A method for extracting vanadium from vanadium slag roasting products and preparing vanadium oxide or vanadyl sulfate

Abstract

The present invention relates to a method for extracting vanadium from a vanadium slag roasting product and preparing vanadium oxide or vanadyl sulfate. The method comprises the following steps: crushing and grinding the vanadium slag roasting product, performing simultaneous reduction leaching in a leaching medium, adjusting the pH of the leachate to 0-5 with a pH regulator, and maintaining a liquid-to-solid ratio of the leaching medium to the roasted product of 3-20:1, thereby obtaining a leachate containing positive divalent vanadium oxide ions. While leaching the vanadium slag roasting product, the present invention adds a reducing agent to directly control the valence of vanadium in the leachate, achieving a stable tetravalent valence, thereby resolving the problem of low vanadium leaching efficiency caused by the easy hydrolysis of pentavalent vanadium ions. Vanadyl hydroxide is then prepared by a selective precipitation method, which is then calcined under different atmospheres to produce VO2 or V2O5, or vanadyl sulfate is produced by in-situ extraction-stripping in the reducing leachate. The method of the present invention has a short process flow, a high vanadium yield, a compact process interface, reduces the three wastes, and is easily industrialized.

Description

Technical Field

[0001] The invention relates to a method for extracting vanadium from vanadium slag roasting products to prepare vanadium oxide or vanadyl sulfate, belonging to the technical field of vanadium chemical technology and resource recycling. Background Art

[0002] With the advancement of science and technology and the development of the times, the energy storage industry has developed rapidly, resulting in an increasing demand for vanadium supply. Vanadium slag is one of the most important raw materials for vanadium extraction in China. The traditional vanadium slag extraction process is generally a roasting and leaching process to obtain V 5+ Due to the hydrolysis of pentavalent vanadium ions within a specific pH range, it is difficult to achieve both purity and leaching efficiency in acidic leaching processes. Industrial production generally adopts a constant pH leaching process of 2.5-3.5, with a vanadium leaching efficiency of approximately 85%. At the same time, the traditional process for preparing vanadium oxides is a vanadium-containing leachate-acidic ammonium salt vanadium precipitation process, which produces a large amount of ammonia nitrogen wastewater and is not in line with the national green development strategy. Currently, there are no reports of an integrated process for directly reducing and leaching divalent vanadium oxide ions from the vanadium slag roasting product, followed by precipitation to prepare vanadium oxides or extraction to prepare vanadium sulfate. Therefore, it is necessary to develop a green, low-cost vanadium extraction process to avoid the above problems.

[0003] Patent 202410421010.6 discloses a method for deep vanadium extraction from vanadium slag, achieving a vanadium leaching rate of up to 94.1% and a vanadium yield of up to 92.52%. However, the process is complex, requiring the vanadium slag and calcium salt to undergo a primary oxidation roasting, a primary acid leaching, a secondary oxidation roasting, a secondary acid leaching, solution precipitation, and further vanadium extraction from the vanadium-containing leachate to obtain V2O5.

[0004] Patent 202310297718.0 discloses a method for preparing vanadium pentoxide or vanadium dioxide by hydrolysis precipitation of vanadium, wherein the tetravalent vanadium (VO 2+ ) to obtain a vanadium-loaded organic phase, which is then stripped with hydrochloric acid and precipitated with alkaline vanadium precipitation to obtain vanadium pentoxide or vanadium dioxide. The raw material vanadium-containing liquid (pentavalent vanadium liquid) is obtained by leaching and then reducing tetravalent vanadium (VO 2+ ), in the process from vanadium slag roasting product to vanadium-containing liquid, pentavalent vanadium liquid inevitably undergoes hydrolysis reaction, and the problem of easy hydrolysis of pentavalent vanadium ions in the vanadium extraction process from vanadium slag is not solved.

[0005] Patent application 201910791690.X discloses a method for preparing solid vanadyl sulfate. This invention uses a vanadium-containing solution (pentavalent vanadium solution) obtained by roasting and leaching vanadium slag from a steelmaking converter. Through targeted iron removal, vanadium valence is converted, followed by solvent extraction and purification. A high-purity vanadyl sulfate solution is obtained through mixed acid stripping, and solid vanadyl sulfate is obtained by direct negative pressure distillation and crystallization. However, this process also involves leaching vanadium from the vanadium slag before reduction. The pentavalent vanadium solution inevitably undergoes hydrolysis, resulting in a relatively low vanadium yield and a lengthy process.

[0006] All of the above methods can produce vanadium dioxide, vanadium pentoxide or vanadyl sulfate, but their raw materials are all vanadium-containing leachate, in which the vanadium is mainly pentavalent, and then a reduction process is used to obtain divalent vanadium oxide ions. From the source, these processes have the common feature of long processes, and the acquisition of vanadium-containing raw materials is also the product of traditional processes. More importantly, none of the above existing methods can solve the problem of easy hydrolysis of pentavalent vanadium ions in the process of extracting vanadium from vanadium slag. Due to the hydrolysis of pentavalent vanadium ions within a specific pH range, the leaching efficiency of vanadium is low, and the problem of vanadium extraction cannot be solved from the source. This is also a bottleneck problem that urgently needs to be solved in this field. Summary of the Invention

[0007] (1) Technical issues to be resolved

[0008] In order to solve the above problems in the prior art, the present invention provides a method for extracting vanadium from a vanadium slag roasting product and preparing vanadium oxide or vanadyl sulfate.

[0009] (2) Technical solution

[0010] In order to achieve the above objectives, the main technical solutions adopted by the present invention include:

[0011] The first aspect of the present invention provides a method for extracting vanadium from a vanadium slag roasting product, comprising the following steps: crushing and grinding the vanadium slag roasting product, performing simultaneous reduction leaching in a leaching medium, adjusting the pH of the leaching solution to a constant pH value between 0 and 5 with a pH regulator, and the leaching medium and the roasting product are in a liquid-solid ratio of 3 to 20:1 to obtain a vanadium ion containing positive divalent vanadium oxide (VO 2+ ) of the leachate.

[0012] Furthermore, the leaching medium is a reducing agent and water, wherein the molar ratio of the reducing agent to the vanadium in the roasted clinker is 1 to 2:1, and the rest is water; the reducing agent is at least one or more of sulfurous acid and sulfites such as sodium sulfite, sodium bisulfite, sodium metabisulfite, sulfur dioxide, hydrogen sulfide and its salts, organic reducing acids (oxalic acid, citric acid, ascorbic acid, glucose, histidine) and their salts, and hydrazine hydrate or any other reducing agents, and the temperature of the synchronous reduction leaching process is 10°C to 100°C, and the time is 10min to 120min.

[0013] Furthermore, the vanadium slag can be one or more of ordinary vanadium slag (chromium content less than 5%), vanadium-chromium slag (chromium content greater than 5% and less than 15%), or high-calcium and high-phosphorus vanadium slag (CaO / V2O5≥0.4, P≥0.1%).

[0014] Furthermore, the vanadium slag roasting product can be one or more of a blank roasting product, a sodium roasting product, a calcium roasting product, a manganese roasting product, a magnesium roasting product, a calcium-manganese composite roasting product, a calcium-magnesium composite roasting product, and a roasting product with various additives; the roasting product is ground to: 38μm < particle size < 74μm.

[0015] Furthermore, the pH is preferably 0-2, and the pH adjuster is at least one of sulfuric acid, hydrochloric acid, and nitric acid. In addition, an organic reducing acid can also be used to adjust the pH, such as one or more of citric acid, oxalic acid, and ascorbic acid.

[0016] A second aspect of the present invention provides a method for extracting vanadium from a vanadium slag roasting product and preparing vanadium oxide, which comprises the following steps:

[0017] S1. Obtaining a leachate containing positive divalent vanadium oxide ions by the vanadium extraction method of claim 1;

[0018] S2, adjusting the pH of the leachate obtained in step S1 to 4-8 (within this pH range to control the region of vanadium precipitation) with an alkaline substance, performing a precipitation reaction at a reaction temperature of 10° C. to 100° C. for 10 min to 120 min, and cooling to room temperature after the reaction is completed;

[0019] S3, separating the product obtained in step S2 by solid-liquid separation, repeatedly washing the precipitate and drying it to obtain vanadium oxyhydroxide;

[0020] S4. calcining the vanadium oxyhydroxide obtained in step S3 in an inert atmosphere at 300° C. to 500° C. for 2 to 3 hours to obtain vanadium dioxide, or heating in an air atmosphere at 500° C. to 670° C. to obtain vanadium pentoxide.

[0021] Furthermore, in step S2, the alkaline substance is one or more of sodium hydroxide, saturated lime water, sodium carbonate, sodium carbonate salt, bicarbonate or ammonia water; in step S3, the solid-liquid separation is filtration or centrifugation or filter pressing, and the repeated rinsing is repeated rinsing with water and anhydrous ethanol several times; the drying condition is vacuum drying at 30°C to 70°C for 24 to 48 hours.

[0022] Furthermore, in step S4, the calcination atmosphere for preparing vanadium dioxide is an inert atmosphere, preferably argon or nitrogen, and the calcination temperature is 300°C to 500°C; the temperature for preparing vanadium pentaoxide is 500°C to 670°C, and the air atmosphere is one or both of air and oxygen.

[0023] A third aspect of the present invention provides a method for extracting vanadium from a vanadium slag roasting product and preparing vanadyl sulfate, comprising the following steps:

[0024] S1. Obtaining a leachate containing positive divalent vanadium oxide ions by the vanadium extraction method of claim 1;

[0025] S2. The leachate obtained in step S1 is extracted with an extractant to obtain a first organic phase loaded with vanadium. The first organic phase loaded with vanadium is washed with water and acid to obtain a second organic phase loaded with vanadium. The obtained second organic phase is stripped with sulfuric acid to obtain a vanadyl sulfate solution.

[0026] Furthermore, the extractant is P5O7 (2-ethylhexylphosphonic acid mono-2-ethylhexyl ester), P2O4 di(2-ethylhexyl) phosphate or bis(2,4,4-trimethylpentyl)phosphonic acid (cynaex272); the extraction stage is one or more stages; the back extraction stage is single or multiple stages, and the extraction method is multi-stage countercurrent extraction or staggered extraction.

[0027] Compared with the existing technology, the beneficial effects of the present invention are:

[0028] The present invention provides a method for extracting vanadium from vanadium slag roasting products to directly prepare vanadium oxide or vanadyl sulfate, wherein the vanadium slag roasting products are simultaneously subjected to reduction leaching to obtain a vanadium-containing leaching solution (VO 2+), by selectively precipitating vanadium, recovering vanadium, preparing vanadium oxide, or extracting vanadyl sulfate. The present invention creatively proposes adding a reducing agent to directly control the valence of vanadium in the leachate while leaching the vanadium slag roasting product, obtaining a stable tetravalent valence. This solves the problem of pentavalent vanadium ions being easily hydrolyzed, which leads to low vanadium leaching efficiency. Vanadyl hydroxide is then prepared by selective precipitation, which is then calcined under different atmospheres to produce VO2 or V2O5, or vanadyl sulfate is produced by in-situ extraction and stripping in the reducing leachate. This process differs from the traditional process of extracting vanadium from vanadium slag to produce V2O5 and the process of reducing and extracting vanadyl sulfate from the leachate of vanadium slag roasting. The entire process achieves efficient vanadium recovery and valence transformation during leaching. Finally, through different vanadium recovery methods, the vanadium product is diversified and highly valued. The method of the present invention has a vanadium conversion rate of 95.24% to 98.46% and a vanadium recovery rate of 97.65% to 99.3%, which are much higher than those of the existing vanadium extraction method. It effectively solves the problem of easy hydrolysis of pentavalent vanadium ions in the vanadium slag vanadium extraction process and improves the vanadium conversion rate and recovery rate.

[0029] The proposed process of vanadium slag roasting product - simultaneous reduction leaching - selective vanadium precipitation / extraction - calcination / strip extraction achieves efficient and high-value utilization of vanadium slag and pollution-free vanadium metallurgy. This invention achieves a short process flow, low cost, easy industrialization, high-value, and green vanadium extraction process from vanadium slag. DETAILED DESCRIPTION

[0030] In order to better explain the present invention and facilitate understanding, the present invention is described in detail below through specific implementation methods.

[0031] Example 1

[0032] (1) The vanadium slag calcified roasted product was crushed and ground to about 200 mesh, i.e., to a particle size of less than 74 μm. Water was added to the roasted product at a liquid-to-solid ratio of 10 mL / g for leaching. During the leaching process, sulfuric acid was used to adjust the pH. The reducing agent was Na2SO3 to keep the pH of the leachate at 2. The leaching temperature was 60°C and the leaching time was 60 min. After the leaching was completed, the solid and liquid were separated by filtration to obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate was blue, and the vanadium leaching rate was calculated to be 98.15%;

[0033] (2) Take the vanadium leaching solution, adjust the pH of the system to 4 with NaOH, and maintain it for 10 minutes;

[0034] (3) The reaction product was separated by filtration, washed repeatedly with water and anhydrous ethanol several times to obtain a black solid, and dried in an oven at 60°C for 24 h to obtain VO(OH)2;

[0035] (4) The obtained VO(OH)2 was calcined in a tube furnace at 400°C for 2 h under nitrogen protection to obtain VO2 with a purity of 99.14%, a vanadium precipitation rate of 98.33%, and a total vanadium recovery rate of 96.51%.

[0036] Example 2

[0037] (1) The vanadium slag calcified roasted product was crushed and ground to a particle size of less than 74 μm, and water was added to the roasted product at a liquid-to-solid ratio of 3 mL / g for leaching. During the leaching process, sulfuric acid was used to adjust the pH of the leachate to maintain the pH at 1.8. The leaching temperature was 90°C, the leaching time was 60 min, and the reducing agent was ascorbic acid. After the leaching was completed, the solid and liquid were separated by filtration to obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate was blue, and the calculated vanadium leaching rate was 98.46%;

[0038] (2) Take 50 mL of vanadium-containing leachate, adjust its pH to 4.5 with ammonia water, and carry out precipitation reaction at 30°C for 30 minutes;

[0039] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol several times, and drying in a vacuum oven at 60°C for 24 hours to obtain vanadium oxyhydroxide;

[0040] (4) The obtained VO(OH)2 was calcined in a tubular furnace at 400°C for 2 h under nitrogen protection to obtain VO2 with a purity of 98.14% and a vanadium precipitation rate of 98.15%.

[0041] Example 3

[0042] (1) The vanadium slag calcium manganese roasted product was crushed and ground to a particle size of less than 74 μm, and water was added to the roasted product at a liquid-to-solid ratio of 3 mL / g for leaching. During the leaching process, sulfuric acid was used to adjust the pH of the leachate to keep the pH at 0. The leaching temperature was 60°C, the leaching time was 60 min, and the reducing agent was ascorbic acid. After the leaching was completed, the solid and liquid were separated by filtration to obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate was blue, and the calculated vanadium conversion rate was 95.46%;

[0043] (2) Take 100 mL of vanadium-containing leachate, adjust its pH to 5.5 with ammonia water, and carry out precipitation reaction at 30°C for 10 minutes;

[0044] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol several times, and drying in a vacuum oven at 60°C for 24 hours to obtain vanadium oxyhydroxide;

[0045] (4) The obtained vanadium oxyhydroxide was heated in a muffle furnace at 600° C. for 180 min and oxidized and calcined to obtain V 2 O 5 with a purity of 99.11% and a vanadium precipitation rate of 98.14%.

[0046] Example 4

[0047] (1) The sodium-treated roasted product of vanadium slag was crushed and ground to a particle size of less than 74 μm. Water was added to the roasted product at a liquid-to-solid ratio of 10 mL / g for leaching. During the leaching process, sulfuric acid was used to adjust the pH of the leachate to maintain the pH at 1.5. The leaching temperature was 90° C., the leaching time was 60 min, and the reducing agent was sodium sulfite. After the leaching was completed, the solid and liquid were separated by filtration to obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate was blue, and the calculated vanadium conversion rate was 97.24%;

[0048] (2) taking 100 mL of vanadium-containing leachate, removing chromium from the leachate, adjusting its pH to 6 with ammonia water, and carrying out precipitation reaction at 30°C for 30 minutes;

[0049] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol several times, and drying in a vacuum oven at 60°C for 24 hours to obtain vanadium oxyhydroxide;

[0050] (4) The obtained vanadium oxyhydroxide was heated in a muffle furnace at 600° C. for 240 min to obtain V2O5 with a purity of 99.13% and a vanadium precipitation rate of 98.14%.

[0051] Example 5

[0052] (1) The vanadium slag calcium-magnesium composite roasted product was crushed and ground to a particle size of less than 74 μm, and water was added to the roasted product at a liquid-to-solid ratio of 10 mL / g for leaching. During the leaching process, sulfuric acid was used to adjust the pH of the leachate to maintain the pH at 1.5. The leaching temperature was 90°C, the leaching time was 60 min, and the reducing agent was sodium sulfite. After the leaching was completed, the solid and liquid were separated by filtration to obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate was blue, and the calculated vanadium leaching rate was 95.24%;

[0053] (2) Take 100 mL of vanadium-containing leachate, adjust its pH to 6 with ammonia water, and carry out precipitation reaction at 30°C for 30 minutes;

[0054] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol several times, and drying in a vacuum oven at 60°C for 24 hours to obtain vanadium oxyhydroxide;

[0055] (4) The obtained vanadium oxyhydroxide was heated in a muffle furnace at 600° C. for 240 min to obtain V2O5 with a purity of 99.15% and a vanadium recovery rate of 98.17%.

[0056] Example 6

[0057] The difference between this embodiment and embodiment 2 is that 20 mL of vanadium leaching solution and 25 mL of organic extractant (20% P2O4 + 5% TBP + 75% sulfonated kerosene) were mixed in a beaker for 10 minutes, and then the vanadium-containing organic phase was separated and washed with 25°C water for 10 minutes, with an O / A ratio of 1:5. The washed organic phase was further washed with 0.02 mol / L sulfuric acid for 2 minutes, with an O / A ratio of 1:1. The organic phase after the acid wash was stripped with sulfuric acid at a sulfuric acid concentration of 2.5 mol / L, an O / A ratio of 5:1, and a stripping time of 20 minutes. After three-stage extraction and stripping, the vanadium recovery rate was 99.1%. The vanadium concentration in the vanadyl sulfate solution obtained by stripping was tested. The resulting vanadyl sulfate was evaporated and concentrated to 80g / L, with a measured impurity content of 35.9mg / L (standard for a composite high-purity vanadyl sulfate solution: impurity ions less than 40mg / L, vanadium ion concentration greater than 76g / L). Compared to traditional vanadyl sulfate preparation from vanadium slag, the overall vanadium recovery rate exceeded 98%, far exceeding the current 90%.

[0058] Example 7

[0059] The difference between this embodiment and embodiment 3 is that 20 mL of vanadium leaching solution and 20 mL of organic extractant (25% P5O7 + 5% TBP + 70% sulfonated kerosene) are taken and mixed evenly in a beaker for 10 minutes. The vanadium-containing organic phase is then separated and washed with 25°C water for 10 minutes, with a phase ratio (O / A) of 1:5. The washed organic phase is further washed with 0.02 mol / L sulfuric acid for 2 minutes, with a phase ratio (O / A) of 1:1. The organic phase after the acid wash is subjected to sulfuric acid stripping with a sulfuric acid concentration of 2.5 mol / L, a phase ratio (O / A) of 7:1, and a stripping time of 20 minutes. After three-stage extraction and stripping, the vanadium recovery rate is 99.3%. The vanadium concentration in the vanadium sulfate solution obtained by stripping is tested. The obtained vanadium sulfate is evaporated and concentrated to 82 g / L, and the impurity content is measured to be 34.3 mg / L.

[0060] In order to demonstrate that the method of the present invention has a high conversion rate for vanadium, the present invention directly roasts the vanadium slag without additives, and conducts a direct acid leaching experiment on a blank roasted product, as shown in Example 8.

[0061] Example 8

[0062] (1) The vanadium slag is directly roasted without additives to obtain a blank roasted clinker. The blank roasted product of the vanadium slag is crushed and ground to a particle size of less than 74 μm. Water is added to the roasted product at a liquid-to-solid ratio of 5 mL / g for leaching. During the leaching process, sulfuric acid is used to adjust the pH of the leachate to maintain the pH at 1.5. The leaching temperature is 90°C, the leaching time is 60 min, and the reducing agent is sulfurous acid. After the leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate is blue, and the calculated vanadium leaching rate is 85.47%;

[0063] (2) Take 50 mL of vanadium-containing leachate, adjust the pH of the system to 5.0 with ammonia water, and carry out vanadium precipitation reaction at 30°C for 10 minutes;

[0064] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol several times, and drying in an oven at 60° C. for 24 h to obtain vanadium oxyhydroxide;

[0065] (4) The obtained vanadium oxyhydroxide was heated in a muffle furnace at 600° C. for 240 min to obtain V 2 O 5 with a purity of 99.13% and a vanadium precipitation rate of 97.65%.

[0066] Control experiment: The difference from Example 8 is that no reducing agent is added during leaching. In this case, when the leaching pH is 1.5, the vanadium leaching rate is 10.31%; when the leaching pH is 2.8 (conventional leaching pH), the vanadium leaching rate is 50.24%.

[0067] In a direct acid leaching experiment on a blank roasted product, the vanadium leaching rate of the present invention method compared to a no-additive roasted product was examined. The present invention method (Example 8) achieved a leaching rate of 85.47%, which was approximately 75% higher than the 10.31% (pH 1.5) and approximately 35% higher than the conventional leaching rate of 50.24% (pH 2.8). This demonstrates that the present invention method has a high conversion rate for vanadium in the blank roasted product of vanadium slag.

[0068] Comparative Example 1

[0069] The calcified vanadium slag roasted product was crushed and ground to approximately 200 mesh, i.e., to a particle size of less than 74 μm. Water was then added at a liquid-to-solid ratio of 10 mL / g of the roasted product for leaching. Sulfuric acid was used to adjust the pH of the leachate, maintaining it between 2.5 and 2.8. The leaching temperature was 60°C, and the leaching time was 60 minutes. After leaching, the solid and liquid were filtered to separate the yellow vanadium-containing leachate, yielding a calculated vanadium recovery of 87.15%. The resulting vanadium-containing leachate was then subjected to ammoniacal salt precipitation and calcination to produce vanadium dioxide (V2O5), with a total vanadium recovery of 85.07%. Compared to Example 1, the reduction leaching process achieved an approximately 11% higher vanadium recovery than the conventional non-reduction leaching process. The total vanadium recovery was increased by approximately 10%. While the conventional process yielded V2O5, the reduction leaching process of the present invention produced VO2.

[0070] Comparative Example 2

[0071] The only difference between Comparative Example 2 and Example 1 is that in step (1), the liquid-to-solid ratio of the leaching medium to the calcined product is 2:1. The resulting vanadium leaching rate is 91.13%, which is lower than that of Example 1. When the liquid-to-solid ratio of the leaching medium to the calcined product is 20:1, the vanadium leaching efficiency is 98.21%, which is slightly higher than that of Example 1. However, further increasing the liquid-to-solid ratio does not significantly improve the vanadium leaching efficiency. In terms of economy and cost, the present application preferably uses a liquid-to-solid ratio of the leaching medium to the calcined product of less than 3 to 20:1.

[0072] Comparative Example 3

[0073] The only difference between Comparative Example 3 and Example 1 is that the temperature of the simultaneous reduction leaching process in step (1) is 5°C. The vanadium leaching efficiency is 90.21%, which is approximately 8% lower than that of Example 1. Leaching temperatures above 100°C are difficult to implement industrially, and further increasing the temperature does not yield significant benefits. Therefore, the preferred temperature for the simultaneous reduction leaching process in this application is 10°C to 100°C.

[0074] In summary, the method of the present invention achieves a vanadium conversion rate of 95.24% to 98.46% and a vanadium recovery rate of 97.65% to 99.3%, significantly exceeding existing vanadium extraction methods and Comparative Examples 1 to 3. This method effectively addresses the issue of pentavalent vanadium ions being easily hydrolyzed during vanadium slag extraction, improving both vanadium conversion and recovery rates. Furthermore, the VO2 obtained by the present invention can be used in fields such as smart windows, infrared stealth, and supercapacitors. The obtained vanadyl sulfate can be used in energy storage, and vanadium pentoxide can be used in the steel industry. The vanadium products obtained by the present invention have multiple applications and high value-added benefits.

[0075] The preferred embodiments of the present invention have been described in detail above, but the present invention is not limited thereto. Within the technical scope of the present invention, the present invention may include various simple variations of the technical solution of the present invention, including combining the various technical features in any other appropriate manner. These simple variations and combinations should also be considered as disclosed in the present invention and fall within the scope of protection of the present invention.

Claims

1. A method for extracting vanadium from a vanadium slag roasting product and preparing vanadium oxide or vanadyl sulfate, characterized in that: The steps include: S1. Crushing and grinding the vanadium slag roasting product, and performing simultaneous reduction leaching in a leaching medium, wherein the leaching medium is a reducing agent and water, and the pH of the leachate is adjusted to a constant pH value between 0 and 5 by a pH regulator. The liquid-solid ratio of the leaching medium to the roasting product is 3 to 20 mL: 1 g, and the molar ratio of the reducing agent to the vanadium in the roasted clinker is 1 to 2: 1; obtaining a leachate containing positive divalent vanadium oxide ions; the vanadium slag roasting product is one or more of a blank roasting product, a sodium roasting product, a calcium roasting product, a manganese roasting product, a magnesium roasting product, a calcium-manganese composite roasting product, and a calcium-magnesium composite roasting product; and the vanadium slag is one or more of ordinary vanadium slag, vanadium-chromium slag, or high-calcium and high-phosphorus vanadium slag; S2. The leachate obtained in step S1 is adjusted to a pH of 4 to 8 with an alkaline substance, and a precipitation reaction is performed at a temperature of 10° C. to 100° C. for 10 min to 120 min. After the reaction is completed, the solution is cooled to room temperature; S3, separating the product obtained in step S2 by solid-liquid separation, repeatedly washing the precipitate and drying it to obtain vanadium oxyhydroxide; S4, calcining the vanadium oxyhydroxide obtained in step S3 in an inert atmosphere at 300° C. to 500° C. for 2 to 3 hours to obtain vanadium dioxide or heating it in an air atmosphere at 500° C. to 670° C. to obtain vanadium pentoxide; or, The leachate obtained in step S1 is extracted with an extractant to obtain a first organic phase loaded with vanadium. The first organic phase loaded with vanadium is washed with water and acid to obtain a second organic phase loaded with vanadium. The obtained second organic phase is stripped with sulfuric acid to obtain a vanadyl sulfate solution.

2. The method according to claim 1, characterized in that The temperature of the synchronous reduction leaching is 10°C to 100°C, and the time is 10min to 120min.

3. The method according to claim 1, characterized in that The calcined product was ground to a particle size of <74 μm.

4. The method according to claim 1, wherein The pH value of the leaching solution is 0-2, and the pH regulator is one or more of sulfuric acid, hydrochloric acid, nitric acid, citric acid, oxalic acid and ascorbic acid.

5. The method according to claim 1, characterized in that In step S2, the alkaline substance is one or more of sodium hydroxide, saturated lime water, sodium carbonate, sodium carbonate salt, bicarbonate or ammonia water; in step S3, the solid-liquid separation is filtration or centrifugation or filter pressing, and the repeated rinsing is repeated rinsing with water and anhydrous ethanol for several times; the drying condition is vacuum drying at 30°C to 70°C for 24 to 48 hours.

6. The method according to claim 1, wherein In step S4, the calcination atmosphere for preparing vanadium dioxide is an inert atmosphere of argon or nitrogen, and the calcination temperature is 300°C to 500°C; the temperature for preparing vanadium pentaoxide is 500°C to 670°C, and the atmosphere is one or both of air and oxygen.

7. The method according to claim 1, characterized in that The extractant is P507, P204 diphosphate or bis(2,4,4-trimethylpentyl)phosphonic acid; the extraction stage is one-stage or multi-stage extraction; the back extraction stage is single-stage or multi-stage, and the extraction method is multi-stage countercurrent extraction or staggered extraction.