Method for extracting vanadium from vanadium slag roasting product and preparing vanadium oxide or vanadyl sulfate

Abstract

The invention relates to a method for extracting vanadium from a vanadium slag roasting product and preparing vanadium oxide or vanadyl sulfate, which comprises the following steps of: crushing the vanadium slag roasting product, grinding, carrying out synchronous reduction leaching in a leaching medium, regulating the pH value of a leaching solution to 0-5 by using a pH regulator, and carrying out solid-liquid separation on the leaching medium and the roasting product according to the liquid-solid ratio of (3-20): 1, so as to obtain the vanadium oxide or vanadyl sulfate. And leaching liquid containing positive divalent vanadium oxygen ions is obtained. According to the method, when the vanadium slag roasting product is leached, the reducing agent is added to directly control the valence state of vanadium in the leaching solution, tetravalent vanadium with the stable valence state is obtained, the problem that the vanadium leaching efficiency is low due to the fact that pentavalent vanadium ions are prone to hydrolysis is solved, vanadium oxyhydroxide is prepared in combination with a selective precipitation method, VO2 or V2O5 is prepared through calcination under different atmospheres, and the method is suitable for industrial production. Or in-situ extraction-reverse extraction is carried out in the reduction leachate to prepare vanadyl sulfate. The method is short in process, high in vanadium yield, compact in process linkage, less in three wastes and easy to industrialize.

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 vanadium extraction process is generally a roasting and leaching process to obtain V 5+ Ions, due to the hydrolysis of pentavalent vanadium ions in a specific pH range, it is difficult to achieve both purity and leaching efficiency in the acidic leaching process. Industrial production generally adopts leaching at a constant pH of 2.5-3.5, and the vanadium leaching efficiency is about 85%. At the same time, the traditional process for preparing vanadium oxides is a vanadium-containing leaching solution-acidic ammonium salt vanadium precipitation process, which produces a large amount of ammonia nitrogen wastewater, which is not in line with the national green development strategy. At present, there is no report on the overall process of directly reducing and leaching the vanadium slag roasting product to obtain positive divalent vanadium oxide ions, and then precipitating to prepare vanadium oxides or extracting to prepare vanadium sulfate. Therefore, it is necessary to develop a green and low-cost vanadium extraction process to avoid the above problems.

[0003] Patent 202410421010.6 discloses a method for deep vanadium extraction from vanadium slag, with a vanadium leaching rate of up to 94.1% and a vanadium recovery rate of up to 92.52%. However, the process is complicated, and the vanadium slag and calcium salt are sequentially subjected to a process of primary oxidation roasting, primary acid leaching, secondary oxidation roasting, secondary acid leaching, solution precipitation of vanadium, and vanadium extraction from the vanadium-containing leaching solution to obtain V2O5.

[0004] Patent 202310297718.0 discloses a method for preparing vanadium pentoxide or vanadium dioxide by hydrolyzing vanadium precipitation, wherein the tetravalent vanadium (VO 2+ ) to obtain an organic phase loaded with vanadium, which is then stripped with hydrochloric acid and precipitated with alkaline vanadium, and then washed and calcined 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 the vanadium slag roasting product to the vanadium-containing liquid, the 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] The patent application with the patent number 201910791690.X discloses a method for preparing solid vanadyl sulfate. The invention is based on the vanadium solution (pentavalent vanadium liquid) obtained after roasting and leaching the vanadium slag from the steelmaking converter. Through targeted iron removal, vanadium valence is transferred, and then impurities are removed and purified by solvent extraction. A high-purity vanadyl sulfate solution is obtained by mixed acid stripping, and solid vanadyl sulfate is obtained by direct negative pressure distillation and crystallization. However, this process also first leaches vanadium from the vanadium slag and then reduces it. The pentavalent vanadium liquid inevitably undergoes hydrolysis reaction, the vanadium yield is relatively low, and the process flow is long.

[0006] The above methods can all obtain vanadium dioxide, vanadium pentoxide or vanadium oxysulfate, but their raw materials are all vanadium-containing leaching solutions, 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, the above existing methods cannot 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 needs to be solved urgently 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] (II) Technical solution

[0010] In order to achieve the above object, 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, which comprises the following steps: crushing the vanadium slag roasting product and then grinding it, performing synchronous 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 ions (VO 2+ )'s 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 pyrosulfite, 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 agent, 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 calcined product can be one or more of a blank calcined product, a sodium calcined product, a calcium calcined product, a manganese calcined product, a magnesium calcined product, a calcium-manganese composite calcined product, a calcium-magnesium composite calcined product, and a calcined product with various additives therein; the calcined product can be 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] The 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 leaching solution containing positive divalent vanadium oxide ions by the vanadium extraction method of claim 1;

[0018] S2, adjusting the pH of the leaching solution 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, the reaction temperature is 10°C-100°C, the reaction time is 10min-120min, and the reaction is cooled to room temperature after completion;

[0019] S3, separating the product obtained in step S2 from solid and liquid, and repeatedly washing and precipitating it, and then drying it to obtain vanadium hydroxide;

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

[0024] S1. Obtaining a leaching solution 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, and 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-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.

[0027] Compared with the prior art, the present invention has the following beneficial effects:

[0028] The present invention provides a method for extracting vanadium from a vanadium slag roasting product to directly prepare vanadium oxide or vanadyl sulfate, wherein the vanadium slag roasting product is simultaneously reduced and leached to obtain a vanadium-containing leaching solution (VO 2+), by selectively precipitating vanadium, recovering vanadium, preparing vanadium oxide or extracting vanadium oxysulfate. The present invention creatively proposes that while leaching the vanadium slag roasting product, a reducing agent is added to directly control the valence state of vanadium in the leachate to obtain a tetravalent vanadium with a stable valence state, thereby solving the problem of easy hydrolysis of pentavalent vanadium ions, which leads to the problem of low leaching efficiency of vanadium, and then combining the selective precipitation method to prepare vanadium oxyhydroxide, and then calcining it under different atmospheres to prepare VO2 or V2O5, or extracting and stripping in situ in the reducing leachate to obtain vanadium oxysulfate. This process is different from the traditional process of extracting vanadium from vanadium slag to prepare V2O5 and the process of reducing and extracting vanadium oxysulfate from the leachate of vanadium slag roasting. The whole process realizes efficient recovery of vanadium and valence state transformation during leaching, and finally makes the vanadium product multi-path and high-value utilization through different ways of recovering vanadium. The vanadium conversion rate of the method of the invention is 95.24% to 98.46%, and the vanadium recovery rate is 97.65% to 99.3%, which are much higher than the existing vanadium extraction method, effectively solve the problem of easy hydrolysis of pentavalent vanadium ions in the vanadium slag vanadium extraction process, and improve the vanadium conversion rate and recovery rate.

[0029] The vanadium slag roasting product-synchronous reduction leaching-selective vanadium precipitation / extraction-calcination / stripping process proposed in the present invention realizes efficient and high-value utilization of vanadium slag and pollution-free metallurgy of vanadium. The present invention realizes the short process, low cost, easy industrialization, high value and greenness of the vanadium slag vanadium extraction process. 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) crushing the vanadium slag calcification roasting product and grinding it to about 200 mesh, that is, to a particle size of less than 74 μm, adding water with a liquid-to-solid ratio of water to the roasting product of 10 mL / g for leaching, using sulfuric acid to adjust the pH, using Na2SO3 as a reducing agent to keep the pH of the leachate at 2, the leaching temperature is 60°C, the leaching time is 60 min, filtering to separate the solid and liquid after the leaching is completed, and obtaining a vanadium-containing leachate, which is blue at this time, and the vanadium leaching rate is calculated to be 98.15%;

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

[0034] (3) The reaction product was separated by filtration, and the solid was repeatedly washed with water and anhydrous ethanol for several times to obtain a black solid, which was then 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 tubular furnace at 400°C for 2 hours 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) crushing the vanadium slag calcification roasting product and grinding it to a particle size of less than 74 μm, adding water at a liquid-to-solid ratio of water to the roasting product of 3 mL / g for leaching, using sulfuric acid to adjust the pH during the process so that the pH of the leachate is always maintained at 1.8, the leaching temperature is 90° C., the leaching time is 60 min, and the reducing agent is ascorbic acid. After the leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate, which is blue at this time, and the vanadium leaching rate is calculated to be 98.46%;

[0038] (2) taking 50 mL of vanadium-containing leaching solution, adjusting its pH to 4.5 with ammonia water, and performing precipitation reaction at 30° C. for 30 min;

[0039] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol for several times, and drying in a vacuum oven at 60° C. for 24 h 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) crushing the vanadium slag calcium manganese roasting product and grinding it to a particle size of less than 74 μm, adding water at a liquid-to-solid ratio of water to the roasting product of 3 mL / g for leaching, using sulfuric acid to adjust the pH so that the pH of the leachate is always maintained at 0, the leaching temperature is 60° C., the leaching time is 60 min, and the reducing agent is ascorbic acid. After the leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate, which is blue at this time, and the conversion rate of vanadium is calculated to be 95.46%;

[0043] (2) taking 100 mL of vanadium-containing leaching solution, adjusting its pH to 5.5 with aqueous ammonia, and performing precipitation reaction at 30° C. for 10 min;

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

[0045] (4) The obtained vanadium hydroxide was heated in a muffle furnace at 600° C. for 180 min, and oxidatively calcined to obtain V2O5 with a purity of 99.11% and a vanadium precipitation rate of 98.14%.

[0046] Example 4

[0047] (1) crushing the sodium-treated roasted product of vanadium slag and grinding it to a particle size of less than 74 μm, adding water at a liquid-to-solid ratio of water to the roasted product of 10 mL / g for leaching, using sulfuric acid to adjust the pH of the leaching solution to always maintain the pH at 1.5, the leaching temperature is 90° C., the leaching time is 60 min, and the reducing agent is sodium sulfite. After the leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leaching solution, which is blue at this time, and the conversion rate of vanadium is calculated to be 97.24%;

[0048] (2) taking 100 mL of vanadium-containing leaching solution, removing chromium from the leaching solution, adjusting the pH value to 6 with ammonia water, and performing a precipitation reaction at 30° C. for 30 min;

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

[0050] (4) The obtained vanadium hydroxide 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) crushing the vanadium slag calcium-magnesium composite roasting product and grinding it to a particle size of less than 74 μm, adding water at a liquid-solid ratio of water to the roasting product of 10 mL / g for leaching, using sulfuric acid to adjust the pH during the process so that the pH of the leachate is always maintained at 1.5, the leaching temperature is 90° C., the leaching time is 60 min, and the reducing agent is sodium sulfite. After the leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate, which is blue at this time, and the vanadium leaching rate is calculated to be 95.24%;

[0053] (2) taking 100 mL of vanadium-containing leaching solution, adjusting its pH to 6 with ammonia water, and performing precipitation reaction at 30° C. for 30 min;

[0054] (3) separating the reaction product by filtration, repeatedly rinsing with water and anhydrous ethanol for several times, and drying in a vacuum oven at 60° C. for 24 h 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 implementation case and implementation case 2 is that 20 mL of vanadium leaching solution and 25 mL of organic extractant (20% P2O4 + 5% TBP + 75% sulfonated kerosene) are taken and mixed evenly in a beaker for 10 minutes, and then the vanadium-containing organic phase is separated and washed with 25°C water for 10 minutes, with a phase ratio (O / A) of 1:5, and the organic phase after water washing 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 washing is subjected to sulfuric acid stripping, with a sulfuric acid concentration of 2.5 mol / L, a phase ratio (O / A) of 5:1, and a stripping time of 20 minutes. After three-stage extraction and stripping, the recovery rate of vanadium is 99.1%, and the concentration of vanadium in the vanadium oxysulfate solution obtained by stripping is tested. The obtained vanadyl sulfate was evaporated and concentrated to 80g / L, and the impurity content was measured to be 35.9mg / L (the standard of the composite high-purity vanadyl sulfate solution: impurity ions are less than 40mg / L, and the vanadium ion concentration is greater than 76g / L). Compared with the traditional preparation of vanadyl sulfate from vanadium slag, the overall recovery rate of vanadium is greater than 98%, far exceeding the existing 90%.

[0058] Example 7

[0059] The difference between this implementation case and implementation case 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, and then the vanadium-containing organic phase is separated and washed with 25°C water for 10 minutes, with a phase ratio (O / A) of 1:5. The organic phase after water washing 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 above acid washing is subjected to sulfuric acid stripping, the sulfuric acid concentration is 2.5 mol / L, the phase ratio (O / A) is 7:1, and the stripping time is 20 minutes. After three-stage extraction and stripping, the recovery rate of vanadium is 99.3%, and the concentration of vanadium 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 the blank roasting product, see Example 8.

[0061] Example 8

[0062] (1) directly roasting vanadium slag without additives to obtain blank roasted clinker, crushing the blank roasted product of vanadium slag and grinding it to a particle size of less than 74 μm, adding water at a liquid-solid ratio of water to the roasted product of 5 mL / g for leaching, during which sulfuric acid is used to adjust the pH so that the pH of the leachate is always maintained 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, which is blue at this time, and the vanadium leaching rate is calculated to be 85.47%;

[0063] (2) taking 50 mL of vanadium-containing leaching solution, adjusting the pH of the system to 5.0 with ammonia water, and performing vanadium precipitation reaction at 30° C. for 10 min;

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

[0065] (4) The obtained vanadium hydroxide 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 97.65%.

[0066] Control group experiment: The difference from Example 8 is that no reducing agent is added during leaching. At this time, 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 the direct acid leaching experiment on the blank roasted product, the leaching rate of vanadium by the method of the present invention for the roasting without additives was investigated. The leaching rate of the method of the present invention (Example 8) was 85.47%, which was increased by about 75% compared with 10.31% (pH was 1.5), and was increased by about 35% compared with the conventional leaching rate of 50.24% (pH was 2.8). It shows that the method of the present invention has a high conversion rate for vanadium in the blank roasted product of vanadium slag.

[0068] Comparative Example 1

[0069] The vanadium slag calcification roasted product is crushed and ground to about 200 meshes, that is, to a particle size of less than 74μm, and water is added to the roasted product at a liquid-solid ratio of 10 mL / g for leaching. During the process, sulfuric acid is used to adjust the pH, and the pH of the leachate is always maintained at 2.5-2.8. The leaching temperature is 60°C and the leaching time is 60min. After the leaching is completed, the solid and liquid are filtered to separate and obtain a vanadium-containing leachate. At this time, the vanadium-containing leachate is yellow, and the calculated vanadium leaching rate is 87.15%; the obtained vanadium-containing leachate is subjected to ammonia salt precipitation and calcination to obtain V2O5, and the total recovery rate of vanadium is 85.07%. Compared with Example 1, the reduction leaching process is about 11% higher than the traditional non-reduction leaching rate. The total recovery rate of vanadium is increased by about 10%, and the V2O5 obtained by the traditional process is VO2.

[0070] Comparative Example 2

[0071] The difference between this comparative example 2 and example 1 is that in step (1), the liquid-solid ratio of the leaching medium to the roasted product is 2:1. The obtained vanadium leaching rate is 91.13%, and the vanadium leaching efficiency is lower than that of example 1. When the liquid-solid ratio of the leaching medium to the roasted product is 20:1, the vanadium leaching efficiency is 98.21%, which is slightly higher than that of example 1. However, if the liquid-solid ratio is further increased, the vanadium leaching efficiency is not significantly improved. In terms of economy and cost, the present application preferably uses a liquid-solid ratio of the leaching medium to the roasted product of less than 3 to 20:1.

[0072] Comparative Example 3

[0073] The difference between this comparative example 3 and Example 1 is that the temperature of the synchronous reduction leaching process in step (1) is 5°C. The leaching efficiency of vanadium is 90.21%. Compared with Example 1, the vanadium leaching rate is reduced by about 8%. When the leaching temperature is higher than 100°C, it is difficult to implement in industry, and further increasing the temperature will not bring much benefit. Therefore, the preferred temperature of the synchronous reduction leaching process in this application is 10°C to 100°C.

[0074] In summary, the vanadium conversion rate of the method of the present invention is 95.24% to 98.46%, and the vanadium recovery rate is 97.65% to 99.3%, which is much higher than the existing vanadium extraction method and comparative examples 1 to 3. It effectively solves the problem of easy hydrolysis of pentavalent vanadium ions in the vanadium slag vanadium extraction process, and improves the conversion rate and recovery rate of vanadium. Moreover, the VO2 obtained by the present invention can be used in the fields of smart windows, infrared stealth and supercapacitors, the obtained vanadium oxysulfate can be used in the field of energy storage, and vanadium pentoxide can be used in the field of steel. The vanadium product obtained by the present invention has multiple ways and high value utilization.

[0075] The preferred embodiments of the present invention are described in detail above, but the present invention is not limited thereto. Within the technical concept of the present invention, the present invention can make a variety of simple modifications to the technical solution of the present invention, including combining various technical features in any other suitable manner, and these simple modifications and combinations should also be regarded as the contents disclosed by the present invention and belong to the protection scope of the present invention.

Claims

1. A method for extracting vanadium from a vanadium slag roasting product, characterized in that: The method comprises the following steps: crushing and grinding the vanadium slag roasting product, performing synchronous reduction leaching in a leaching medium, adjusting the pH of the leaching solution to 0-5 with a pH regulator, and the leaching medium and the roasting product have a liquid-to-solid ratio of 3-20:1 to obtain a leaching solution containing positive divalent vanadium oxide ions.

2. The method according to claim 1, characterized in that The leaching medium is a reducing agent and water, wherein the molar ratio of the reducing agent to vanadium in the roasted clinker is 1-2:1; the temperature of the synchronous reduction leaching is 10-100°C, and the time is 10-120 minutes.

3. The method according to claim 1, characterized in that: The vanadium slag can be one or more of ordinary vanadium slag, vanadium-chromium slag or high-calcium and high-phosphorus vanadium slag.

4. The method according to claim 1, characterized in that: 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 therein; the roasting product is ground to a particle size of less than 74 μm.

5. The method according to claim 1, characterized in that The pH is 0-2, and the pH adjuster is one or more of sulfuric acid, hydrochloric acid, nitric acid, citric acid, oxalic acid, and ascorbic acid.

6. A method for extracting vanadium from vanadium slag roasting products and preparing vanadium oxides, characterized in that: It includes the following steps: S1. Obtaining a leaching solution containing positive divalent vanadium oxide ions by the vanadium extraction method of claim 1; S2, adjusting the pH of the leachate obtained in step S1 to 4-8 with an alkaline substance, performing a precipitation reaction, the reaction temperature is 10°C-100°C, the reaction time is 10min-120min, and the reaction is cooled to room temperature after completion; S3, separating the product obtained in step S2 from solid and liquid, and repeatedly washing and precipitating it, and then drying it to obtain vanadium hydroxide; S4, calcining the vanadium oxyhydroxide obtained in step S3 in an inert atmosphere at 300° C. to 500° C. for 2-3 hours to obtain vanadium dioxide, or heating in an air atmosphere at 500° C. to 670° C. to obtain vanadium pentoxide.

7. The method according to claim 6, 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 washing 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.

8. The method according to claim 6, characterized in that 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 atmosphere is one or both of air and oxygen.

9. A method for extracting vanadium from a vanadium slag roasting product and preparing vanadyl sulfate, characterized in that: It includes the following steps: S1. Obtaining a leaching solution containing positive divalent vanadium oxide ions by the vanadium extraction method of claim 1; 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, and the obtained second organic phase is stripped with sulfuric acid to obtain a vanadyl sulfate solution.

10. The method according to claim 9, characterized in that The extractant is P5O7, P2O4 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.