step 2: Add hydrogen peroxide with a ferrous ion content of 2 times to the leachate and oxidize at room temperature for 30 min. Then add sodium phosphate to the mixed solution to control the phosphorus-iron molar ratio of 1.4. Adjust the pH of the solution to 1.8 using ammonia and react at 120°C for 1 h. Separate the solid and liquid to obtain crude ferric phosphate and precipitate A. Mix the crude ferric phosphate with a hydrochloric acid solution (approximately 0.15 mol / L) with an initial pH of 0.8 at a liquid-to-solid ratio of 1 mL / g in a high-temperature and high-pressure reactor and react at 180°C for 3 h to obtain refined ferric phosphate.

[0081] Step 3: Adjust the phosphorus-aluminum molar ratio in the solution to 2.2 using sodium phosphate, adjust the pH of the precipitated solution A to 2.5 using ammonia, react at 120°C for 2 h, and separate the solid and liquid to obtain industrial-grade aluminum phosphate and precipitated solution B. Step 4: Adjust the pH of the precipitated solution B to 2.0 using dilute sulfuric acid. Extract for 10 min using a mixture of 15 vol.% di(2-ethylhexyl) phosphate (P204) and 85 vol.% sulfonated kerosene as the extractant at an extraction ratio of 0.5 (O:A). After phase separation, obtain the raffinate and the supported organic phase. Mix the supported organic phase with 5 mol / L sulfuric acid solution at a volume ratio of 4, and back-extract for 10 min to obtain a vanadium-rich solution with a vanadium concentration of 18 g / L. Adjust the pH to 2.5, add ammonium chloride to the vanadium-rich solution with an ammonium addition coefficient of 3, precipitate at room temperature for 2 h, and obtain ammonium polyvanadate after solid-liquid separation. Calcine this polyvanadate at 500°C for 1 h to prepare vanadium pentoxide.

[0082] The leaching rates of vanadium, iron, and aluminum, the product recovery rates, and the purity of vanadium pentoxide products in Examples 1-4 and Comparative Examples 1-4 were tested, and the results are shown in Table 2.

[0083] Table 2. Process effects of Examples 1-4 and Comparative Examples 1-4

[0084] As shown in Table 2 and Examples 1-4, the ultrasonic-enhanced leaching technology employed in this invention can achieve simultaneous and efficient leaching of vanadium, iron, and a portion of aluminum without the use of highly corrosive fluorides and oxidants. The leaching rates of vanadium and iron both exceed 90%, while the leaching rate of aluminum exceeds 40%. Furthermore, through a stepwise directional phosphate precipitation technology based on precise control of solution components, iron ions in the leachate are preferentially converted into crude ferric phosphate, and aluminum ions are subsequently recovered as industrial-grade aluminum phosphate. Finally, a simple extraction-vanadium precipitation process is used to achieve effective separation and precipitation of vanadium. This method can ultimately produce battery-grade iron phosphate (HG / T 4701-2021) that meets the requirements of chemical industry standards, industrial aluminum phosphate (T / ZJCX 0002-2022) that meets the requirements of group standards, and vanadium pentoxide (YB / T 5304-2017) that meets the requirements of ferrous metallurgy industry standards. The recovery rates of vanadium and iron exceed 85% and 80%, respectively, and the purity of vanadium pentoxide products can reach up to 99.5% or more.

[0085] Comparing Example 1 and Comparative Example 1, it is evident that, compared to the process of this invention, the traditional mechanical stirring leaching process, without the addition of oxidants or fluorides as leaching aids, exhibits significantly poorer leaching performance, with vanadium leaching rates below 55% and iron leaching rates below 70%. If the leachate is directly subjected to solvent extraction to separate and enrich vanadium, the iron and aluminum contained within cannot be effectively separated, resulting in resource waste. Furthermore, the coexistence of iron and aluminum severely restricts the vanadium extraction efficiency, leading to a final vanadium pentoxide product purity below 99%. This further demonstrates that, compared to traditional processes, this invention can achieve highly efficient leaching and resource utilization of vanadium, iron, and aluminum in vanadium shale, with the vanadium pentoxide product exhibiting higher purity and recovery rate.

[0086] By comparing Example 1 with Comparative Examples 2 and 3, it can be seen that the solution pH value during the iron precipitation process has a significant impact on the technical effect of the present invention. Too low an iron precipitation pH value will significantly reduce the recovery rate of battery-grade iron phosphate products; too high an iron precipitation pH value will lead to a decrease in the recovery rate of vanadium pentoxide and industrial-grade aluminum phosphate, and under the same purification conditions, it will be impossible to obtain battery-grade iron phosphate that meets the standards.

[0087] By comparing Example 1 with Comparative Examples 4 and 5, it can be seen that the phosphorus-iron molar ratio in the iron precipitation solution has a significant impact on the technical effect of the present invention. An excessively low phosphorus-iron molar ratio will reduce the recovery rate of battery-grade iron phosphate; an excessively high phosphorus-iron molar ratio will lead to a decrease in the recovery rates of vanadium pentoxide and industrial-grade aluminum phosphate, and under the same purification conditions, it will be impossible to obtain qualified battery-grade iron phosphate.

[0088] By comparing Example 1 with Comparative Examples 6 and 7, it can be seen that the initial pH value of the acid solution in the crude ferric phosphate purification process has a significant impact on the technical effect of the present invention. Too low a pH value results in a large loss of battery-grade ferric phosphate, while too high a pH value makes it impossible to obtain qualified battery-grade ferric phosphate.

[0089] By comparing Example 1 with Comparative Examples 8 and 9, it can be seen that the pH of the solution in the aluminum precipitation process has a significant impact on the technical effect of the present invention. Too low a pH will reduce the recovery rate of industrial aluminum phosphate products and also reduce the purity of vanadium pentoxide products; too high a pH will significantly reduce the recovery rate of vanadium pentoxide and will not yield qualified industrial-grade aluminum phosphate.

[0090] By comparing Example 1 with Comparative Examples 10 and 11, it can be seen that the phosphorus-aluminum molar ratio in the aluminum precipitation process has a significant impact on the technical effect of the present invention. If the phosphorus-aluminum molar ratio is too low, the recovery rate of industrial-grade aluminum phosphate will be too low, and the purity of vanadium pentoxide products will also be reduced. If the phosphorus-aluminum molar ratio is too high, the recovery rate of vanadium pentoxide will be significantly reduced, and industrial-grade aluminum phosphate that meets the standards cannot be obtained.

[0091] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale, characterized in that, Includes the following steps: Step 1: Crush and grind the vanadium shale ore until the proportion of particles smaller than 200 mesh is not less than 75%, then mix it with sulfuric acid solution, leach it with ultrasonic waves, and separate the solid and liquid to obtain leachate. Step 2: Add hydrogen peroxide to the leachate, and after the oxidation reaction, add phosphate. Then adjust the pH of the solution to 1.0~2.0, and then react at a set temperature. Then perform solid-liquid separation to obtain crude iron phosphate and precipitated liquid A. Then purify the crude iron phosphate to prepare battery-grade iron phosphate. Step 3: Add phosphate to the precipitated solution A, adjust the pH of the mixed solution to 2.4~3.0, then react at a set temperature, and then perform solid-liquid separation to obtain industrial-grade aluminum phosphate and precipitated solution B; Step 4: Adjust the pH value of the precipitated liquid B, then extract and separate to obtain the loaded organic phase. The loaded organic phase is back-extracted to obtain a vanadium-rich solution, which is then treated with ammonium salt precipitation to prepare vanadium pentoxide product.

2. The method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 1, characterized in that, In step one, the chemical composition of the vanadium shale ore, by mass percentage, includes: V2O5 0.5%~1.2%, Al2O3 6%~10%, Fe2O3 2%~5%, SiO2 45%~73%, K2O 1%~3%, MgO 0.5~2%, and CaO 0.3~5%.

3. The method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 2, characterized in that, The concentration of the sulfuric acid solution is 10-30 vol.%, and the liquid-to-solid ratio of the sulfuric acid solution to the vanadium shale ore is 1-5 mL / g. During the ultrasonic leaching process, the ultrasonic power is set to 100~500 W, the leaching temperature is 25~100°C, and the leaching time is 0.5~4 h.

4. The method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 1, characterized in that, In step two, the amount of hydrogen peroxide added is 1 to 4 times the amount of ferrous ions in the leachate; the oxidation reaction time is 10 to 60 minutes, and the oxidation reaction temperature is 25 to 50°C.

5. The method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 4, characterized in that, The phosphate is selected from one or more of sodium phosphate, potassium phosphate, and ammonium phosphate; after the phosphate is added, the phosphorus-iron molar ratio in the solution is 1.0~1.6; The set temperature is 25~120°C, and the reaction time is 0.1~3 h.

6. A method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 4, characterized in that, The purification of crude ferric phosphate specifically includes: The crude ferric phosphate and acid solution are mixed, followed by solid-liquid separation and drying to obtain refined ferric phosphate; wherein the pH value of the acid solution is 0.2~1.2; And / or, the liquid-to-solid ratio of the acid solution to the crude ferric phosphate is 1~10 mL / g; And / or, the temperature for preparing battery-grade iron phosphate is 140~200°C and the time is 2~4 h.

7. The method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 1, characterized in that, In step three, the phosphorus-aluminum molar ratio in the mixed solution is 1.5-2.0; the set temperature is 25-120°C; and the reaction time is 0.1-3 h.

8. The method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 1, characterized in that, In step four, the extractant is selected from di(2-ethylhexyl) phosphate and / or 2-ethylhexylphosphonic acid mono(2-ethylhexyl) ester; During the extraction process, the volume ratio of the organic phase containing the extractant to the precipitated liquid B is 0.25~1, and the extraction time is 5~20 min.

9. A method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 8, characterized in that, The stripping agent is selected from dilute hydrochloric acid and / or dilute sulfuric acid; During the back-extraction process, the volume ratio of the supported organic phase to the aqueous phase containing the back-extraction agent is 1 to 4, and the extraction time is 5 to 20 minutes.

10. A method for resource-based recovery of vanadium, iron, and aluminum from vanadium shale as described in claim 8, characterized in that, The preparation of vanadium pentoxide from the vanadium-rich solution specifically includes: After adjusting the pH of the vanadium-rich solution to 2-3, ammonium salt is added, and after precipitation, solid-liquid separation is performed to obtain ammonium polyvanadate, which is then calcined to obtain vanadium pentoxide product. The ammonium salt is selected from ammonium chloride and / or ammonium sulfate, and the molar ratio of ammonium ions to vanadium ions in the ammonium salt is 2 to 4. The precipitation temperature is 20~30℃, and the precipitation time is 1~2 h; The calcination temperature is 500~600°C, and the calcination time is 30~120 min.

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